Trilogy Metals (TMQ) 8-K/A Regulation FD Disclosure

T A B L E  O F  C O N T E N T S

1.0	SUMMARY			1-1
	1.1	INTRODUCTION		1-1
	1.2	PROPERTY DESCRIPTION AND LOCATION		1-3
	1.3	GEOLOGY AND MINERALIZATION		1-3
	1.4	RESOURCE ESTIMATE		1-5
	1.5	MINERAL PROCESSING AND METALLURGICAL TESTING		1-6
	1.6	MINING METHODS		1-7
	1.7	RECOVERY METHODS		1-7
	1.8	PROJECT INFRASTRUCTURE		1-8
		1.8.1	POWER SUPPLY	1-10
		1.8.2	TAILINGS STORAGE FACILITY	1-11
	1.9	ENVIRONMENTAL		1-11
	1.10	CAPITAL COSTS (INITIAL AND SUSTAINING)		1-12
	1.11	OPERATING COSTS		1-13
	1.12	ECONOMIC ANALYSIS		1-14
	1.13	INTERPRETATION AND CONCLUSIONS		1-15
	1.14	OPPORTUNITIES AND RECOMMENDATIONS		1-15
2.0	INTRODUCTION			2-1
	2.1	TERMS OF REFERENCE		2-1
		2.1.1	UNITS OF MEASUREMENT	2-1
	2.2	QUALIFIED PERSONS		2-2
	2.3	SITE VISITS		2-4
	2.4	SCOPE OF PERSONAL INSPECTIONS		2-4
	2.5	INFORMATION SOURCES		2-4
3.0	RELIANCE ON OTHER EXPERTS			3-1
	3.1	MICHAEL F. O’BRIEN, M.SC., PR.SCI.NAT, FGSSA, FAUSIMM, FSAIMM		3-1
	3.2	SABRY ABDEL HAFEZ, PH.D., P.ENG		3-1
4.0	PROPERTY DESCRIPTION AND LOCATION			4-1
	4.1	LOCATION		4-1
	4.2	MINERAL TENURE		4-4
	4.3	ROYALTIES, AGREEMENTS AND ENCUMBRANCES		4-6
		4.3.1	KENNECOTT AGREEMENTS	4-6
		4.3.2	NANA AGREEMENT	4-6
	4.4	ENVIRONMENTAL LIABILITIES		4-8

iv

	4.5	PERMITS		4-8
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			5-1
	5.1	ACCESSIBILITY		5-1
		5.1.1	AIR	5-1
		5.1.2	WATER	5-1
		5.1.3	ROAD	5-1
	5.2	CLIMATE		5-1
	5.3	LOCAL RESOURCES		5-2
	5.4	INFRASTRUCTURE		5-3
	5.5	PHYSIOGRAPHY		5-3
6.0	HISTORY			6-1
	6.1	PRIOR OWNERSHIP AND OWNERSHIP CHANGES – ARCTIC DEPOSIT AND THE AMBLER LANDS
				6-2
	6.2	PREVIOUS EXPLORATION AND DEVELOPMENT RESULTS – ARCTIC DEPOSIT		6-3
		6.2.1	INTRODUCTION	6-3
		6.2.2	GEOCHEMISTRY	6-8
		6.2.3	GEOPHYSICS	6-8
		6.2.4	DRILLING	6-9
		6.2.5	SPECIFIC GRAVITY	6-9
		6.2.6	PETROLOGY, MINERALOGY, AND RESEARCH STUDIES	6-9
		6.2.7	GEOTECHNICAL, HYDROLOGICAL AND ACID-BASE ACCOUNTING STUDIES	6-10
		6.2.8	METALLURGICAL STUDIES	6-11
	6.3	HISTORICAL MINERAL RESOURCE ESTIMATES		6-11
		6.3.1	RUSSELL – KENNECOTT (1975/1976) RESOURCE ESTIMATE	6-11
		6.3.2	BROWN – KENNECOTT (1985) RESOURCE ESTIMATE	6-11
		6.3.3	RANDOLF – KENNECOTT (1990) RESOURCE ESTIMATE	6-12
		6.3.4	KENNECOTT (1995) RESOURCE ESTIMATE	6-12
		6.3.5	DEAD CREEK, SUNSHINE AND HORSE CLIFF HISTORICAL RESOURCES	6-13
	6.4	DEVELOPMENT STUDIES		6-13
		6.4.1	KENNECOTT TENURE	6-13
7.0	GEOLOGICAL SETTING AND MINERALIZATION			7-1
	7.1	REGIONAL GEOLOGY – SOUTHERN BROOKS RANGE		7-1
		7.1.1	TERRANE DESCRIPTIONS	7-1
		7.1.2	REGIONAL TECTONIC SETTING	7-3
	7.2	AMBLER SEQUENCE GEOLOGY		7-4
		7.2.1	GENERAL STRATIGRAPHY OF THE AMBLER SEQUENCE	7-5
		7.2.2	STRUCTURAL FRAMEWORK OF THE AMBLER DISTRICT	7-9
	7.3	ARCTIC DEPOSIT GEOLOGY		7-10
		7.3.1	LITHOLOGIES AND LITHOLOGIC DOMAIN DESCRIPTIONS	7-11
		7.3.2	STRUCTURE	7-13
		7.3.3	ALTERATION	7-14
	7.4	ARCTIC DEPOSIT MINERALIZATION		7-15
	7.5	GENESIS		7-16

v

	7.6	DEPOSITS AND PROSPECTS		7-17
8.0	DEPOSIT TYPES			8-1
9.0	EXPLORATION			9-1
	9.1	GRIDS AND SURVEYS		9-2
	9.2	GEOLOGICAL MAPPING		9-2
	9.3	GEOCHEMISTRY		9-4
	9.4	GEOPHYSICS		9-7
	9.5	BULK DENSITY		9-9
	9.6	PETROLOGY, MINERALOGY AND RESEARCH STUDIES		9-9
	9.7	GEOTECHNICAL, HYDROGEOLOGICAL AND ACID BASE ACCOUNTING STUDIES		9-9
		9.7.1	GEOTECHNICAL STUDIES	9-9
		9.7.2	HYDROLOGICAL STUDIES	9-9
		9.7.3	ACID-BASE ACCOUNTING STUDIES	9-10
		9.7.4	METALLURGICAL STUDIES	9-11
10.0	DRILLING			10-1
	10.1	DRILL COMPANIES		10-5
	10.2	DRILL CORE PROCEDURES		10-6
		10.2.1	KENNECOTT TENURE	10-6
		10.2.2	NOVAGOLD/NOVACOPPER TENURE	10-6
	10.3	GEOTECHNICAL DRILL HOLE PROCEDURES		10-7
	10.4	METALLURGICAL DRILL HOLE PROCEDURES		10-8
	10.5	COLLAR SURVEYS		10-9
		10.5.1	KENNECOTT TENURE	10-9
		10.5.2	NOVAGOLD/NOVACOPPER TENURE	10-9
	10.6	DOWNHOLE SURVEYS		10-9
	10.7	RECOVERY		10-10
	10.8	DRILL INTERCEPTS		10-10
	10.9	DRILLING AT OTHER PROSPECTS		10-11
11.0	SAMPLE PREPARATION, ANALYSES, AND SECURITY			11-1
	11.1	SAMPLE PREPARATION		11-1
		11.1.1	CORE DRILLING SAMPLING	11-1
		11.1.2	ACID-BASE ACCOUNTING SAMPLING	11-3
		11.1.3	DENSITY DETERMINATIONS	11-3
	11.2	SECURITY		11-4
	11.3	ASSAYING AND ANALYTICAL PROCEDURES		11-4
	11.4	QUALITY ASSURANCE/QUALITY CONTROL		11-5
		11.4.1	CORE DRILLING SAMPLING QA/QC	11-5
		11.4.2	ACID-BASE ACCOUNTING SAMPLING QA/QC	11-12
		11.4.3	DENSITY DETERMINATIONS QA/QC	11-12
	11.5	AUTHOR’S OPINION		11-15
12.0	DATA VERIFICATION			12-1

vi

	12.1	DRILL HOLE COLLAR VERIFICATION		12-1
	12.2	TOPOGRAPHY VERIFICATION		12-1
	12.3	CORE LOGGING VERIFICATION		12-2
	12.4	DATABASE VERIFICATION		12-2
	12.5	QA/QC REVIEW		12-3
	12.6	QP OPINION		12-4
13.0	MINERAL PROCESSING AND METALLURGICAL TESTING			13-1
	13.1	METALLURGICAL TEST WORK REVIEW		13-1
		13.1.1	INTRODUCTION	13-1
		13.1.2	MINERALOGICAL AND METALLURGICAL TEST WORK – SGS 2012	13-1
		13.1.3	HISTORICAL TEST WORK REVIEW	13-14
	13.2	RECOMMENDED TEST WORK		13-19
14.0	MINERAL RESOURCE ESTIMATES			14-1
	14.1	INTRODUCTION		14-1
	14.2	PREVIOUS RESOURCE ESTIMATES		14-2
		14.2.1	KENNECOTT (1990) RESOURCE ESTIMATE	14-2
		14.2.2	SRK (2008) RESOURCE ESTIMATION	14-2
		14.2.3	SRK (2011) RESOURCE ESTIMATION UPDATE	14-5
		14.2.4	SRK (2012) RESOURCE ESTIMATION UPDATE	14-6
	14.3	RESOURCE ESTIMATION PROCEDURES		14-7
	14.4	DATABASE		14-7
	14.5	GEOLOGICAL MODELLING		14-8
		14.5.1	MINERALIZATION MODEL	14-8
		14.5.2	SPECIFIC GRAVITY MODEL	14-8
		14.5.3	LITHOGEOCHEMICAL MODEL	14-8
		14.5.4	ESTIMATION DOMAINS	14-9
	14.6	EXPLORATORY DATA ANALYSIS		14-11
		14.6.1	CONTACT PROFILES	14-11
		14.6.2	RAW ASSAY DATA AND STATISTICS	14-12
	14.7	COMPOSITING		14-13
		14.7.1	OUTLIER MANAGEMENT AND CAPPING STRATEGY	14-13
		14.7.2	COMPOSITE STATISTICS	14-15
	14.8	SPECIFIC GRAVITY ANALYSIS		14-16
		14.8.1	SPECIFIC GRAVITY STATISTICS AND SPATIAL ANALYSIS	14-16
		14.8.2	SPECIFIC GRAVITY INTERPOLATION PLAN	14-16
	14.9	ACID BASE ACCOUNTING ANALYSIS		14-17
	14.10	VARIOGRAPHY AND SPATIAL ANALYSIS		14-18
	14.11	RESOURCE ESTIMATION METHODOLOGY		14-20
	14.12	RESOURCE BLOCK MODEL		14-22
		14.12.1	CONFIGURATION	14-22
		14.12.2	CELL ATTRIBUTES	14-22
	14.13	MODEL VALIDATION		14-25
		14.13.1	VISUAL VALIDATION	14-25

vii

		14.13.2	STATISTICAL COMPARISONS BETWEEN BLOCK AND COMPOSITE DATA	14-28
	14.14	MINERAL RESOURCE CLASSIFICATION		14-37
		14.14.1	MEASURED RESOURCE CLASSIFICATION	14-38
		14.14.2	INDICATED RESOURCE CLASSIFICATION	14-38
		14.14.3	INFERRED RESOURCE CLASSIFICATION	14-38
	14.15	MINERAL RESOURCE ESTIMATE		14-38
	14.16	GRADE SENSITIVITY ANALYSIS		14-40
15.0	MINERAL RESERVE ESTIMATES			15-1
16.0	MINING METHODS			16-1
	16.1	INTRODUCTION		16-1
	16.2	PIT OPTIMIZATION		16-1
		16.2.1	BLOCK MODEL	16-1
		16.2.2	PIT SLOPE ANGLE	16-1
		16.2.3	SURFACE TOPOGRAPHY	16-1
		16.2.4	PIT OPTIMIZATION PARAMETERS	16-1
		16.2.5	PIT OPTIMIZATION RESULTS	16-5
	16.3	MINE DESIGN		16-7
		16.3.1	BENCH HEIGHT AND PIT WALL SLOPE	16-7
		16.3.2	MINIMUM WORKING AREA	16-8
		16.3.3	HAUL ROAD	16-8
		16.3.4	PIT HYDROLOGY/DEWATERING	16-9
		16.3.5	PIT DESIGN RESULTS	16-9
	16.4	PRODUCTION SCHEDULE		16-10
	16.5	MINE WASTE ROCK MANAGEMENT		16-18
	16.6	MINING EQUIPMENT		16-18
		16.6.1	MINE EQUIPMENT FLEET	16-18
		16.6.2	OPERATING HOURS	16-18
		16.6.3	PRIMARY EQUIPMENT	16-18
		16.6.4	SUPPORT AND ANCILLARY EQUIPMENT	16-19
	16.7	MINING LABOUR		16-20
17.0	RECOVERY METHODS			17-1
	17.1	MINERAL PROCESSING		17-1
		17.1.1	FLOWSHEET DEVELOPMENT	17-1
		17.1.2	PROCESS PLANT DESCRIPTION	17-4
		17.1.3	COARSE MATERIAL STORAGE	17-5
		17.1.4	GRINDING AND CLASSIFICATION	17-5
		17.1.5	FLOTATION	17-6
	17.2	PLANT PROCESS CONTROL		17-15
		17.2.1	OVERVIEW	17-15
		17.2.2	PRODUCTION PROJECTION	17-16
18.0	PROJECT INFRASTRUCTURE			18-1
	18.1	OVERVIEW		18-1
	18.2	ROADS AND AIRSTRIP		18-4

viii

		18.2.1	AMBLER MINING DISTRICT INDUSTRIAL ACCESS ROAD	18-4
		18.2.2	ACCESS ROAD	18-4
		18.2.3	HAUL ROADS, SITE ROADS, AND PAD AREAS	18-5
		18.2.4	AIRSTRIP	18-6
	18.3	BUILDINGS AND RELEVANT BUILDING SERVICES		18-7
		18.3.1	PRIMARY CRUSHING BUILDING	18-7
		18.3.2	CRUSHED MATERIAL STOCKPILE	18-7
		18.3.3	CONVEYING	18-7
		18.3.4	MILL BUILDING, MAINTENANCE, TRUCK SHOP, ASSAY AND METALLURGY LAB18-7
		18.3.5	ADMINISTRATION AND MINE DRY	18-8
		18.3.6	ARCTIC CORRIDORS	18-8
		18.3.7	COLD STORAGE WAREHOUSE	18-8
		18.3.8	HVAC AND FIRE PROTECTION	18-8
		18.3.9	FIRE PROTECTION	18-9
		18.3.10	DUST CONTROL	18-9
	18.4	WATER SUPPLY AND DISTRIBUTION		18-9
	18.5	WASTE MANAGEMENT		18-10
		18.5.1	SEWAGE DISPOSAL	18-10
		18.5.2	DOMESTIC WASTE DISPOSAL	18-10
	18.6	FUEL STORAGE		18-10
	18.7	ON SITE EXPLOSIVES STORAGE		18-11
		18.7.1	EMULSION PLANT	18-11
		18.7.2	DETONATOR AND EXPLOSIVE STORAGE MAGAZINE	18-11
	18.8	POWER SUPPLY TO PLANT SITE		18-11
	18.9	TAILINGS STORAGE FACILITY		18-11
		18.9.1	TSF EMBANKMENT	18-12
		18.9.2	TAILINGS IMPOUNDMENT	18-13
		18.9.3	RECLAIM WATER SYSTEM	18-13
		18.9.4	SEEPAGE COLLECTION SUMP	18-14
	18.10	WASTE ROCK FACILITIES		18-14
		18.10.1	NON-ACID GENERATING WASTE ROCK	18-14
		18.10.2	POTENTIALLY ACID GENERATING WASTE ROCK	18-14
	18.11	WATER MANAGEMENT		18-14
		18.11.1	SITE FACILITIES	18-16
	18.12	WATER TREATMENT PLANT		18-18
	18.13	CONSTRUCTION AND PERMANENT CAMP ACCOMMODATION		18-18
	18.14	COMMUNICATIONS		18-18
19.0	MARKET STUDIES AND CONTRACTS			19-1
20.0	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT			20-1
	20.1	ENVIRONMENTAL STUDIES		20-1
		20.1.1	HYDROLOGY AND WATER QUALITY DATA	20-1
		20.1.2	WETLANDS DATA	20-2
		20.1.3	AQUATIC LIFE DATA	20-2
		20.1.4	SUBSISTENCE DATA	20-4

ix

		20.1.5	ACID BASE ACCOUNTING DATA	20-5
		20.1.6	ADDITIONAL BASELINE DATA REQUIREMENTS	20-6
	20.2	PERMITTING		20-6
		20.2.1	EXPLORATION PERMITS	20-6
		20.2.2	MAJOR MINE PERMITS	20-7
	20.3	SOCIAL OR COMMUNITY CONSIDERATIONS		20-9
	20.4	MINE RECLAMATION AND CLOSURE		20-10
		20.4.1	RECLAMATION AND CLOSURE PLAN	20-11
		20.4.2	RECLAMATION AND CLOSURE FINANCIAL ASSURANCE	20-12
21.0	CAPITAL AND OPERATING COSTS			21-1
	21.1	INTRODUCTION		21-1
	21.2	CAPITAL COST ESTIMATE		21-1
		21.2.1	INTRODUCTION	21-1
		21.2.2	CAPITAL COST SUMMARY	21-1
		21.2.3	TOTAL SUSTAINING CAPITAL COSTS	21-1
		21.2.4	PURPOSE AND CLASS OF ESTIMATE	21-2
		21.2.5	ESTIMATE BASE DATE, EXCHANGE RATE, AND VALIDITY PERIOD	21-2
		21.2.6	CONTRIBUTORS TO THE ESTIMATE	21-3
	21.3	ELEMENTS OF COST		21-3
		21.3.1	DIRECT COSTS	21-3
		21.3.2	INDIRECT COSTS	21-4
		21.3.3	OWNER’S COSTS	21-4
		21.3.4	CONTINGENCIES	21-5
		21.3.5	EXCLUSIONS	21-5
		21.3.6	COSTS INCURRED PRIOR TO RELEASE OF DETAIL ENGINEERING AND CONSTRUCTION
			ASSUMPTIONS	21-5
	21.4	MINING CAPITAL COST ESTIMATE		21-6
	21.5	OPERATING COST ESTIMATE		21-9
		21.5.1	SUMMARY	21-9
		21.5.2	MINING OPERATING COST ESTIMATE	21-10
		21.5.3	PROCESSING OPERATING COSTS	21-11
		21.5.4	GENERAL AND ADMINISTRATIVE COSTS AND SURFACE SERVICES COSTS	21-13
		21.5.5	TAILINGS STORAGE FACILITY COST	21-15
		21.5.6	ROAD TOLL COST	21-15
22.0	ECONOMIC ANALYSIS			22-1
	22.1	PRE-TAX MODEL		22-2
		22.1.1	MINE/METAL PRODUCTION IN FINANCIAL MODEL	22-2
		22.1.2	BASIS OF FINANCIAL EVALUATION	22-2
	22.2	SUMMARY OF FINANCIAL RESULTS		22-4
	22.3	SENSITIVITY ANALYSIS		22-5
	22.4	POST-TAX FINANCIAL ANALYSIS		22-7
		22.4.1	US FEDERAL TAX	22-7
		22.4.2	ALASKA STATE TAX	22-8
		22.4.3	ALASKA MINING LICENSE TAX	22-8
		22.4.4	TAXES AND POST-TAX RESULTS	22-8

x

	22.5	ROYALTIES		22-9
	22.6	SMELTER TERMS		22-10
	22.7	TRANSPORTATION LOGISTICS		22-11
	22.8	INSURANCE		22-11
	22.9	REPRESENTATION AND MARKETING		22-11
23.0	ADJACENT PROPERTIES			23-1
	23.1	SUN DEPOSIT		23-1
	23.2	SMUCKER DEPOSIT		23-3
24.0	OTHER RELEVANT DATA AND INFORMATION			24-1
	24.1	BORNITE DEPOSIT		24-1
25.0	INTERPRETATION AND CONCLUSIONS			25-1
	25.1	GENERAL		25-1
	25.2	PROPERTY DESCRIPTION AND LOCATION		25-1
	25.3	GEOLOGY		25-1
	25.4	MINERAL RESOURCE ESTIMATION		25-2
	25.5	METALLURGICAL TEST WORK AND PROCESS DESIGN		25-2
	25.6	MINING METHODS		25-4
	25.7	PROJECT INFRASTRUCTURE		25-4
	25.8	TAILINGS AND WATER MANAGEMENT		25-5
	25.9	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT		25-5
	25.10	ECONOMIC ANALYSIS		25-5
26.0	RECOMMENDATIONS			26-1
	26.1	GENERAL		26-1
	26.2	GEOLOGY		26-1
	26.3	MINERAL PROCESSING AND METALLURGICAL TESTING		26-2
	26.4	PROCESS PLANT DESIGN		26-2
	26.5	MINING METHODS		26-3
	26.6	PROJECT INFRASTRUCTURE		26-3
		26.6.1	PROCESS PLANT AND INFRASTRUCTURE LOCATION	26-3
		26.6.2	ACCESS ROAD	26-4
		26.6.3	ELECTRICAL POWER GENERATION	26-4
		26.6.4	AIRSTRIP	26-4
		26.6.5	TAILINGS AND WASTE MANAGEMENT	26-4
	26.7	ENVIRONMENTAL		26-6
	26.8	ECONOMIC ANALYSIS		26-6
27.0	REFERENCES			27-1
	27.1	GEOLOGY		27-1
	27.2	ENVIRONMENTAL		27-5
	27.3	MINING		27-6

xi

	27.4	CAPITAL COST ESTIMATE	27-6
	27.5	ECONOMIC ANALYSIS	27-6
	27.6	ELECTRICAL POWER GENERATION	27-6
28.0	CERTIFICATES OF QUALIFIED PERSONS		28-1
	28.1	MICHAEL F. O’BRIEN, M.SC., PR.SCI.NAT, FGSSA, FAUSIMM, FSAIMM	28-1
	28.2	JIANHUI (JOHN) HUANG, PH.D., P.ENG	28-3
	28.3	SABRY ABDEL HAFEZ, PH.D., P.ENG	28-4
	28.4	HASSAN GHAFFARI, P.ENG	28-5
	28.5	HARVEY WAYNE STOYKO, P.ENG	28-6
	28.6	MICHAEL CHIN, P.ENG	28-7
	28.7	GRAHAM WILKINS, P.ENG	28-8
	28.8	MARVIN SILVA, PH.D., PE, P.ENG.	28-9
	28.9	JACK DIMARCHI, CPG	28-10

LISTOFAPPENDICES

APPENDIX A	CLAIMS MAP AND LIST OF CLAIMS
APPENDIX B	ACCESS ROAD
APPENDIX C	AIRSTRIP
APPENDIX D	TSF AND WATER MANAGEMENT
APPENDIX E	CAPITAL COST BASIS OF ESTIMATE

LIST OFTABLES

Table 1.1	General Project Information	1-3
Table 1.2	Resource Estimate for the Arctic Project (NSR Cut-off of $35/t)	1-6
Table 1.3	Summary of Key Mining Results	1-7
Table 1.4	Capital Cost Summary	1-13
Table 1.5	Summary of Total Sustaining Capital Cost for LOM	1-13
Table 1.6	Overall Operating Cost Estimate	1-14
Table 2.1	Qualified Persons	2-2
Table 6.1	Known Mapping, Geochemical, and Geophysical Programs Targeting VMS Prospects in the Ambler Mining District	6-4
Table 6.2	Russell 1976 Resource Estimation	6-11
Table 6.3	Brown 1985 Resource Estimation	6-12
Table 6.4	Randolf 1990 Resource Estimation	6-12

xii

Table 6.5	Kennecott 1995 Resource Estimation	6-12
Table 6.6	Historical Resources for the Dead Creek, Sunshine and Horse Cliff Prospects	6-13
Table 9.1	Summary of NovaCopper/NovaGold Exploration Activities Targeting VMS-style Mineralization in the Ambler Sequence Stratigraphy and the Arctic Deposit	9-1
Table 9.2	TDEM Loops and Locations	9-7
Table 10.1	Companies, Campaigns, Drill Holes and Metres Drilled at the Arctic Deposit	10-1
Table 10.2	Summary of NovaCopper/NovaGold Drilling	10-2
Table 10.3	Drill Contractors, Drill Holes, Metreage and Core Sizes by Drill Campaign at the Arctic Deposit	10-5
Table 10.4	Geotechnical Holes and Instrumentation	10-8
Table 10.5	Recovery and RQD 2004 to 2008 Arctic Drill Campaigns	10-10
Table 10.6	Drill Intercept Summary Table	10-11
Table 10.7	Drill, Metreage and Average Drill Depth for NovaCopper Ambler Sequence VMS Targets	10-11
Table 10.8	Significant Drill Intercepts – NovaCopper Ambler Sequence Prospects	10-12
Table 10.9	NovaCopper Exploration Drilling – Ambler Schist Belt	10-14
Table 11.1	Analytical Laboratories Used by Operators of the Arctic Project	11-5
Table 12.1	Assay Certificates Reviewed	12-2
Table 13.1	Metallurgical Test Work Programs	13-1
Table 13.2	Head Grades – Composite Samples – 2012	13-2
Table 13.3	Mineral Modal Abundance for Composite Samples – 2012	13-3
Table 13.4	Bond Ball Mill Grindability and Abrasion Index Test Results	13-4
Table 13.5	Locked Cycle Metallurgical Test Results	13-8
Table 13.6	Copper and Lead Separation Test Results	13-11
Table 13.7	Multi-element Assay Results – Copper Concentrate	13-13
Table 13.8	Multi-element Assay Results – Lead Concentrate	13-13
Table 13.9	Multi-element Assay Results – Zinc Concentrate	13-14
Table 13.10	Metallic Mineral Identified in Arctic Project Samples	13-14
Table 13.11	Bond Ball Mill Work Index	13-16
Table 13.12	Head Analyses	13-18
Table 13.13	Flotation Test on Ambler Low Talc Composite	13-19
Table 14.1	Historical Resource Estimate	14-2
Table 14.2	Drill Hole Assay Statistics – 2008 SRK Resource Estimation	14-3
Table 14.3	Historical SG Data Statistics – Arctic Deposit: 1998–2004	14-3
Table 14.4	SG Measurements Categorized by Rock Type	14-3
Table 14.5	Arctic Deposit Resources at $100 GMV Cut-off – 2008 SRK	14-4
Table 14.6	Arctic Deposit Resources at $75 NSR Cut-off – 2011 SRK	14-5
Table 14.7	Arctic Deposit Resources at $75 NSR Cut-off – 2012 SRK	14-6
Table 14.8	Exploration Data within the Resource Area	14-7
Table 14.9	Modifications to Database for Resource Estimation	14-7
Table 14.10	Arctic Grade Estimation Domains	14-9
Table 14.11	Arctic SG Estimation Domains	14-10
Table 14.12	Arctic ABA Domains	14-10
Table 14.13	Raw Metal Sample Statistics (Length Weighted)	14-13
Table 14.14	Sample Length Statistics	14-13
Table 14.15	Arctic Deposit Drill Hole Composite Statistics	14-15
Table 14.16	Arctic SG Determinations Summary Statistics	14-16
Table 14.17	Arctic Density Interpolation Summary Statistics	14-16
Table 14.18	PAG and NAG Ratio Assignment for the ABA Domains	14-17
Table 14.19	Variogram Models	14-18
Table 14.20	Search Ellipse Orientation and Dimensions for Mineralization Domains	14-21
Table 14.21	Block Model Cell Attributes	14-23

xiii

Table 14.22	Indicated Mineral Resource Table Restated by Mineralization Zone, Arctic Project, Alaska, Tetra Tech (June 30, 2013)	14-39
Table 14.23	Inferred Mineral Resource Table Restated by Mineralization Zone, Arctic Project, Alaska, Tetra Tech (June 30, 2013)	14-39
Table 14.24	Material Comprising the Indicated Mineral Resource (Table 14.22) Diluted to Full Blocks, Restated by Mineralization Zone, Arctic Project, Alaska	14-40
Table 14.25	Material Comprising the Inferred Mineral Resource (Table 14.23) Diluted to Full Blocks, Restated by Mineralization Zone, Arctic Project, Alaska	14-40
Table 14.26	Indicated Mineral Resource Table Sensitivity to NSR Within Whole Blocks	14-41
Table 14.27	Inferred Mineral Resource Table Sensitivity to NSR, Arctic Project Within Whole Blocks	14-41
Table 16.1	Pit Optimization Parameters	16-3
Table 16.2	Metallurgical Recovery	16-5
Table 16.3	Pit Optimization Results	16-6
Table 16.4	Haul Road Width	16-8
Table 16.5	Pit Design Results	16-9
Table 16.6	Production Schedule	16-12
Table 16.7	Operational Delays per Shift	16-18
Table 16.8	Haulage Cycle Times	16-19
Table 16.9	Primary Equipment Requirements	16-19
Table 16.10	Support Equipment Requirements	16-19
Table 16.11	Ancillary Equipment Requirements	16-20
Table 16.12	Operator and Maintenance Staff on Payroll	16-21
Table 17.1	Major Design Criteria	17-4
Table 17.2	Annual Concentrate Production Projection	17-17
Table 18.1	Components for Management of Non-contact Surface Waters	18-16
Table 18.2	Components for Management of Contact Surface Waters and Sediments	18-16
Table 20.1	Captured or Observed Fish Species	20-3
Table 20.2	Additional Recommended Environmental Baseline Studies	20-6
Table 20.3	Major Mine Permits Required for the Arctic Project	20-8
Table 21.1	Summary of Capital and Operating Costs	21-1
Table 21.2	CAPEX Summary	21-2
Table 21.3	Summary of Total Sustainable Capital Costs for LOM	21-2
Table 21.4	Foreign Exchange Rate Summary	21-3
Table 21.5	Summary of Mining Costs	21-6
Table 21.6	Mining Pre-production Unit Costs	21-7
Table 21.7	Initial Mining Equipment Fleet	21-7
Table 21.8	Sustaining Capital Mining Equipment Fleet	21-8
Table 21.9	Overall Operating Cost Estimate	21-10
Table 21.10	Relevant Consumables Prices	21-10
Table 21.11	Mining LOM and Unit Operating Cost Summary	21-11
Table 21.12	Summary of Process Operating Cost	21-12
Table 21.13	G&A Operating Costs	21-13
Table 21.14	Surface Services Operating Costs	21-14
Table 22.1	Mine/Metal Production from the Arctic Mine	22-2
Table 22.2	Summary of Pre-tax Financial Results	22-4
Table 22.3	Components of the Various Taxes	22-9
Table 22.4	Summary of Post-tax Financial Results	22-9
Table 23.1	Sun Deposit Historical Resources	23-2
Table 23.2	Smucker Deposit Historical Resources	23-4
Table 24.1	Bornite NI 43-101 Resources	24-1
Table 25.1	Resource estimate for the Arctic Project (NSR Cut-off of $35/t)	25-2

xiv

Table 25.2 Projected Recoveries and Concentrate Grades 25-3

LISTOFFIGURES

Figure 1.1	Property Location Map	1-2
Figure 1.2	Arctic Project Site Layout	1-10
Figure 4.1	Location Map of the UKMP – Northwest Alaska	4-2
Figure 4.2	Upper Kobuk Mineral Projects Lands	4-3
Figure 4.3	Mineral Tenure Plan	4-5
Figure 7.1	Geologic Terranes of the Southern Brooks Range	7-2
Figure 7.2	Geology of the Ambler Mining District	7-5
Figure 7.3	Ambler Sequence Stratigraphy in the Arctic Deposit Area	7-7
Figure 7.4	Generalized Geology of the Central Ambler District	7-8
Figure 7.5	Typical F1 Isoclinal Folds Developed in Calcareous Gnurgle Gneiss	7-9
Figure 7.6	Generalized Geologic Map of the Arctic Deposit	7-11
Figure 7.7	Typical Massive Sulphide Mineralization at the Arctic Deposit	7-16
Figure 7.8	Prospects of the Ambler Mining District	7-18
Figure 9.1	Mapping Campaigns in and around the Arctic Deposit	9-3
Figure 9.2	Arctic Deposit Area Geology	9-4
Figure 9.3	Copper Distribution in Silt and Soil Samples in the Dead Creek Area	9-5
Figure 9.4	Zinc Distribution in Silt and Soil Samples in the Arctic Deposit Area	9-6
Figure 9.5	TDEM Loops and Contoured Resistivity – Dead Creek Prospect	9-8
Figure 9.6	PAG versus NAG by Rock Type at the Arctic Deposit	9-11
Figure 10.1	Plan Map of Drill Holes Utilized in the Mineral Resource Estimation	10-3
Figure 10.2	Drill Holes Utilized for Metallurgical, Geotechnical, Hydrological and Acid-Base Accounting Studies	10-4
Figure 10.3	Known Collar Locations and Principal Target Areas – Ambler District	10-13
Figure 10.4	Sunshine Prospect and Drill Hole Locations	10-15
Figure 11.1	Spatial Availability of QA/QC Data	11-6
Figure 11.2	Graph Showing Good Agreement between Wet-dry Measured Specific Gravity and Pycnometer Measured Specific Gravity	11-13
Figure 11.3	Measured versus Stoichiometric Specific Gravities	11-14
Figure 11.4	Scatter Plot Showing the Measured Specific Gravity versus Multiple (Copper, Iron, Zinc, Barium) Regression Estimate	11-15
Figure 12.1	Distribution of the Differences Between GPS Elevations and the DTM	12-1
Figure 14.1	Cross-section Illustrating the Arctic Deposit Geological Model	14-9
Figure 14.2	Contact Profile between Zone 1 SMS and Country Rock for Copper	14-11
Figure 14.3	Contact Profile between Zone 1 SMS and MS for Zinc	14-12
Figure 14.4	Lead Histogram	14-14
Figure 14.5	Gold Histogram	14-15
Figure 14.6	Copper: Variography in SMS Domains	14-18
Figure 14.7	Zinc: Variography in SMS Domains	14-19
Figure 14.8	Copper: Variography in MS Domains	14-19
Figure 14.9	Zinc: Variography in MS Domains	14-20
Figure 14.10	Cross-section for Copper	14-25
Figure 14.11	Cross-section for Zinc	14-26
Figure 14.12	Cross-section for Lead	14-27

xv

Figure 14.13	Cross-section for Gold	14-27
Figure 14.14	Cross-section for Silver	14-28
Figure 14.15	Swath Plot of Block Copper Grade Values by Northing for the Arctic Deposit	14-29
Figure 14.16	Swath Plot of Block Copper Grade Values by Easting for the Arctic Deposit	14-29
Figure 14.17	Swath Plot of Block Zinc Grade Values by Northing for the Arctic Deposit	14-30
Figure 14.18	Swath Plot of Block Zinc Grade Values by Easting for the Arctic Deposit	14-31
Figure 14.19	Copper: Scatter Plot of OK versus ID Estimates	14-32
Figure 14.20	Zinc: Scatterplot of OK versus ID Estimates	14-32
Figure 14.21	Lead: Scatterplot of OK versus ID Estimates	14-33
Figure 14.22	Gold: Scatterplot of OK versus ID Estimates	14-33
Figure 14.23	Silver: Scatterplot of OK versus ID Estimates	14-34
Figure 14.24	Block Model Grade Tonnage Comparison: Copper OK (red line) and Copper ID2(broken line)	14-35
Figure 14.25	Block Model Grade Tonnage Comparison: Zinc OK (red line) and Zinc ID2 (broken line)	14-35
Figure 14.26	Block Model Grade Tonnage Comparison: Lead OK (red line) and Lead ID2 (broken line)	14-36
Figure 14.27	Block Model Grade Tonnage Comparison: Gold OK (red line) and Gold ID2 (broken line)	14-36
Figure 14.28	Block Model Grade Tonnage Comparison: Silver OK (red line) and Silver ID2 (broken line)	14-37
Figure 16.1	Arctic Project Topography	16-2
Figure 16.2	Pit Wall Slope	16-7
Figure 16.3	One-way Haul Road	16-8
Figure 16.4	Two-way Haul Road	16-9
Figure 16.5	Pit Design	16-10
Figure 16.6	Production Schedule	16-13
Figure 16.7	Pre-production Year Mine Status Map	16-14
Figure 16.8	Year 4 Mine Status Map	16-15
Figure 16.9	Year 8 Mine Status Map	16-16
Figure 16.10	Year 12 Mine Status Map	16-17
Figure 17.1	Simplified Process Flow Diagram	17-3
Figure 18.1	Final General Arrangement Layout after 12 years of Operation	18-3
Figure 18.2	Access to the Arctic Project	18-4
Figure 18.3	Tailings Storage Facility	18-12
Figure 18.4	Tailings Embankment Cross Section	18-13
Figure 20.1	Comparison of Benthic Macroinvertebrate Community Measures	20-4
Figure 21.1	LOM Average Operating Cost Distribution	21-10
Figure 22.1	Pre-tax Undiscounted Annual and Cumulative NCF	22-4
Figure 22.2	Pre-tax NPV Sensitivity Analysis	22-6
Figure 22.3	Pre-tax IRR Sensitivity Analysis	22-6
Figure 22.4	Pre-tax Payback Period Sensitivity Analysis	22-7
Figure 23.1	Adjacent Properties and Land Status	23-1
Figure 23.2	Sun Project Prospect Location Map	23-3

xvi

GLOSSARY

UNITS OFMEASURE
above mean sea level	amsl
acre	ac
ampere	A
annum (year)	a
billion	B
billion tonnes	Bt
billion years ago	Ga
British thermal unit	BTU
centimetre	cm
cubic centimetre	cm3
cubic feet per minute	cfm
cubic feet per second	ft3/s
cubic foot	ft3
cubic inch	in3
cubic metre	m3
cubic yard	yd3
Coefficients of Variation	CVs
day	d
days per week	d/wk
days per year (annum)	d/a
dead weight tonnes	DWT
decibel adjusted	dBa
decibel	dB
degree	°
degrees Celsius	°C
diameter	ø
dollar (American)	US$
dollar (Canadian)	Cdn$
dry metric ton	dmt
foot	ft
gallon	gal
gallons per minute (US)	gpm
Gigajoule	GJ
gigapascal	GPa
gigawatt	GW
gram	g
grams per litre	g/L

xvii

grams per tonne	g/t
greater than	>
hectare (10,000 m2)	ha
hertz	Hz
horsepower	hp
hour	h
hours per day	h/d
hours per week	h/wk
hours per year	h/a
inch	in
kilo (thousand)	k
kilobar	kb
kilogram	kg
kilograms per cubic metre	kg/m3
kilograms per hour	kg/h
kilograms per square metre	kg/m2
kilometre	km
kilometres per hour	km/h
kilopascal	kPa
kilotonne	kt
kilovolt	kV
kilovolt-ampere	kVA
kilovolts	kV
kilowatt	kW
kilowatt hour	kWh
kilowatt hours per tonne	kWh/t
kilowatt hours per year	kWh/a
less than	<
litre	L
litres per minute	L/m
megabytes per second	Mb/s
megapascal	MPa
megavolt-ampere	MVA
megawatt	MW
metre	m
metres above mean sea level	mamsl
metres above sea level	masl
metres Baltic sea level	mbsl
metres per minute	m/min
metres per second	m/s
microns	µm
milligram	mg
milligrams per litre	mg/L
millilitre	mL
millimetre	mm

xviii

million	M
million bank cubic metres	Mbm3
million bank cubic metres per annum	Mbm3/a
million tonnes	Mt
minute (plane angle)	'
minute (time)	min
month	mo
ounce	oz
pascal	Pa
centipoise	mPa·s
parts per million	ppm
parts per billion	ppb
percent	%
pound(s)	lb
pounds per square inch	psi
revolutions per minute	rpm
second (plane angle)	"
second (time)	s
short ton (2,000 lb)	st
short tons per day	st/d
short tons per year	st/y
specific gravity	SG
square centimetre	cm2
square foot	ft2
square inch	in2
square kilometre	km2
square metre	m2
three-dimensional	3D
tonne (1,000 kg) (metric ton)	t
tonnes per day	t/d
tonnes per hour	t/h
tonnes per year	t/a
tonnes seconds per hour metre cubed	ts/hm3
volt	V
week	wk
weight/weight	w/w
wet metric ton	wmt
ABBREVIATIONS ANDACRONYMS
abrasion index	Ai
acid potential	AP
acid rock drainage	ARD
acid-base accounting	ABA
Acme Analytical Laboratories Ltd	AcmeLabs

xix

Alaska Department of Environmental Conservation	ADEC
Alaska Department of Fish and Game	ADF&G
Alaska Department of Natural Resources	ADNR
Alaska Department of Transportation	ADOT
Alaska Industrial Development Export Authority	AIDEA
Alaska Mining Licence Tax	AMLT
Alaska National Interest Lands Conservation Act	ANILCA
Alaska Native Claims Settlement Act	ANCSA
Alaska Native Regional Corporations	ANCSA Corporations
Alaska Pollution Discharge Elimination System	APDES
Alaska State Tax	AST
ALS Chemex Laboratories	ALS Chemex
Alternative Minimum Tax	AMT
aluminium oxide	Al2O3
Ambler Mining District Industrial Access Road	AMDIAR
ammonium nitrate/fuel oil	ANFO
Analytical Spectral Devices	ASD
Andover Mining Corp.	Andover
Annual Hardrock Exploration Activity	AHEA
Arctic Property	the Property
atomic absorption spectroscopy	AAS
atomic absorption	AA
average relative difference	AD
barium	Ba
Bear Creek Mining Corporation	BCMC
BGC Engineering Inc.	BGC
Bond ball mill work index	BWi
Canadian Institute of Mining, Metallurgy and Petroleum	CIM
capital cost estimate	CAPEX
carboxymethyl cellulose	CMC
Center of the Universe	COU
Clean Water Act	CWA
Controlled Source Audio Magnetotelluric	CSAMT
copper equivalent	CuEq
copper	Cu
cumulative net cash flow	CNCF
Democratic Republic of the Congo	DRC
digital terrain model	DTM
distributed control system	DCS
Document Management System	DMS
domain electromagnetic	DEM
EBA, A Tetra Tech Company	EBA
effective grinding length	EGL
Electromagnetic	EM
engineering, procurement and construction management	EPCM

xx

environmental impact statement	EIS
Environmental Protection Agency	EPA
Ephemeroptera, Plecoptera, Trichoptera	EPT
Ernst & Young LLP	EY
Exploration Incentive Credits	EICs
Fine Sediment Biotic Index	FSBI
general and administrative	G&A
GeoSpark Consulting Inc.	GeoSpark
global positioning system	GPS
gold	Au
Gross Metal Value	GMV
heating, ventilation and air conditioning	HVAC
inductively coupled plasma	ICP
inductively coupled plasma-atomic emission spectroscopy	ICP-AES
inductively coupled plasma-mass spectrometry	ICP-MS
input/output	I/O
Interior Energy Project	IEP
internal rate of return	IRR
International Electrotechnical Commission	IEC
International Organization for Standardization	ISO
inverse distance squared	ID2
iron	Fe
Kennecott Mining Company	Kennecott
Kennecott Research Center	KRC
kriging efficiency	KE
Lakefield Research Ltd.	Lakefield
lead	Pb
Lerchs-Grossmann	LG
life-of-mine	LOM
light-emitting diode	LED
liquefied natural gas	LNG
Localizer Performance with Vertical Guidance	LPV
magnesium oxide	MgO
magnesium	Mg
massive sulphide	MS
memorandum of understanding	MOU
metal leaching	ML
methyl isobutyl carbinol	MIBC
motor control centres	MCCs
Multi-Sector General Permit	MSGP
NANA Regional Corporation Inc.	NANA
National Environmental Policy Act	NEPA
National Instrument 43-101	NI 43-101
nearest neighbour	NN
net cash flow	NCF

xxi

net present value	NPV
neutralization potential	NP
non-acid generating	NAG
North American Datum	NAD
Northwest Alaska Native Association	NANA
Northwest Arctic Borough	NWAB
NovaCopper Inc.	NovaCopper
NovaGold Resources Inc.	NovaGold
operating cost estimate	OPEX
operator interface stations	OIS
ordinary kriging	OK
potentially acid generating	PAG
Precision Approach Path Indicators	PAPIs
preliminary economic assessment	PEA
programmable logic controllers	PLCs
Qualified Person	QP
quality assurance/quality control	QA/QC
Quantitative Evaluation of Minerals by Scanning Electronic Microscopy	QEMSCAM®
Quantitative Kriging Neighbourhood Analysis	QKNA
Resource Associates of Alaska	RAA
Robertson Geoconsultants Inc	Robertson
rock mass rating	RMR
rock quality designation	RQD
run-of-mine	ROM
Runway End Identification Lighting System	REILS
SAG mill and ball mill	SAB
semi-autogenous grinding	SAG
semi-massive sulphide	SMS
SGS Mineral Services	SGS
short wave infrared	SWIR
silver	Ag
sodium isopropyl xanthate	SIPX
sodium	Na
SRK Consulting	SRK
sulphur	S
tailings storage facility	TSF
Teck Resources Ltd.	Teck
theoretical slope of regression	ZZ*
thermal ionization mass spectrometry	TIMS
Thompson-Howarth Precision Versus Concentration	THPVC
time domain electromagnetic	TDEM
titanium dioxide	TiO2
total inorganic carbon	TIC
Universal Transverse Mercator	UTM
Upper Kobuk Mineral Projects	UKMP

xxii

US Army Corps of Engineers	USACE
US Geological Survey	USGS
vibrating wire piezometers	VWPs
volcanogenic massive sulphide	VMS
water quality standards	WQS
Watts, Griffis and McOuat Ltd.	WGM
Wide Area Augmentation System	WAAS
x-ray fluorescence	XRF
zinc	Zn

xxiii

1.0 SUMMARY

1.1 INTRODUCTION

NovaCopper Inc. (NovaCopper) retained Tetra Tech and EBA, a Tetra Tech Company (EBA) to prepare a preliminary economic assessment (PEA) for the Arctic Project and disclose it in a technical report prepared in accordance with National Instrument 43-101 (NI 43-101) and Form 43-101F1. The Arctic Property (the Property) is located in the Ambler mining district (Ambler District) of the southern Brooks Range, in the Northwest Arctic Borough (NWAB) of Alaska. The Property is located 270 km east of the town of Kotzebue, 36 km northeast of the village of Kobuk, and 260 km west of the Dalton Highway, an all-weather state maintained highway. Figure 1.1 illustrates the location of the Property.

The PEA describes the potential technical and economic viability of establishing a conventional open-pit mine, a mill complex and related infrastructure to process massive and semi-massive copper-zinc-lead-silver-gold mineralization from the Arctic Deposit. A minimum 12-year mine life supporting a nominal 10,000 t/d conventional grinding mill-and-flotation circuit is envisaged. The base case scenario assumes long-term metal prices of $2.90/lb for copper, $0.85/lb for zinc, $0.90/lb for lead, $22.70/oz for silver and $1,300/oz for gold. The PEA was prepared on a 100% ownership basis and all amounts are stated in US dollars unless otherwise noted.

The effective date of this report is September 12, 2013 and the effective date of the resource model is July 30, 2013.

General information for the Arctic Project is summarized in Table 1.1.

NovaCopper Inc.	1-1	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Figure 1.1 Property Location Map

NovaCopper Inc.	1-2	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Table 1.1 General Project Information

Description	Unit	Amount
Mill Feed	Mt	35.7
Life-of-mine (LOM)	years	12
Milling Rate	t/d	10,000
Strip Ratio	waste:mineralized material	8.4
Total Project Capital Cost	$ million	717.7
Average Overall Operating Cost*	$/t milled	63.93
Pre-tax Net Present Value (NPV) at 8% Discount Rate**	$ million	927.7
Pre-tax Internal Rate of Return (IRR)	%	22.8
Pre-tax Payback Period	years	4.6
Post-tax NPV at 8% Discount Rate	$ million	537.2
Post-tax IRR	%	17.9
Post-tax Payback Period	years	5.0

Notes:	*Excludes pre-production cost.
	**The base case utilizes long-term metal prices of $2.90/lb for copper, $0.85/lb for zinc, $0.90/lb for lead, $22.70/oz for silver and $1,300/oz for gold.

The PEA should not be considered to be a prefeasibility or feasibility study, as the economics and technical viability of the Arctic Project have not been demonstrated at this time. The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Furthermore, there is no certainty that the PEA results will be realized.

1.2 PROPERTY DESCRIPTION AND LOCATION

The Arctic Project is located in the Ambler District of the southern Brooks Range, in the NWAB of Alaska. The Property is geographically isolated with no current road access or nearby power infrastructure. The Property is located 270 km east of the town of Kotzebue, 36 km northeast of the village of Kobuk, and 260 km west of the Dalton Highway, an all-weather state maintained highway.

The Property is part of the Upper Kobuk Mineral Projects (UKMP), which includes the Arctic and Bornite deposits. In October 2011, NovaCopper entered into an exploration agreement with NANA Regional Corporation, Inc. (NANA) for the development of the parties’ collective resource interests in the Ambler District. The agreement consolidates certain land holdings of the parties into an area of interest consisting of an approximately 143,000 ha land package.

1.3 GEOLOGY AND MINERALIZATION

The Ambler District is located on the southern margin of the Brooks Range and hosts: 1) a belt of Devonian volcanogenic massive sulphide (VMS) deposits, including the Property which contains the Arctic Deposit; and 2) a belt of Devonian epigenetic carbonate-hosted copper deposits including the Bornite Deposit.

The district encompasses an east-west trending zone of Devonian to Jurassic age submarine volcanic and sedimentary rocks occurring as structurally bound, imbricate allochtons (Hitzman et al. 1986) and further characterized by increasing metamorphic grade to the north. The district shows isoclinal folding in the northern portion and thrust faulting to south (Schmidt 1983).

NovaCopper Inc.	1-3	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

On the Property, VMS-style deposits and prospects (including the Arctic Deposit) are hosted in the Ambler Sequence, a group of Middle Devonian to Early Mississippian, metamorphosed, bimodal volcanic rocks with interbedded tuffaceous, graphitic, and calcareous volcaniclastic metasediments. The Ambler sequence occurs in the upper part of the regional Anirak Schist. VMS-style mineralization is found along the entire 110 km strike length of the district.

The Ambler Sequence has undergone two periods of intense, penetrative deformation. The first deformation period is characterised by upper greenschist-facies metamorphism and formation of a penetrative schistosity. Folding varies from isoclinal folding with local transposition of bedding units, to pervasive upright or slightly overturned folds verging north on all scales. This fold event deforms the transposed bedding and schistosity, and defines the subsequent event.

Stratigraphically, the Ambler Sequence consists of variably metamorphosed calc-turbidites, overlain by calcareous schists with irregularly distributed mafic sills and pillow lavas. These are overlain by the Arctic-sulphide host section which consists mainly of fine-grained, carbonaceous siliciclastic rocks which are in turn overlain by reworked silicic volcanic rocks, including meta-rhyolite porphyries and most notably the regionally extensive Button Schist with its characteristically large relic phenocrysts. Greywacke sandstones, interpreted to be turbidites, occur throughout the section but are concentrated higher in the stratigraphy. Several rock units within the stratigraphy show substantial variation in local thickness as a consequence of basin morphology at the time of deposition.

Alteration at the Arctic Deposit is characterized by magnesium alteration, primarily as talc, chlorite and phengite alteration products associated with the sulphide-bearing horizons and continuing in the footwall. Stratigraphically above the sulphide-bearing horizons, significant muscovite as paragonite is developed and results in a marked shift in Na/Mg (sodium/magnesium) ratios across the sulphide bearing horizons.

Mineralization occurs as stratiform semi-massive sulphide (SMS) to massive sulphide (MS) beds within primarily graphitic chlorite schists and fine-grained quartz sandstones. The sulphide beds average 4 m in thickness but vary from less than 1 m up to as much as 18 m in thickness.

The bulk of the mineralization is within six modelled SMS zones cored by five MS zones lying along the upper and lower limbs of the Arctic isoclinal anticline. Wireframes of the mineralized horizons have been modeled based on MS defined by more than 50% sulphide minerals and SMS defined by 35 to 50% sulphide minerals. All of the zones are within an area of roughly 1 km² with mineralization extending to a depth of approximately 250 m below the surface. Mineralization is predominately coarse-grained sulphides consisting mainly of chalcopyrite, sphalerite, galena, tetrahedrite, arsenopyrite, pyrite and pyrrhotite. Trace amounts of electrum and enargite are also present.

NovaCopper Inc.	1-4	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

1.4 RESOURCE ESTIMATE

A new mineral resource has been estimated by Mr. Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM, who is an independent Qualified Person (QP), as defined in section 1.5 of NI 43-101. The mineral resource is listed in Table 1.2 and the effective date of this resource estimate is July 30, 2013.

The mineral resource estimate prepared by Tetra Tech considers diamond drill holes drilled by different operators from 1965 to 2011. The majority of the drilling has been completed in recent years by NovaCopper and its previous parent company NovaGold Resources Inc. (NovaGold). The mineral resource for the Arctic Project is supported by 43 core holes (approximately 13,500 m) drilled by NovaGold and 92 core holes (approximately 17,600 m) drilled by previous owners Kennecott Mining Company (Kennecott), and/or a Kennecott subsidiary. The geological and assay database have been reviewed and verified by Tetra Tech. It is Tetra Tech’s opinion that the current drilling information is adequate to support a mineral resource estimate.

Leapfrog^™ software (version 2.5.1) was used to review and verify the resource estimation domains, prior to being imported into Isatis^™ software (version 2012.1) to prepare assay data for geostatistical analysis, variography, block model construction, metal grade estimation and mineral resource tabulation. Mineral Resources were estimated into five massive-sulphide and six semi-massive sulphide lenses and then grades for the mineralized material portion of a 10 m by 10 m by 5 m block were calculated based on the proportion of MS and SMS within each block. Whole block model grades were calculated based on the weighted value of the estimated grade item based on the proportion of the mineralization domain. Whole blocks, including internal dilution, were used in pit optimization. Extreme lead and gold assays were capped prior to compositing. Ordinary kriging (OK) and inverse distance squared (ID²) estimates were run, with OK used for resource reporting and ID² used for validation. Search parameters were constrained within each mineralized domain and required an optimum number of 15 composites, minimum number of 5 composites, minimum number of 2 drill holes, and maximum search distance range of 200 m. In general, blocks categorized as Indicated were supported by at least 2 drill holes within a 75 m search radii, and blocks categorized as Inferred were supported by at least 2 drill holes within a 150 m search radii.

Differences between the previously reported mineral resource estimate (as reported in the Technical Report dated April 24, 2012) are primarily related to additional drilling, updated geological interpretation, additional specific gravity determinations, and reporting of grades and tonnes within an open pit designed to support the requirements for reasonable prospects for economic extraction.

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Table 1.2 Resource Estimate for the Arctic Project (NSR Cut-off of $35/t)

Category	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Au (Moz)	Ag (Moz)
Indicated	23.848	3.26	4.45	0.76	0.71	53.2	1,713	2,338	400.9	0.55	40.8
Inferred	3.363	3.22	3.84	0.58	0.59	41.5	239	285	43.2	0.06	4.5

Notes:	1.	These resource estimates have been prepared in accordance with NI 43-101 and the Canadian Institute of Mining, Metallurgy and Petroleum ( CIM) Definition Standards. Mineral resources that are not mineral reserves do not have demonstrated economic viability. Inferred resources have a great amount of uncertainty as to their existence and whether they can be mined legally or economically. It cannot be assumed that all or any part of the Inferred resources will ever be upgraded to a higher category.
	2.	Mineral Resources are reported within mineralization wireframes, contained within an Indicated and Inferred pit design using an assumed copper price of $2.90/lb, zinc price of $0.85/lb, lead price of $0.90/lb, silver price of $22.70/oz, and gold price of $1,300/oz.
	3.	Appropriate mining costs, processing costs, metal recoveries and inter ramp pit slope angles were used to generate the pit design.
	4.	The $35.01/t milled cut-off is calculated based on a process operating cost of $19.03/t, G&A of $7.22/t and site services of $8.76/t. NSR equals payable metal values, based on the metal prices outlined in Note 2 above, less applicable treatment, smelting, refining costs, penalties, concentrate transportation costs, insurance and losses and royalties.
	5.	The LOM strip ratio is 8.39.
	6.	Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content.
	7.	Tonnage and grade measurements are in metric units. Contained copper, zinc and lead pounds are reported as imperial pounds, contained silver and gold ounces as troy ounces.

1.5 MINERAL PROCESSING AND METALLURGICAL TESTING

Since 1970, metallurgical test work has been conducted to determine the flotation response of various samples extracted from the Arctic Deposit. In general, the samples tested produced similar metallurgical performances. In 2012, SGS Mineral Services (SGS) conducted a metallurgical test program to further study metallurgical responses of the samples produced from Zones 1, 2, 3, and 5 of the Arctic Deposit. The flotation test procedures used talc pre-flotation, conventional copper-lead bulk flotation and zinc flotation, followed by copper and lead separation. In general, the 2012 test results indicated that the samples responded well to the flowsheet tested. The average results of the locked cycle tests (without copper and lead separation) were as follows:

• The copper recoveries to the bulk copper-lead concentrates ranged from 89 to 93% excluding the Zone 1 & 2 composite which produced a copper recovery of approximately 84%; the copper grades of the bulk concentrates were 24 to 28%.

• Approximately 92 to 94% of the lead was recovered to the bulk copper-lead concentrates containing 9 to 13% lead.

• The zinc recovery was 84.2% from Composite Zone 1 & 2, 93.0% from Composite Zone 3 and 90.5% from Composite Zone 5. On average, the zinc grades of the concentrates produced were higher than 55%, excluding the concentrate generated from Composite Zone 1 & 2, which contained only 44.5% zinc.

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• Gold and silver were predominantly recovered into the bulk copper-lead concentrates. Gold recoveries to this concentrate ranged from 65 to 80%, and silver recoveries ranged from 80 to 86%.

Using an open circuit procedure, the copper and lead separation tests on the bulk copper-lead concentrate produced from the locked cycle tests generated reasonable copper and lead separation. The copper concentrates produced contained approximately 28 to 31% copper, while the grades of the lead concentrates were in the range of 41% to 67% lead. Also, it appears that most of the gold reported to the copper concentrate and on average the silver was equally recovered into the copper and lead concentrates.

The 2012 grindability test results showed that the Bond ball millwork index (BWi) tests ranged from 6.5 to 11 kWh/t and abrasion index (Ai) tests fluctuated from 0.017 to 0.072 g for the mineralized samples. The data indicates that the samples are neither resistant nor abrasive to ball mill grinding. The materials are considered to be soft or very soft in terms of grinding requirements.

1.6 MINING METHODS

The PEA is based on a conventional truck-and-shovel, open-pit mine design at a single pit. The mining schedule was developed based on a maximum mill capacity of 10,000 t/d. The Arctic Project’s total mine life is 13 years, including 1 year of pre-stripping followed by 12 years of production. The pit uses four pushbacks and a minimum mining width of 40 m. Over the 13-year life, the pit is producing 35.7 Mt of mineralized material and 299.4 Mt of waste rock. The LOM stripping ratio is 8.39 and the stripping ratio excluding the pre-stripping waste rock is 7.94. The pit design incorporates a bench height of 5 m and a 45° inter-ramp angle. After adding the ramps, the overall slope angle will be within 43°, as recommended by EBA (2013). Key mining results are summarized in Table 1.3.

Table 1.3 Summary of Key Mining Results

Item	Units	Value
Mining Pre-stripping	years	1
LOM	years	12
Mineralized Material	Mt	35.7
Waste Material	Mt	299.4
Stripping Ratio	waste:mineralized material	8.4
Average Mining Operating Cost	US$/t milled	28.40

1.7 RECOVERY METHODS

A 10,000 t/d process plant has been designed to process the massive/semi-massive sulphide mineralization that will be supplied from the open pit mine. The main economic elements found in the deposit are copper, lead, zinc, and associated gold and silver. The process plant will operate two shifts per day and 365 d/a with an overall plant availability of 92%. The process plant will produce copper, lead, and zinc concentrates.

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The mill feed will be hauled from the pit to a primary crushing facility where the material will be crushed by a jaw crusher to a particle size of 80% passing 125 mm.

The crushed material will be ground by two stages of grinding, consisting of one semi-autogenous grinding (SAG) mill and one ball mill in closed circuit with hydrocyclones (SAG mill and ball mill (SAB) circuit). The hydrocyclone overflow with a grind size of approximately 80% passing 70 µm will be processed by talc pre-flotation, conventional bulk flotation (to recover copper, lead, and associated gold and silver), and then zinc flotation. The rougher bulk copper-lead concentrate will be reground and cleaned, and then followed by copper and lead separation to produce a lead concentrate and a copper concentrate. The zinc rougher flotation concentrate will be reground and upgraded by cleaner flotation to produce a zinc concentrate. The final tailings from the zinc flotation circuit will be pumped to the tailings storage facility (TSF). Copper, lead, and zinc concentrates will be thickened and pressure-filtered separately before being shipped to smelters.

The LOM average mill feed is expected to contain 2.28% copper, 0.53% lead, 3.13% zinc, 0.5 g/t gold, and 37 g/t silver. According to the mine plan developed for the PEA study and metallurgical test results, the LOM average metal recoveries and concentrate grades are projected below:

• copper concentrate:

 - recovery: 87.1% copper; 57.9% gold; 40.2% silver

 - copper grade: 29%

• lead concentrate:

 - recovery: 74.0% lead; 6.8% gold; 40.2% silver

 - lead grade: 50%

• zinc concentrate:

 - recovery: 86.8% zinc

 - zinc grade: 56%.

1.8 PROJECT INFRASTRUCTURE

The proposed Arctic mine site is spread over a distance of approximately 6 km within the upper reaches of the Sub-Arctic Creek Valley.

Primary access to the Property is currently by air, using both fixed wing aircraft and helicopters. No surface access is currently in place. A memorandum of understanding (MOU) has been signed with the Alaska Industrial Development Export Authority (AIDEA) for the development of the Ambler Mining District Industrial Access Road (AMDIAR) which would provide access from the Dalton Highway (Highway 11) to within 17 km of the Arctic Project.

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The proposed development for the Arctic Project consists of the following major infrastructure:

• roads and an airstrip

• mill buildings and related services facilities including maintenance and truck shops, and assay lab

• water supply and distribution

• waste management

• fuel storage

• on site explosive storage

• power supply

• TSF and water management

• water treatment plant

• construction and permanent camp accommodation

• waste rock storage facilities

• communication.

All buildings and facilities will be constructed with appropriate heating, ventilation and air conditioning (HVAC) and fire protection systems, water distribution and plumbing systems, and dust control systems. Figure 1.2 illustrates the overall Project site layout.

A series of mine haul roads will be constructed from the open pit to the primary crusher, and as well as site roads to and from the truck shop, TSF, and to the Arctic Project airstrip.

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Figure 1.2 Arctic Project Site Layout

1.8.1 POWER SUPPLY

The Arctic Project requires 15 MW of peak load for 10,000 t/d operation demand. Power will be generated by 4+1 self-contained 3.6 MW Prime diesel generators. Four units will be in service with the fifth unit reserved for maintenance. Heat will be recovered from the generators and used to heat the mill.

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1.8.2 TAILINGS STORAGE FACILITY

The co-disposal TSF will be a fully lined facility consisting of rockfill embankment constructed across the Sub-Arctic Creek drainage, creating an impoundment that will extend up the drainage. The rockfill embankment will be constructed to an ultimate crest elevation of 655 mamsl with the embankment being raised in stages to minimize the initial capital construction cost. During operations, potentially acid generating (PAG) waste rock will be placed at the bottom and sides of the basin forming layers with consecutive disposal on tailings that will be filling the voids. The tailings has the potential to generate acid and, therefore, the tailings and the PAG waste rock will be placed under water and remain permanently submerged in order to reduce the potential for acid generation. Additional studies will be required to determine the most suitable method of co-disposal and potential requirements for acid rock drainage (ARD) management and mitigation programs will need to be part of the design of the TSF.

The TSF will be required to contain 110.5 Mm³ total over the 12-year LOM, with 23.8 Mm³to accommodate the tailings at an assumed stored dry density of 1.5 t/m³ and 86.7 Mm³ of PAG waste rock at an assumed stored dry density of 1.9 t/m³. The TSF will be sited as a staged rockfill embankment with an upstream geomembrane liner. The starter embankment will have a crest elevation of 560 m and impound 1 year of mining production, which is approximately 670,000 m³ of tailings and 12.3 Mm³ of waste rock.

1.9 ENVIRONMENTAL

NovaCopper initiated baseline environmental data collection in 2007, including surface water quality sampling, wetlands mapping, stream flow monitoring, aquatic life surveys, subsistence, meteorological monitoring, and acid base accounting sampling. Additional baseline environmental data in the Ambler Lowlands, the Subarctic Creek drainage, the Shungnak River drainage and downstream receiving environments will be required to support future mine design, development of an environmental impact statement (EIS), permitting, construction and operations.

The Arctic Project has the potential to significantly improve work opportunities for local and regional residents. In October 2011, NovaCopper signed an agreement with NANA. In addition to consolidating landholdings in the Ambler District, the agreement has language establishing native hiring preferences and preferential use of NANA subsidiaries for contract work. Furthermore, the agreement formalized an Oversight Committee, with equal representation from NovaCopper and NANA, to regularly review project plans and activities. In addition, a Subsistence Subcommittee has been formed to protect subsistence and the Iñupiaq way of life and a Workforce Development Subcommittee is also in place to address current and future employment needs. NovaCopper meets monthly, during summer months, with the residents of Kobuk, Shungnak and Ambler, the three villages closest to the project area. NovaCopper also meets annually with eight other NANA region villages including Noatak, Kivalina, Kotzebue, Kiana, Deering, Buckland, Selawik and Noorvik, for the purpose of updating residents on project plans and fielding their questions and concerns. NovaCopper has also developed a good working relationship with the NWAB government.

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The Arctic Project will be subject to a mine permitting process which will include compliance with theNational Environmental Policy Act (NEPA) and will require a number of major mine permits from state and federal agencies as well as a significant number of minor permits. Although a number of federal conservation units are located in the general vicinity of the Arctic Project, including but not limited to the Gates of the Arctic National Parks, Kobuk Preserve, Selawik National Wildlife Refuge, and Kobuk Valley and Selawik Wilderness areas, there presence does not change the permitting process nor add to the number of permits required for the Arctic Project.

NovaCopper will be required to develop a formal project description and detailed reclamation and closure plan to support a successful permit application strategy. The mine plan will embrace the concept of “design for closure”. In order to reduce any lasting risk of environmental impacts, the plan will minimize surface disturbances during operations and promote long-term stability of the site after closure.

1.10 CAPITAL COSTS (INITIAL AND SUSTAINING)

The total estimated capital cost for the design, construction, installation and commissioning of the Arctic Project is estimated to be $717.7 million. A summary of the estimated capital cost is shown in Table 1.4. The total sustaining capital costs of $164.4 million for the 12 year LOM including equipment, tailings and other items are summarized in Table 1.5.

This capital cost estimate has been prepared in accordance with the recommended practices of the AACE International. In accordance with the AACE’s International Estimate Classification System, this cost estimate meets or exceeds the specifications for a Class 5 Estimate and has a deemed accuracy of ±35%.

Tetra Tech prepared this estimate with a base date of Q2 2013. No escalation beyond Q2 2013 was applied to the estimate. Quotations provided by vendors are budgetary and non-binding.

The capital cost estimate uses US dollars as the base currency.

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Table 1.4 Capital Cost Summary

	Item	Total Cost ($ million)
	Direct Costs
10	Overall Site	82.5
20	Open Pit Mining	119.7
35	Mineralized Material Handling	17.4
40	Process	122.2
50	Tailings and Water Management	21.0
70	On-Site Infrastructure	49.1
75	Airstrip	14.2
86	External Access Roads	27.2
87	Temporary Services	23.1
Subtotal Direct Costs		476.4
90	Indirect Costs	130.9
98	Owner’s Costs	18.6
99	Contingency	91.9
Total Capital Costs		717.7

Note: Total may not add up due to rounding.

Table 1.5 Summary of Total Sustaining Capital Cost for LOM

Item	Total Cost ($ million)
Mining Equipment	45.6
Tailings	112.8
Other Equipment	6.0
Total Sustaining Capital	164.4

1.11 OPERATING COSTS

The total LOM average operating cost for the proposed mine is estimated at $63.93/t milled. The estimate includes mining, processing, tailings management, general and administrative (G&A), surface services and public road toll costs. The cost is estimated based on a total LOM mill feed of 35.68 Mt from the open pit mine. The nominal annual process rate is approximately 3,650,000 t/a (LOM average annual process rate is approximately 2,973,435 t/a) or 10,000 t/d (LOM daily average rate is 8,146 t/d) at 365 d/a.

Tetra Tech has not estimated the road toll cost that NovaCopper will pay to use the AMDIAR proposed to be built by the Government of Alaska. Since this cost is determined by confidential negotiations between NovaCopper and the AIDEA, a State-owned private corporation, Tetra Tech has relied on NovaCopper management to provide the road toll cost. For the purposes of this PEA study, it has been assumed that a toll would be paid based on a $150 million 30-year bond at a 5% interest rate, which would result in the Arctic Project paying approximately $9.7 million each year for its 12-year mine life. The toll payments are assumed in the PEA to commence when the Arctic Project has reached commercial production.

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The accuracy for the operating cost estimate is expected to be within a range of ±35%. The operating cost estimate uses US dollars as the base currency. The breakdown of the estimated operating costs is presented in Table 1.6.

Table 1.6 Overall Operating Cost Estimate

Area	LOM Average Unit Operating Cost ($/t milled)
Mining*	28.40
Processing**	19.86
G&A	8.92
Plant Services	3.48
Road Toll	3.27
Total	63.93

Notes:	*Excluding preproduction cost
	**Including tailings management operation cost

1.12 ECONOMIC ANALYSIS

This PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Furthermore, there is no certainty that this PEA will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Tetra Tech prepared an economic evaluation of the Arctic Project based on a pre-tax financial model. NPV was estimated at the beginning of the two-year construction period.

As of May 30, 2013, the long-term metal prices applied in the economic analysis are as follows:

• copper: $2.90/lb

• lead: $0.90/lb

• zinc: $0.85/lb

• gold: $1,300.00/oz

• silver: $22.70/oz.

The pre-tax financial results are:

• 22.8% IRR

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• 4.6-year payback on the $717.7 million initial capital costs

• $927.7 million NPV at an 8% discount rate.

NovaCopper engaged the Canadian firm of Ernst & Young LLP (EY) in Vancouver, BC to prepare tax calculations for use in the post-tax economic evaluation of the Arctic Project with the inclusion of US federal and Alaska income taxes and Alaska Mining License tax (Section 22.0) ..

The following post-tax financial results were calculated:

• 17.9% IRR

• 5.0-year payback on the $717.7 million initial capital costs

• $537.2 million NPV at an 8% discount rate.

1.13 INTERPRETATION AND CONCLUSIONS

Based on the work carried out in this PEA and the resultant economic evaluation, this study should be followed by further technical and economic studies leading to a prefeasibility study.

1.14 OPPORTUNITIES AND RECOMMENDATIONS

Tetra Tech recommends the following actions further outlined in Section 26.0 which may support the Arctic Project advancement should NovaCopper proceed with a prefeasibility study. As part of the recommended work program, the following areas of work should be considered:

• resource model integration and additional drilling to upgrade the Inferred Resource

• geotechnical studies, including geotechnical investigations of the pit area, plant site, TSF site, airstrip and other project related locations

• engineering studies, including power supply and optimization of the layout of the process and service related facilities

• metallurgical studies, including process flowsheet and condition optimization and determination of process design related parameters

• additional waste characterization studies

• additional baseline studies and environmental permitting activities

• marketing studies

• trade-off studies to maximize project economics.

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2.0 INTRODUCTION

2.1 TERMS OF REFERENCE

NovaCopper retained Tetra Tech to conduct a PEA on its Arctic Project in the Ambler District of northwestern Alaska (Figure 1.1). The results of that PEA are disclosed in this technical report, which has been prepared in accordance with NI 43-101 and Form 43-101F1.

The PEA describes the potential technical and economic viability of establishing a conventional open-pit copper-zinc-lead-silver-gold mine-and-mill complex for the Arctic Project.

This report supersedes and replaces a previous PEA, prepared for NovaGold by SRK in 2012. The SRK report was based on an underground mine plan with significantly different parameters. On this basis, the previous PEA is considered neither current nor relevant to current development planning on the Property. For completeness, the previous PEA report is available to the interested reader on NovaCopper’s’s SEDAR profile.

Tetra Tech and EBA (a Tetra Tech company) QPs are responsible for all sections of the current technical report. NovaCopper also engaged the Canadian firm EY in Vancouver, BC to prepare tax calculations for use in the post-tax economic evaluation of the Arctic Project (Section 22.0). Tetra Tech used the information completed by these contributors to support information in this current technical report.

The PEA has been prepared on a 100% ownership basis and all amounts are stated in US dollars unless otherwise noted.

2.1.1 UNITS OF MEASUREMENT

All units of measurement in this technical report and resource estimate are metric, unless otherwise stated.

The monetary units are in US dollars, unless otherwise stated.

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2.2 QUALIFIED PERSONS

As shown in Table 2.1, the following people served as the QPs as defined in NI 43-101,Standards of Disclosure for Mineral Projects, and in compliance with Form 43-101F1:

Table 2.1 Qualified Persons

	Report Section		Company	QP
1.0	Summary		All	Sign-off by Section
2.0	Introduction		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
3.0	Reliance on Other Experts		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
4.0	Property Description and Location		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
5.0	Accessibility, Climate, Local Resources, Infrastructure, and Physiography		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
6.0	History		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
7.0	Geological Setting and Mineralization		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
8.0	Deposit Types		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
9.0	Exploration		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
10.0	Drilling		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
11.0	Sample Preparation, Analyses, and Security		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
12.0	Data Verification		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
13.0	Mineral Processing and Metallurgical Testing		Tetra Tech	John Huang, Ph.D., P.Eng.
14.0	Mineral Resource Estimates		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
15.0	Mineral Reserve Estimates		Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
16.0	Mining Methods		Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
17.0	Recovery Methods		Tetra Tech	John Huang, Ph.D., P.Eng.
18.0	Project Infrastructure		-	-
	18.1	Overview	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.2	Roads and Airstrip	-	-
	18.2.1	Ambler Mining District Industrial Access Road	Tetra Tech	Michael Chin, P.Eng.
	18.2.2	Access Road	EBA	Graham Wilkins, P.Eng.
	18.2.3	Haul Roads, Site Roads, and Pad Areas	Tetra Tech	Michael Chin, P.Eng.
	18.2.4	Airstrip	EBA	Graham Wilkins, P.Eng.

table continues…

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		Report Section	Company	QP
	18.3	Buildings and Relevant Building Services	Tetra Tech	John Huang, Ph.D., P.Eng.
	18.4	Water Supply and Distribution	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.5	Waste Management	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.6	Fuel Storage	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.7	On Site Explosives Storage	Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
	18.8	Power Supply to Plant Site	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.9	Tailings Storage Facility	Tetra Tech	Marvin Silva, Ph.D., PE, P.Eng.
	18.10	Waste Rock Facilities	Tetra Tech	Marvin Silva, Ph.D., PE, P.Eng
	18.11	Water Management	Tetra Tech	Marvin Silva, Ph.D., PE, P.Eng.
	18.12	Water Treatment Plant	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.13	Construction and Permanent Camp Accommodation	Tetra Tech	Hassan Ghaffari, P.Eng.
	18.14	Communications	Tetra Tech	Hassan Ghaffari, P.Eng.
19.0	Market Studies and Contracts		Tetra Tech	John Huang, Ph.D., P.Eng.
20.0	Environmental Studies, Permitting, and Social or Community Impact		Tetra Tech	Jack DiMarchi, CPG
21.0	Capital and Operating Costs		-	-
	21.1	Introduction	Tetra Tech	Harvey Wayne Stoyko, P.Eng./ John Huang, Ph.D., P.Eng.
	21.2	Capital Cost Estimate	Tetra Tech	Harvey Wayne Stoyko, P.Eng.
	21.3	Elements of Cost	Tetra Tech	Harvey Wayne Stoyko, P.Eng.
	21.4	Mining Capital Cost Estimate	Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
	21.5	Operating Cost Estimate	-	-
	21.5.1	Summary	Tetra Tech	John Huang, Ph.D., P.Eng.
	21.5.2	Mining Operating Cost Estimate	Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
	21.5.3	Processing Operating Costs	Tetra Tech	John Huang, Ph.D., P.Eng.
	21.5.4	General and Administrative Costs and Surface Services Costs	Tetra Tech	John Huang, Ph.D., P.Eng.
	21.5.5	Tailings Storage Facility Cost	Tetra Tech	Marvin Silva, Ph.D., PE, P.Eng.
	21.5.6	Road Toll Cost	Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
22.0	Economic Analysis		Tetra Tech	Sabry Abdel Hafez, Ph.D., P.Eng.
23.0	Adjacent Properties		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
24.0	Other Relevant Data and Information		Tetra Tech	Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM
25.0	Interpretation and Conclusions		All	Sign-off by Section
26.0	Recommendations		All	Sign-off by Section
27.0	References		All	Sign-off by Section
28.0	Certificates of Qualified Persons		All	Sign-off by Section

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2.3 SITE VISITS

The following QPs conducted site visits to the Arctic Project:

• Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM, completed a site visit on June 19, 2013 for three days.

• Sabry Abdel Hafez, Ph.D., P.Eng., completed a site visit on June 19, 2013 for three days.

• Graham Wilkins, P.Eng., completed a one day helicopter flyover in the fall of 2011.

2.4 SCOPE OF PERSONAL INSPECTIONS

During a June 2013 site visit, Mr. O’Brien undertook a helicopter inspection of the proposed access road, airstrip, Bornite exploration camp, and the Arctic valley. A total of nine drill sites were inspected in the field, during two foot traverses over representative areas of the Arctic Deposit. Collar locations were checked using a hand-held global positioning system (GPS), and within the limits of such instrumentation, collar locations matched those of the 2013 drill database. Several zones of outcrop were inspected, including a zone of “talc” on the backside of the Arctic ridge. There was no active drilling at Arctic; however, NovaCopper logging and sampling facilities were in operation and a review of company procedures found staff to be following industry best practices. Mr. O’Brien also visited the NovaCopper core storage facility in Fairbanks and reviewed three drill holes for lithology, mineralization and the quality of storage. Core boxes were found to be in good condition and intervals were easily retrieved. Furthermore, Mr. O’Brien also held meetings with NovaCopper project staff including Scott Petsel (UKMP Project Manager), Stuart Morris (Senior Project Geologist) and Andy West (Senior Project Geologist).

During a June 2013 site visit, Mr. Abdel Hafez undertook a high-level helicopter traverse along the proposed access road, airstrip and Arctic valley, as well as a visual, on-the-ground inspection of surface conditions near the proposed open pit, waste dump, mill and mine complex, camp, and TSF. During the site visit, Mr. Abdel Hafez viewed the Arctic Project topography and the locations of the proposed infrastructure.

2.5 INFORMATION SOURCES

Reports and documents listed in Section 27.0 were used to support the preparation of the technical report. Additional information was sought from NovaCopper personnel where required.

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3.0 RELIANCEONOTHEREXPERTS

The QPs who prepared this report relied on information provided by other experts as noted below. The relevant QPs believe that it is reasonable to rely on these experts, based on the assumption that the experts have the necessary education, professional designations, and relevant experience on matters relevant to the technical report.

Tetra Tech has relied on NovaCopper’s management team for estimating the annual toll to be paid by the Arctic Project for use of the AMDIAR. This estimate is disclosed in Section 21.0 and is based on discussions between NovaCopper’s management and senior executives of the AIDEA.

Tetra Tech has also relied on NovaCopper’s public disclosures for estimating the applicable royalties on the Arctic Deposit used in the economic analysis. This reliance is based on NovaCopper’s Second Quarter 2013 Management Discussion & Analysis, dated May 31, 2013 (http://www.novacopper.com/i/pdf/financials/2013-Q2-MDA.pdf).

3.1 MICHAEL F. O’BRIEN, M.SC., PR.SCI.NAT, FGSSA, FAUSIMM, FSAIMM

Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM, relied on Erin Workman, P.Geo., Director, Technical Services and Joe R. Piekenbrock, VP Exploration, both from NovaCopper, for matters relating to mineral tenure and mining rights permits, surface rights, royalties, agreements and encumbrances relevant to this report.

3.2 SABRY ABDEL HAFEZ, PH.D., P.ENG.

Sabry Abdel Hafez, Ph.D., P.Eng., relied on EY, concerning tax matters relevant to this report. The reliance is based on a letter to Tetra Tech titled “Assistance with income tax and mineral tax portions of economic analysis prepared by Tetra Tech in connection with the Preliminary Economic Assessment report on NovaCopper’s Arctic Project” and dated July 26, 2013.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 LOCATION

The Property is part of the UKMP mineral tenure package, which includes the Arctic Deposit, as well as numerous additional mineral showings/deposits (Figure 4.1 and Figure 4.2). The Property is located in the Ambler District of the southern Brooks Range, in the NWAB of Alaska,. The Property is located in Ambler River A-2 quadrangle, Kateel River Meridian T 20N R 11E, section 2, and T 21N, R 11E, sections 34 and 35.

The Arctic Project is located 260 km east of the town of Kotzebue, 30 km north of the village of Kobuk, 260 km west of the Dalton Highway, an all-weather state maintained public road, at geographic coordinates N67.17° latitude and W156.38° longitude (Universal Transverse Mercator (UTM) North American Datum (NAD) 83, Zone 4 coordinates 7453080N, 613110E).

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Figure 4.1 Location MapoftheUKMP–Northwest Alaska

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Figure 4.2UpperKobukMineralProjectsLands

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4.2 MINERALTENURE

The Property comprises 45,348 ha of State of Alaska mining claims and US Federal patented mining claims in the Kotzebue Recording District. The Arctic Project land tenure consists of 1,358 contiguous claims, including 875 40-acre State claims, 481 160-acre State claims, and two Federal patented claims comprising 272 acres (110 ha) held in the name of NovaCopper US Inc. These claims are shown in Figure 4.3 and listed in Appendix A. The Arctic Project is located near the southern edge of the centre of the claim block. The Federal patented claim corners were located by the US Geological Survey (USGS). There is no expiration date or labor requirement on the Federal patented claims. Rent for each State claim is paid annually to the Alaska Department of Natural Resources (ADNR). An Annual Labor Statement must be submitted annually to maintain the State claims in good standing.

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Figure 4.3MineralTenurePlan

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4.3 ROYALTIES,AGREEMENTSANDENCUMBRANCES

The Arctic Deposit was discovered in 1965, and originally staked by Bear Creek Mining Corporation (BCMC), a subsidiary of Kennecott Copper Corporation at the time. Until 1993, BCMC was the operator, and from 1993 to 1998 Kennecott Minerals was the operator.

4.3.1 KENNECOTTAGREEMENTS

On March 22, 2004, Alaska Gold Company, a wholly-owned subsidiary of NovaGold Resources Inc. (NovaGold) completed an Exploration and Option to Earn an Interest Agreement with Kennecott Exploration Company and Kennecott Arctic Company (collectively, Kennecott) on the Ambler land holdings.

On December 18, 2009, a Purchase and Termination Agreement was entered into between Alaska Gold Company and Kennecott whereby NovaGold agreed to pay Kennecott a total purchase price of $29 million for a 100% interest in the Ambler land holdings, which included the Arctic Project, to be paid as: $5 million by issuing 931,098 NovaGold shares, and two installments of $12 million each, due 12 months and 24 months from the closing date of January 7, 2010. The NovaGold shares were issued in January 2010, the first $12 million payment was made on January 7, 2011, and the second $12 million payment was made in advance on August 5, 2011; this terminated the March 22, 2004 exploration agreement between NovaGold and Kennecott. Under the Purchase and Termination Agreement, the seller retained a 1% net smelter return (NSR) royalty that is purchasable at any time by the land owner for a one-time payment of $10 million.

During 2011, NovaGold incorporated the NovaCopper entities and transferred its Ambler land holdings, including the Arctic Project from Alaska Gold Company to NovaCopper US Inc. In April 2012, NovaGold completed a spin-out of NovaCopper, with the Ambler lands, to the NovaGold shareholders and made NovaCopper an independent stand alone public company.

4.3.2 NANAAGREEMENT

In 1971, the US Congress passed the Alaska Native Claims Settlement Act (ANCSA) which settled land and financial claims made by the Alaska Natives and provided for the establishment of 13 regional corporations to administer those claims. These are known as the Alaska Native Regional Corporations (ANCSA Corporations). One of these 13 regional corporations is the Northwest Alaska Native Association (NANA) Regional Corporation, Inc. ANCSA Lands controlled by NANA bound the southern border of the Property claim block. National Park lands are within 25 km of the northern property border.

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On October 19, 2011, NovaCopper and NANA Regional Corporation, Inc. entered into an Exploration Agreement and Option to Lease (the “NANA Agreement”) for the cooperative development of their respective resource interests in the Ambler mining district. The NANA Agreement consolidates NovaCopper’s and NANA’s land holdings into an approximately 142,831 ha land package and provides a framework for the exploration and development of the area. The NANA Agreement provides that NANA will grant NovaCopper the nonexclusive right to enter on, and the exclusive right to explore, the Bornite Lands and the ANCSA Lands (each as defined in the NANA Agreement) and in connection therewith, to construct and utilize temporary access roads, camps, airstrips and other incidental works. The NANA Agreement has a term of 20 years, with an option in favour of NovaCopper to extend the term for an additional 10 years. The NANA Agreement may be terminated by mutual agreement of the parties or by NANA if NovaCopper does not meet certain expenditure requirements on NANA’s lands.

If, following receipt of a feasibility study and the release for public comment of a related draft environmental impact statement, NovaCopper decides to proceed with construction of a mine on the lands subject to the NANA Agreement, NovaCopper will notify NANA in writing and NANA will have 120 days to elect to either (a) exercise a non-transferrable back-in-right to acquire between 16% and 25% (as specified by NANA) of that specific project; or (b) not exercise its back-in-right, and instead receive a net proceeds royalty equal to 15% of the net proceeds realized by NovaCopper from such project. The cost to exercise such back-in-right is equal to the percentage interest in the Arctic Project multiplied by the difference between (i) all costs incurred by NovaCopper or its affiliates on the project, including historical costs incurred prior to the date of the NANA Agreement together with interest on the historical costs; and (ii) $40 million (subject to exceptions). This amount will be payable by NANA to NovaCopper in cash at the time the parties enter into a joint venture agreement and in no event will the amount be less than zero.

In the event that NANA elects to exercise its back-in-right, the parties will, as soon as reasonably practicable, form a joint venture with NANA’s interest being between 16% to 25% and NovaCopper owning the balance of the interest in the joint venture. Upon formation of the joint venture, the joint venture will assume all of the obligations of NovaCopper and be entitled to all the benefits of NovaCopper under the NANA Agreement in connection with the mine to be developed and the related lands. A party’s failure to pay its proportionate share of costs in connection with the joint venture will result in dilution of its interest. Each party will have a right of first refusal over any proposed transfer of the other party’s interest in the joint venture other than to an affiliate or for the purposes of granting security. A transfer by either party of a net smelter royalty return on the project or any net proceeds royalty interest in a project other than for financing purposes will also be subject to a first right of refusal.

In connection with possible development on the Bornite Lands or ANCSA Lands, NovaCopper and NANA will execute a mining lease to allow NovaCopper or the joint venture to construct and operate a mine on the Bornite Lands or ANCSA Lands (the “Mining Lease”). These leases will provide NANA a 2% net smelter royalty as to production from the Bornite Lands and a 2.5% net smelter royalty as to production from the ANCSA Lands.

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If NovaCopper decides to proceed with construction of a mine on its own lands subject to the NANA Agreement, NANA will enter into a surface use agreement with NovaCopper which will afford NovaCopper access to the project along routes approved by NANA (the “Surface Use Agreement”). In consideration for the grant of such surface use rights, NovaCopper will grant NANA a 1% net smelter royalty on production and an annual payment of $755 per acre (as adjusted for inflation each year beginning with the second anniversary of the effective date of the NANA Agreement and for each of the first 400 acres (and $100 for each additional acre) of the lands owned by NANA and used for access which are disturbed and not reclaimed.

4.4 ENVIRONMENTALLIABILITIES

There are no known environmental liabilities due to previous operators or ongoing NovaCopper exploration activities at the Property. There has been no mine development or production on the Property. As a result, NovaCopper has not incurred outstanding environmental liabilities in conjunction with its entry into the NANA Agreement.

4.5 PERMITS

Multiple permits are required during the exploration phase of the Property. Permits are issued from Federal, State, and Regional agencies, including: the Environmental Protection Agency (EPA), the US Army Corps of Engineers (USACE), the Alaska Department of Environmental Conservation (ADEC), the Alaska Department of Fish and Game (ADF&G), the Alaska Department of Natural Resources (ADNR), and the NWAB. The State of Alaska permit for exploration on the Property, the Annual Hardrock Exploration Activity (AHEA) Permit, is obtained and renewed every five years through the ADNR – Division of Mining, Land and Water. NovaCopper holds an AHEA exploration permit in good standing with the Alaska DNR, and has done so each year since 2004 under Alaska Gold Company, a wholly owned subsidiary of NovaGold and now NovaCopper. The Property is within the NWAB thus requiring a Title 9 Miscellaneous Land Use permit for mineral exploration, fuel storage, gravel extraction, and the operation of a landfill. NovaGold held these permits in good standing during the 2004 to 2008 seasons and renewed the permits for the 2010 exploration season to 2015. The Bornite Camp, Bornite Landfill, Dahl Creek Camp, and the to-be-constructed Arctic Camp are permitted by the ADEC.

A number of statutory reports and payments are required to maintain the claims in good standing on an annual basis. As the Arctic Project progresses, additional permits for environmental baseline and detailed engineering studies will be necessary at federal, state, and local levels. A detailed outline of permitting requirements is discussed in Section 20.0.

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5.0	ACCESSIBILITY, CLIMATE, LOCAL
	RESOURCES, INFRASTRUCTURE AND
	PHYSIOGRAPHY

5.1	ACCESSIBILITY
5.1.1	AIR

Primary access to the Property is by air, using both fixed wing aircraft and helicopters.

There are four well maintained, approximately 1,500 m-long gravel airstrips located near the Property, capable of accommodating charter fixed wing aircraft. These airstrips are located 66 km west at Ambler, 46 km southwest at Shungnak, 36 km southwest at Kobuk, and 32 km southwest at Dahl Creek. There is daily commercial air service from Kotzebue to the village of Kobuk, the closest community to the Property. During the summer months, the Dahl Creek Camp airstrip is suitable for larger aircraft, such as C-130 and DC-6.

In addition to the four 1,500 m airstrips, there is a 700 m airstrip located at the Bornite Camp, approximately 25 km southwest of the Property, and a 400 m airstrip located approximately 10 km southwest of the Property. The airstrip at Bornite is suited to smaller aircraft, which support the Bornite Camp with personnel and supplies.

5.1.2	WATER
	There is no direct water access to the Property. During spring runoff, river access is possible by barge from Kotzebue Sound to Ambler, Shungnak, and Kobuk via the Kobuk River.
5.1.3	ROAD
	A winter trail and a one -lane dirt track suitable for high-clearance vehicles or construction equipment links the Arctic Project’s main camp at Bornite to the 400 m Dahl Creek airstrip and camp southwest of the Arctic Deposit. An unimproved gravel track connects the airstrip with the Arctic Deposit.
5.2	CLIMATE
	The climate in the region is typical of a sub-arctic environment. Exploration is generally conducted from late May until late September. Weather conditions on the Property can vary significantly from year to year and can change suddenly. During the summer exploration season, average maximum temperatures range from 10°C to 20°C, while average lows range from -2°C to 7°C (Alaska Climate Summaries: Kobuk 1971 to 2000). By early October, unpredictable weather limits safe helicopter travel to the Property. During winter months, the Property can be accessed by snow machine, track vehicle, or fixed wing aircraft. Winter temperatures are routinely below -25°C and can exceed -50°C. Annual precipitation in the region averages at 395 mm with the most rainfall occurring from June through September, and the most snowfall occurring from November through January.

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5.3 LOCAL RESOURCES

The Property is approximately 270 km east of the town of Kotzebue, on the edge of Kotzebue Sound, 36 km northeast of the village of Kobuk, 260 km west of the Dalton Highway, and 470 km northwest of Fairbanks. Kobuk (population 151; 2010 US Census) is a potential workforce source for the Arctic Project, and is the location of one of the airstrips near the Property. Several other villages are also near the Property, including Shungnak located 45 km to the southwest with a population of 262 (2010 US Census) and Ambler, 64 km to the west with a population of 258 (2010 US Census). Kotzebue has a population of 3,201 (2010 US Census) and is the largest population centre in the Northwest Arctic Borough. Kotzebue is a potential source of limited mining-related supplies and labourers, and is the nearest centre serviced by regularly scheduled, large commercial aircraft (via Nome or Anchorage). In addition, there are seven other villages in the region that will be a potential source of some of the workforce for the Property. Fairbanks (population 32,036; 2011 US Census) has a long mining history and can provide most mining-related supplies and support that cannot be sourced closer to the Property.

Drilling and mapping programs are seasonal and have been supported out of the Main Bornite Camp and Dahl Creek Camp. The main Bornite Camp facilities are located on Ruby Creek on the northern edge of the Cosmos Hills. The camp provides office space and accommodations for the geologists, drillers, pilots, and support staff. There are four 2-person cabins installed by NANA prior to NovaCopper’s tenure.

In 2011, the main Bornite Camp was expanded to 20 sleeping tents, 3 administrative tents, 2 shower/bathroom tents, 1 medical tent, and 1 dining/cooking tent. With these additions, the camp capacity was increased to 49 beds. A 30 m by 9 m core logging facility was also built in summer of 2011. An incinerator was installed near the Bornite airstrip to manage waste created by the Arctic Project. Power for the Arctic Project is supplied by a 175 kW Caterpillar diesel generator. Water is provided by a permitted artesian well located 250 m from the Bornite Camp.

In 2012, the camp was further expanded with the addition of a laundry tent, a women's shower/washroom tent, a recreation tent, several additional sleeping tents, and a 2 x enlargement of the kitchen tent. Camp capacity increased to 76 beds. The septic field was upgraded to accommodate the increase in camp population. One of the two-person cabins was winterized for use by the winter caretaker. A permitted landfill was established to allow for the continued cleanup and rehabilitation of the historic shop facilities and surroundings.

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The Dahl Creek camp is an overflow or alternative facility to the main Bornite Camp. The Dahl Creek camp has a main cabin for dining and administrative duties, and a shower facility. Sleeping facilities include two hard-sided sleeping cabins with seven beds (primarily used for staff), two 4-person sleeping tents, and three 2-person sleeping tents for a total of 21 beds. There are support structures, including a shop and storage facilities.

5.4 INFRASTRUCTURE

Proposed infrastructure is discussed in more detail in Section 18.0. Currently, the Arctic Project does not have access to Alaska power and transportation infrastructure.

Beginning in 2009, the Property has been the focus of the Ambler Mining District Access Corridor study. The State of Alaska has spent approximately US$10 M to identify proposed access routes and to initiate environmental baseline studies. The working group for this study consists of the Alaska Department of Transportation (ADOT), the ADNR, the Governor’s Office, the AIDEA, NANA, and NovaCopper.

The proposed Ambler Access Route is a 322 km road running east from the Property to the Dalton Highway. The environmental baseline study for the route is expected to be completed in 2013, with the environmental impact study to follow in early 2014. Following the completion of these studies, public meetings will be held with the final environmental impact study targeted for late 2014.

5.5 PHYSIOGRAPHY

The Arctic Project is located along the south slope of the Brooks Range, which separates the Arctic region from the interior of Alaska. Nearby surface water includes Subarctic Creek, the Shungnak and Kogoluktuk Rivers, the Kobuk River, and numerous small lakes. The Arctic Project is located at the eastern end of Subarctic Creek, a tributary of the Shungnak River to the west, along a ridge between Subarctic Creek and the Kogoluktuk River Valley. The Property area is marked by steep and rugged terrain with high topographic relief. Elevations range from 30 masl along the Kobuk River to 1,180 masl on a peak immediately north of the Arctic Project area. The divide between the Shungnak and Kogoluktuk Rivers in the Ambler Lowlands is approximately 220 masl.

The Kobuk Valley is located at the transition between boreal forest and Arctic tundra. Spruce, birch, and poplar are found in portions of the valley, with a ground cover of lichens (reindeer moss). Willow and alder thickets and isolated cottonwoods follow drainages, and alpine tundra is found at higher elevations. Tussock tundra and low, heath-type vegetation covers most of the valley floor. Intermittant permafrost exists on the Property.

Permafrost is a layer of soil at variable depths beneath the surface where the temperature has been below freezing continuously from a few to several thousands of years (Climate of Alaska 2007). Permafrost exists where summer heating fails to penetrate to the base of the layer of frozen ground and occurs in most of the northern third of Alaska as well as in discontinuous or isolated patches in the central portion of the state.

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Wildlife in the Property area is typical of Arctic and Subarctic fauna (Kobuk Valley National Park 2007). Larger animals include caribou, moose, Dall sheep, bears (grizzly and black), wolves, wolverines, coyotes, and foxes. Fish species include salmon, sheefish, arctic char, and arctic grayling. The Kobuk River, which briefly enters the UKMP on its southwest corner, is a significant salmon spawning river. Subarctic Creek, which does not contain andromous fish, drains into the Shungnak River, which drains into the Kobuk River. The Caribou on the Property belong to the Western Arctic herd that migrates twice a year – south in August, from their summer range north of the Brooks Range, and north in March from their winter range along the Buckland River.

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6.0 HISTORY

Prospectors first arrived in the Ambler District around 1900, shortly after the discovery of the Nome and Fairbanks gold districts. Several small gold placer deposits were located in the southern Cosmos Hills south of the Arctic Deposit and worked intermittently over the next few years. During this time copper mineralization was observed at Ruby Creek in the northern Cosmos Hills; however, no exploration was undertaken until 1947 when local prospector Rhinehart “Rhiny” Berg located outcropping mineralization along Ruby Creek. Berg subsequently staked claims over the Ruby Creek showings and constructed an airstrip for access (alaskamininghalloffame.org 2012).

BCMC, an exploration subsidiary of Kennecott, optioned the property from Berg in 1957. The prospect became known as Bornite and Kennecott conducted extensive exploration over the next decade, culminating in the discovery of the high-grade No. 1 orebody and the sinking of an exploration shaft to conduct underground drilling.

In conjunction with the discovery of the Bornite Deposit, BCMC greatly expanded their regional reconnaissance exploration in the Cosmos Hills and the southern Brooks Range. Stream silt sampling in 1963 revealed a significant copper anomaly in Arctic Creek roughly 17 km northeast of Bornite. The area was subsequently staked and, in 1967, eight core holes were drilled at the Arctic Deposit yielding impressive massive sulphide intercepts over an almost 500-m strike length.

BCMC conducted intensive exploration on the property until 1977 and then intermittently through 1998. No drilling or additional exploration was conducted on the Arctic Project between 1998 and 2004.

In addition to drilling and exploration at the Arctic Deposit, BCMC also conducted exploration at numerous other prospects in the Ambler District (most notably Dead Creek, Sunshine, Cliff, and Horse). The abundance of VMS prospects in the district resulted in a series of competing companies, including Sunshine Mining Company, Anaconda, Noranda, Teck Cominco, Resource Associates of Alaska (RAA), Watts, Griffis and McOuat Ltd. (WGM), and Houston Oil and Minerals Company, entering into a claim staking war in the district in 1973.

District exploration by Sunshine Mining Company and Anaconda resulted in two additional significant discoveries in the district:

• the Sun Deposit located 60 km east of the Arctic Deposit

• the Smucker Deposit located 40 km west of the Arctic Deposit.

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District exploration continued until the early 1980s on the four larger deposits in the district (Arctic, Bornite, Smucker and Sun) when the district fell into a hiatus due to depressed metal prices.

In 1987, Cominco acquired the claims covering the Sun and Smucker deposits from Anaconda. Teck, as Cominco’s successor company, continues to hold the Smucker Deposit. In 2007, Andover Mining Corporation purchased a 100% interest in the Sun Deposit for US$13 million.

In 1981 and 1983, Kennecott received three US Mineral Survey patents (#2245 totalling 240 acres over the Arctic Deposit – later amended to include another 32 acres; and #2233 and #2234 for 25 claims totalling 516.5 acres at Bornite. The Bornite patented claims and surface development were subsequently sold to NANA Regional Corporation, Inc. in 1986.

No production has occurred at the Arctic Deposit or at any of the other deposits within the district.

6.1 PRIOR OWNERSHIP AND OWNERSHIP CHANGES – ARCTIC DEPOSIT AND THEAMBLER LANDS

BCMC initially staked federal mining claims covering the Arctic Deposit area beginning in 1965. The success of the 1960’s drill programs defined a significant high-grade polymetallic resource at the Arctic Deposit and, in the early 1970s, Kennecott began the patent process to obtain complete legal title to the Arctic Deposit. In 1981, Kennecott received US Mineral Survey patent #2245 covering 16 mining claims totalling 240.018 acres. In 1983, US Mineral Survey patent #2245 was amended to include two additional claims totalling 31.91 acres.

With the passage of theAlaska National Interest Lands Conservation Act (ANILCA) in 1980, which expedited native land claims outlined in the ANSCA and state lands claims under theAlaska Statehood Act, both the state of Alaska and NANA selected significant areas of land within the Ambler District. State selections covered much of the Ambler schist belt, host to the VMS deposits including the Arctic Deposit, while NANA selected significant portions of the Ambler Lowlands to the immediate south of the Arctic Deposit as well as much of the Cosmos Hills including the area immediately around Bornite.

In 1995, Kennecott renewed exploration in the Ambler schist belt containing the Arctic Deposit patented claims by staking an additional 48 state claims at Nora and 15 state claims at Sunshine Creek. In the fall of 1997, Kennecott staked 2,035 state claims in the belt consolidating their entire land position and acquiring the majority of the remaining prospective terrain in the VMS belt. Five more claims were subsequently added in 1998. After a short period of exploration which focused on geophysics and geochemistry combined with limited drilling, exploration work on the Arctic Project again entered a hiatus.

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On March 22, 2004, Alaska Gold Company, a wholly-owned subsidiary of NovaGold completed an Exploration and Option Agreement with Kennecott to earn an interest in the Ambler land holdings. A description of the current mineral tenure, as well as recent royalties, agreements and encumbrances is provided in Section 4.0.

6.2	PREVIOUS EXPLORATION AND DEVELOPMENT RESULTS – ARCTIC DEPOSIT
6.2.1	INTRODUCTION

Kennecott’s tenure at the Arctic Project saw two periods of intensive work from 1965 to 1985 and from 1993 to 1998, before optioning the property to NovaGold in 2004.

Though abundant reports, memos, and files exist in Kennecott’s Salt Lake City office, only limited digital compilation of the data exists for the earliest generation of exploration at the Arctic Deposit and within the VMS belt. Beginning in 1993, Kennecott initiated a reevaluation of the Arctic Deposit and assembled a computer database of previous work at the Arctic Deposit and in the district. A new computer-generated block model was constructed in 1995 and an updated resource of the deposit was calculated from the block model. Subsequently, Kennecott staked a total of 2,035 State of Alaska claims in 1997 and, in 1998 undertook the first field program since 1985.

Due to the plethora of companies and the patchwork exploration that occurred as a result of the 1973 staking war, much of the earliest exploration work on what now constitutes the Ambler Schist belt was lost during the post-1980 hiatus in district exploration. The following subsections outline the best documented data at the Arctic Deposit as summarized in the 1998 Kennecott exploration report, including the assembled computer database; however, this outline is not considered to be either exhaustive or in depth.

In 1982, geologists with Kennecott, Anaconda and the State of Alaska published the definitive geologic map of the Ambler schist belt (Hitzman et al. 1982).

Table 6.1 lists known exploration mapping, geochemical, and geophysical programs conducted for VMS targets in the Ambler District.

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Table 6.1 Known Mapping, Geochemical, and Geophysical Programs Targeting VMS Prospects in the Ambler Mining District

Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
Arctic Center of the Universe (COU) Back Door	Arctic	BCMC-KEX	Two (or more) sulfide bands with thickness up to ~40 m with Zn, Cu, Pb, Ag, Au, ±Ba mineralization.	Proffett 1998; Lindberg and others 2004, 2005; NG personnel 2008 at 1:2,000 scale	Extensive 2006 NG program (>670 samples)	Numerous surveys including the 1998 Dighem EM and Mag aerial surveys, 1998 CSAMT survey, TEM downhole and surface surveys in 2005, TDEM ground survey in 2006	Numerous
	COU Back Door, 4th of July Creek	NG-Anaconda	No exposed or drilled mineralization, target is the projection of the Arctic horizon	NG 1:2,000 mapping in 2006	Extensive 2006 NG program	4 TDEM ground surveys in 2005 and 2006	2005 and 2006 NG Progress Reports; Lindberg's 2005 report
Sunshine Bud CS	Sunshine Creek	BCMC and BCMC- Noranda	Disseminated to semi- massive lens up to 18 m thick. Upper mineralized limb is Ba-rich	BCMC 1983; Paul Lindberg 2006; NG 2011	Numerous eras of soil sampling, most recent 1998 by Kennecott (Have data) and 2006 by NG	BCMC completed Recon IP survey and Crone vertical shoot back EM in 1977, 2 TDEM surveys to the NW	Various BCMC reports; Lindberg's 2006 Sunshine progress report; 2006 NG Progress report
	Bud-CS	SMC and TAC	Au-rich gossan and 3+ m intercept of 1.7% Cu, 0.4% Pb, 1.5% Zn, 2 oz/ton Ag, 0.017 oz/ton Au	Anaconda (TAC) and Sunshine (SMC)	SMC soil sampling	Anaconda completed downhole resistivity survey in 1981 on Bud 7	1981 through 1983 Anaconda Progress reports
Dead Creek Shungnak SK	Shungnak (Dead Creek)	BCMC, Cominco	Thin (0.1 to 3 m) disseminated to semi- massive lenses of Cu, Zn, Pb, Ag mineralization	Bruce Otto and others 2006; Proffett 1998	NG in 2006 (355 samples); KEX in 1998 (~240 samples)	At lE 2 CEM surveys by BCMC at DH with no anomalous responses (do not have data)	2006 NG report; 1982 and 1983 Anaconda Ambler Progress reports

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
Dead Creek Shungnak SK(con’t)	SK	GCO and BCMC/GCO- HOMEX JV	Mineralized float up to 0.4% Cu, 4.8% Pb, 8.7% Zn, 5 oz/ton Ag	BCMC	BCMC 1982 soil grid	CEM and Max-min completed by BCMC (do not have data)	1982 Annual Progress Report, BCMC; Bruce Otto 2006 Memo
Horse Cliff DH	Horse-Cliff DH	Horse - BCMC, Cliff SMC, DH - BCMC and BCMC/GCO- HOMEX	Disseminate to semi- massive with local massive lens, thicknesses up to tens of feet.	KEX 1983 1:1000 prospect map	SMC soil surveys 1976-1978 and 1980	No known ground based survey; occurrences within a large resistivity high	1985 Progress Report BCMC-GCO-Homex J; 1980 Summary of Ambler Field Investigations - Sunshine Mining, Horse Creek Memo - Robinson 1981; 1978 Ellis Geologic Evaluation and Assessment of the Nern Belt Claims
Snow Ambler RB Nani Frost	Snow	Cominco	Ag-Pb-Zn mineralization as massive and semi- massive bands hosted within thin bands of graphitic schist (GS).	Noranda- Cominco scanned map with no georeference; Prospect scale	KEX Soil gird in 1997 or 1998	No known ground based survey; Anaconda completed downhole resistivity survey in 1981 on Ambler-4	“Snow Prospect Miscellaneous Notes and Maps.pdf” is only known report
	Ambler	Anaconda TAC	Massive disseminated chalcopyrite and pyrite associated with chert	Numerous Anaconda geologists; no digitized maps	Only scattered soils in database	Max-min surveys, no data is available	1983 Ambler River Memo (Sunshine Progress Report); 1982 Anaconda Progress Report
	Nani-Frost	BCMC and BCMC- Noranda	Outcrops of 2-3 m of 0.8% Cu, 0.4% Pb, 1.2% Zn, 0.05 oz/ton Ag within felsic schist	BGMC (do not have data)	BCMC identified numerous weak soil anomalies (do not have data)	CEM, Max-min, and PEM completed by BCMC (do not have data)	1982 Annual Progress Report, BCMC

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
Red Nora	Nora	BCMC/GCO- HOMEX	Disseminated chalcopyrite within chlorite altered volcanics in two zones (Sulphide Gulch and Nern Horizon)	Generalized geologic map created by WGM for BCMC- GCO- HOMEX	No known data	Two PEM over the Sulphide Gulch horizon	1984 and 1985 Progress Report BCMC- GCO-Homex JV
	Red	BCMC	Thin discordant bands of sphalerite, chalcopyrite, galena, and pyrrhotite with calcite and fluorite cutting 'siltites' and metacarbonates	None	KEX soil lines 1998	KEX identified EM anomalies 1998, follow-up gravity and Max- min EM; TDEM survey in 2006; DIGHEM helicopter EM and radiometric survey in 2006	Kennecott's final 1998 field report; 2006 NG Progress Report
Other	BT, Jerri Creek	Anaconda, AMC	Massive sulphide bands up to 1.5 m thick extend nearly 2.3 km along an E-W strike	Hitzman and others	Historic soils at Jerri Creek	No known surveys.	Hitzman thesis and Anaconda (BT) and Bear Creek (Jerri) Assessment reports; 1982 and 1983 Anaconda Ambler Progress reports
	Kogo- White Creek	Bud - SMC or AMC	Discovered by hydrochemistry of high Cu ions in White Creek.	SMC?	Soil geochem surveys by SMC in 1978 and KEX in 1998	Recon IP survey in 1977; Max-Min Mag survey in 1980; Follow-up Max-Min and gravity by KEX in 1998; TDEM by NG in 2006.	1980 Summary of Ambler Field Investigations, SMC; Kennecott's Final 1998 Field Report
	Pipe	BCMC and SMC	Podiform zones of sulphide mineralization within calc-schists and QMS	Schmidt in 1978, SMC in 1982	Kennecott soil grid in 1997-1998	Not known	Schmidt's 1978 report (Part IV) for Anaconda's (?) annual report

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
Other (con’t)	Tom Tom	Anaconda and SMC	1982 'Discovery' trench by SMC uncovered massive sulphide boulders with up to 6 oz/ton Ag, 5.4% Pb, 6.3% Zn, only 0.2% Cu	Sunshine in 1982 (?)	SMC soils in 1982	Gamma mag survey by SMC in 1982; TDEM by NG in 2006.	1982 Sunshine Mining Company Memo by E.R. Modroo; Schmidt's 1978 report (Part IV) for Anaconda's (?) annual report
Sun	Sun-Picnic Creek	Anaconda - AMC-Cominco; Andover is current owner	Three (?) zones of sulphide mineralization varying from 1 to 10 m; Upper zone is Zn-Pb-Ag rich while the two lower zones are Cu rich	Various Anaconda geologists	Not known, but most likely extensive	Not known, but most likely extensive	1981 Anaconda progress report; Anaconda 1977 prefeasibility study (not in NG possession)
Smucker	Smucker- Charlie- Puzzle-4B- Patti	Anaconda, Cominco, and Bear Creek; now owned by Teck	A single mineralized Ag-Zn-Pb-Cu horizon varying from 1 to 8 m in thickness	Detailed mapping by Anaconda and GCMC geologists	Strong soil geochem anomalies in lowlands SE of Smucker horizon; Kennecott soil grid in 1997 or 1998	Not known	1985 Progress Report BCMC- GCO-Homex JV

Note: EM = electromagnetic; TDEM = time domain electromagnetic; CSAMT = Controlled Source Audio Magnetotelluric

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6.2.2 GEOCHEMISTRY

Historic geochemistry for the district, compiled in the 1998 Kennecott database, includes 2,255 soil samples, 922 stream silt samples, 363 rock samples, and 37 panned concentrate samples. Data has been sourced from several companies including Kennecott, Sunshine Mining, RAA, and NANA. Sourcing of much of the data had been poorly documented in the database.

During 1998, Kennecott renewed its effort in the district, and, as a follow-up to the 1998 EM survey, undertook directed soil and rock chip sampling in and around EM anomalies generated in the geophysical targeting effort. During this period Kennecott collected 962 soils and 107 rocks and for the first time used extensive multi-element inductively coupled plasma (ICP) analysis.

6.2.3 GEOPHYSICS

Prior to 1998, Kennecott conducted a series of geophysical surveys which are poorly documented or are unavailable to NovaCopper. With the renewed interest in the belt, Kennecott mounted a largely geophysically driven program to assess the district for Arctic-sized targets. Based on an initial review of earlier geophysical techniques employed at the Arctic Deposit, Kennecott initiated an extensive helicopter-supported airborne EM and magnetic survey covering the entire VMS belt in March 1998. The survey was conducted on 400 m line spacing with selective 200 m line spacing at the Arctic Deposit and covered 2,509 total line kilometres. The Arctic Deposit presented a strong 900 Hz EM conductive signature.

Forty-six additional discrete EM conductors were identified, of which, 17 were further evaluated in the field. Eight of the EM anomalies were coincident with anomalous geochemistry and prospective geology, and were deemed to have significant potential for mineralization. As a follow-up, each anomaly was located on the ground using a Maxmin 2 horizontal loop EM system. Gravity lines were subsequently completed utilizing a LaCoste and Romberg Model G gravimeter over each of the eight anomalies.

In addition to the EM and gravity surveys in 1998, five lines of CSAMT data were collected in the Arctic Valley. The Arctic Deposit showed an equally strong conductive response in the CSAMT data as was seen in the EM data. As a result of the survey, Kennecott recommended additional CSAMT for the deposit area.

Field targeting work in 1998 prompted Kennecott to drill two exploration holes on anomaly 98-3, located approximately 6 km northwest of the Arctic Deposit and 2 km east-northeast of the Dead Creek prospect. Hole 98-03-01 was drilled to test the sub-cropping gossan and was roughly coincident with the centre of the geophysical anomaly as defined by airborne and ground EM data. Scattered mineralization was encountered throughout the hole with intervals of chalcopyrite and sphalerite.

Based on the results of the 1998 geophysical program, Kennecott made the following recommendations:

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• anomaly 98-3 requires further drilling

• anomalies 98-7 and 98-22 are drill ready

• anomalies 98-8, -9, -14, -35, and - 38 require additional ground targeting.

Kennecott conducted no further field exploration in the district after 1998 and subsequently optioned the property to NovaGold in 2004.

6.2.4 DRILLING

Between 1967 and July 1985, Kennecott (BCMC) completed 86 holes (including 14 large diameter metallurgical test holes) totalling 16,080 m. In 1998, Kennecott drilled an additional 6 core holes totalling 1,492 m to test for:

• extensions of the known Arctic resource

• grade and thickness continuity

• EM anomaly 98-3.

Drilling for all BCMC/Kennecott campaigns in the Arctic Deposit area (1966 to 1998) totals 92 core holes for a combined 17,572 m. A complete and comprehensive discussion of the all the drilling undertaken at the Arctic Deposit is contained in Section 10.0 of this report.

6.2.5 SPECIFIC GRAVITY

Prior to 1998, no specific gravity (SG) measurements were available for the Arctic Deposit rocks. A “factored” average bulk density was used to calculate a tonnage factor for resource estimations. A total of 38 samples from the 1998 drilling at the Arctic Deposit were measured for SG determinations. This included six samples of unaltered metavolcanics, ten samples of graphitic schist and talc schist lithology, seven samples of SMS, and 15 samples of MS.

A complete and comprehensive discussion of SG determinations captured during both the Kennecott and NovaCopper/NovaGold tenures are discussed in Sections 11.0 and 14.0 of this report.

6.2.6 PETROLOGY, MINERALOGY, AND RESEARCH STUDIES

There have been numerous internal studies done by Kennecott on the petrology and mineralogy of the Arctic Deposit that exist as internal memos, file notes, and reports from as early as 1967. Most notable are Clark et al. 1972; Clark et al. 1976; Hunt 1999; Stephens et al. 1970; and Stevens 1982.

In addition, Jeanine Schmidt completed a doctoral dissertation for Stanford University in 1983 entitled “The Geology and Geochemistry of the Arctic Prospect, Ambler District, Alaska”.

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6.2.7 GEOTECHNICAL, HYDROLOGICAL AND ACID-BASE ACCOUNTING STUDIES

A series of geotechnical, hydrological and acid-base accounting (ABA) studies were conducted by Kennecott before their divestiture of the Arctic Project to NovaGold.

GEOTECHNICALSTUDIES

In December 1998, URSA Engineering prepared a geotechnical study for Kennecott titled “Arctic Project – 1998 Rock Mass Characterization”. Though general in scope, the report summarized some of the basic rock characteristics as follows:

• Compressive strengths average 6,500 psi for the quartz mica schists, 14,500 psi for the graphitic schists, and 4,000 psi for talc schists.

• Rock mass quality can be described as average to good quality, massive with continuous jointing except the talc schist, which was characterized as poor quality. The rock mass rating (RMR) averages 40 to 50 for most units except the talc schist which averages 30.

HYDROLOGICALSTUDIES

In 1998, Robertson Geoconsultants Inc. (Robertson) of Vancouver prepared a report for Kennecott titled “Initial Assessment of Geochemical and Hydrological Conditions at Kennecott’s Arctic Project”. The report presented the results of the acid generation potential of mine waste and wall rock for the Arctic Project in the context of a hydrological assessment of the climate, hydrology and water balance analyses at the Arctic Deposit. Climatic studies at the time were limited to regional analyses as no climatic data had been collected at the Arctic Project site prior to the review. Regional data, most specifically a government installed gauging station about 20 miles to the southwest at Dahl Creek, provided information in assessing the hydrology of the Arctic Project at the time. A total of nine regional gauges were utilized to evaluate the overall potential runoff in the area.

ACID-BASEACCOUNTINGSTUDIES

The 1998 Robertson study documented acid-base accounting results based on the selection of 60 representative core samples from the deposit. Results of the study are summarized as follows:

• Roughly 70% of the waste rock material was deemed to be PAG.

• Mitigation of the acid generating capacity could be affected by submersion of the waste rock. Mitigation of the high wall and pit geometries would make potential pit flooding unlikely and could present a long term mitigation issue.

• Characteristics of the mine tailings were not assessed.

• Based on the study, Robertson recommended underground mining scenarios, or aggressive study including site water balance.

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6.2.8 METALLURGICAL STUDIES

During Kennecott’s tenure on the Property, they undertook an extensive series of studies regarding the metallurgy and processing of the Arctic mineralization. An extensive discussion of the historic and current metallurgical studies is presented in Section 13.0 of this report.

6.3 HISTORICAL MINERAL RESOURCE ESTIMATES

A series of historic mineral resources have been estimated for the Arctic Deposit, including Russell (1975), Brown (1985), Randolf (1990), Kennecott (1995).

The historic resource estimates are considered relevant but not necessarily reliable. In each case, a significant amount of resampling and verification is required to upgrade or verify the historical estimate as current mineral resources or mineral reserves. Please note that a QP has not done sufficient work to classify any of the historical estimates as current mineral resources or mineral reserves and the issuer is not treating the historical estimates as current mineral resources or mineral reserves.

6.3.1 RUSSELL – KENNECOTT (1975/1976) RESOURCE ESTIMATE

The oldest documented estimate is a mineral resource estimated by R.H. Russell in 1975 for Kennecott. This employed a polygonal estimation method. The estimate was updated in 1976 after additional drilling, and constrained within an open -pit mining scenario, with minimum 5 ft blocks with greater than 1% copper cut-off. Equivalencies for cut-off calculations include 5.56% zinc equalling 1% copper, and 2.21 oz/ton silver equalling 1% copper. The estimation utilized a tonnage factor of 8.81 ft³per ton. Table 6.2 summarizes the results of the Russell estimation. This historical estimate is not categorized, and it is not known what the equivalent CIM category would be.

Table 6.2Russell 1976 Resource Estimation

Million Tons	Copper (%)	Zinc (%)	Lead (%)	Silver (oz/ton)	Gold (oz/ton)	Basis
36.8	3.97	5.49	0.78	1.60	0.019	Tonnage Factor = 8.81 Cut-off 5 ft of >1% CuEq

Note: CuEq = copper equivalent

6.3.2 BROWN – KENNECOTT (1985) RESOURCE ESTIMATE

In a 1985 economic study, Brown reports resources based on a polygonal method at 4%, 6%, 8%, and 10% polymetallic cut-offs (copper + zinc + 1/2 lead), using 8 ft minimum mining heights, elevations above 2,300 ft, and 10 to 14% dilution factors. In addition, the resource reports additional resources below 2,300 ft elevation at 6% and 8% cut-offs. The resource is based on zones of mineralization ranging from 8 to 40 ft thick covering an area of 2,400 by 2,800 ft. Table 6.3 summarizes results of the Brown estimation.

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This historical estimate is not categorized, and it is not known what the equivalent CIM category would be.

Table 6.3Brown 1985 Resource Estimation

Million Tons	Copper (%)	Zinc (%)	Lead (%)	Silver (oz/ton)	Gold (oz/ton)	Basis
8.8	4.70	8.30	1.15	2.00	0.026	8% cut-off (Cu+Zn+1/2Pb) + 12.7% dilution
3.9	4.04	6.82	1.17	1.90	0.023	Additional tonnage below 2300 ft elevation
5.2	3.00	2.35	0.35	-	-	Additional tonnage 4 to 8% cut-off

6.3.3 RANDOLF – KENNECOTT (1990) RESOURCE ESTIMATE

Another documented Kennecott resource is by Randolf (1990), who completed an updated economic evaluation of the Property, based on exploration drilling completed through 1975. No supporting data or discussion of the basis for the mineral resource estimation is presented. Table 6.4 summarizes results of the Randolf estimate. This historical estimate is not categorized, and it is not known what the equivalent CIM category would be.

Table 6.4 Randolf 1990 Resource Estimation

Million Tons	Copper (%)	Zinc (%)	Lead (%)	Silver (oz/ton)	Gold (oz/ton)	Basis
36.8	3.70	5.05	0.75	1.60	0.02	Unknown

6.3.4 KENNECOTT (1995) RESOURCE ESTIMATE

The most widely repeated historic mineral resource number is the updated 1995 Kennecott resource, produced in conjunction with the construction of the digital database for the Arctic Project. No supporting data or discussion of the basis of the resource estimation is presented. Table 6.5 summarizes the results of the 1995 estimate. This historical estimate is not categorized, and it is not known what the equivalent CIM category would be.

Table 6.5 Kennecott 1995 Resource Estimation

Million Tons	Copper (%)	Zinc (%)	Lead (%)	Silver (g/ton)	Gold (g/ton)	Basis
36.3	4.0	5.5	0.8	54.9	0.7	Unknown

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6.3.5 DEAD CREEK, SUNSHINE AND HORSE CLIFF HISTORICAL RESOURCES

There are a number of other prospects within the Arctic Project boundary, covering the permissive Ambler Sequence. Three of the prospects (Dead Creek, Sunshine and Horse Cliff) have historical resources, as listed in Table 6.6.

Table 6.6 Historical Resources for the Dead Creek, Sunshine and Horse Cliff Prospects

Area	Resource (Mt)	Status	Cu (%)	Zn (%)	Pb (%)	Ag (g/t)
Dead Creek	1.0	Historical	3.0	3.0	2.0	46.6
Sunshine	20.0	Historical	1.4	2.5	0.5	24.8
Horse Cliff	10.0	Historical	1.5	4.5	1.5	31.3

6.4	DEVELOPMENT STUDIES
6.4.1	KENNECOTT TENURE

During Kennecott’s tenure on the Arctic property, no less than 10 different economic studies were completed internally. These studies include: J.L. Halls, 1974, “Ambler District Evaluation”; J.L. Halls, 1976, “Arctic Deposit Order of Magnitude Evaluation”; P.A. Metz, 1978, “Arctic Prospect Summary File Report”; J.L. Halls, 1978, “Arctic Deposit”; C.D. Broadbent, 1981, “Evaluation of the Arctic and Ruby Creek Deposits”; R.R. Dimrock, 1984, “Evaluation Update”; W.J. Brown, 1985, “Pre-AFD Report”; M.P. Randolf, 1990, “Re-Evaluation”; W.L. Jacobsen, 1997, “Arctic Project Mining Potential”; and J. Earnshaw, 1999, “Interim Report Conceptual Level Economic Evaluations of the Arctic Resource”. In addition to the internal studies, SRK completed a “Preliminary Arctic Scoping Study” in 1998 at the request of Kennecott.

The internal studies contemplated both open pit and underground mining scenarios at a variety of the production rates with prevailing metals pricing, capital and operating costs. The studies also targeted a variety of transportation options including roads to both Cape Krusenstern (north of Kotzebue) and Cape Darby on the Seward Peninsula, as well as railroad options to Cape Krusenstern and Whittier. Two studies also evaluated the air transport of concentrate and supplies. The SRK external study only contemplated various underground mining scenarios and did not address high-level economics.

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7.0	GEOLOGICAL SETTING AND
	MINERALIZATION

7.1 REGIONAL GEOLOGY–SOUTHERN BROOKS RANGE

The Ambler mining district (Ambler District) occurs along the southern margin of Brooks Range within an east-west trending zone of Devonian to Jurassic age submarine volcanic and sedimentary rocks (Hitzman et al. 1986). The district covers both: 1) VMS-like deposits and prospects hosted in the Devonian age Ambler Sequence (or Ambler Schist belt), a group of metamorphosed bimodal volcanic rocks with interbedded tuffaceous, graphitic and calcareous volcaniclastic metasediments; and 2) epigenetic carbonate-hosted copper deposits occurring in Devonian age carbonate and phyllitic rocks of the Bornite Carbonate Sequence. The Ambler Sequence occurs in the upper part of the Anirak Schist, the thickest member of the Schist belt or Coldfoot subterrane (Moore et al. 1994). VMS-like stratabound mineralization can be found along the entire 110 km strike length of the district. Immediately south of the Schist belt in the Cosmos Hills, a time equivalent section of the Anirak Schist includes the approximately 1 km thick Bornite Carbonate Sequence. Mineralization of both the VMS-like deposits of the Schist belt and the carbonate-hosted deposits of the Cosmos Hills has been dated at 375 to 387 Ma (Selby et al. 2009; McClelland et al. 2006).

In addition, the Ambler District is characterized by increasing metamorphic grade north perpendicular to the strike of the east-west trending units. The district shows isoclinal folding in the northern portion and thrust faulting to south (Schmidt 1983). The Devonian to Mississippian age Angayucham basalt and the Triassic to Jurassic age mafic volcanic rocks are in low-angle over thrust contact with various units of the Ambler Schist belt and Bornite Carbonate Sequence along the northern edge of the Ambler Lowlands.

7.1.1 TERRANE DESCRIPTIONS

The terminology of terranes in the southern Brooks Range evolved during the 1980s because of the region’s complex juxtaposition of rocks of various composition, age and metamorphic grade. Hitzman et al. (1986) divided the Ambler District into the Ambler and Angayucham terranes. Recent work (Till et al. 1988; Silberling et al. 1992; Moore et al. 1994) includes the rocks of the previously defined Ambler terrane as part of the regionally extensive Schist belt or Coldfoot subterrane along the southern flank of the Arctic Alaska terrane as shown in Figure 7.1 (Moore et al. 1994). In general, the southern Brooks Range is composed of east-west trending structurally bound allochtons of variable metasedimentary and volcanogenic rocks of Paleozoic age.

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Figure7.1 Geologic Terranes of the Southern Brooks Range

The Angayucham terrane, which lies along southern margin of the Brooks Range, is locally preserved as a klippen within the eastern Cosmos Hills and is composed of weakly metamorphosed to unmetamorphosed massive-to-pillowed basalt rocks with minor radiolarian cherts, marble lenses and isolated ultramafic rocks. This package of Devonian to Late Triassic age (Plafker et al. 1977) mafic and ultramafic rocks is interpreted to represent portions of an obducted and structurally dismembered ophiolite that formed in an ocean basin south of the present day Brooks Range (Hitzman et al. 1986; Gottschalk and Oldow 1988). Locally, the Angayucham terrane overlies the schist belt to the north along a poorly exposed south-dipping structure.

Gottschalk and Oldow (1988) describe the Schist belt as a composite of structurally bound packages composed of dominantly greenschist facies rocks, including pelitic to semi-pelitic quartz-mica schist with associated mafic schists, metagabbro and marbles. Locally, the Schist belt includes the middle Devonian age Bornite Carbonate Sequence, the lower Paleozoic age Anirak pelitic, variably siliceous and graphic schists, and the mineralized Devonian age Ambler sequence consisting of volcanogenic and siliciclastic rocks variably associated with marbles, calc-schists, metabasites and mafic schists (Hitzman et al. 1982; Hitzman et al. 1986). The lithologic assemblage of the Schist belt is consistent with an extensional, epicontinental tectonic origin.

Structurally overlaying the Schist belt to the north is the Central belt. The Central belt is in unconformable contact with the Schist belt along a north-dipping low-angle structure (Till et al., 1988). The Central belt consists of lower Paleozoic age metaclastic and carbonate rocks, and Proterozoic age schists (Dillon et al. 1980). Both the Central belt and Schist belt are intruded by meta-to-peraluminous orthogneisses, which locally yield a slightly discordant U-Pb thermal ionization mass spectrometry (TIMS) zircon crystallization age of middle to late Devonian (Dillon et al. 1980; Dillon et al. 1987). This igneous protolith age is supported by Devonian orthogneiss ages obtained along the Dalton Highway, 161 km to the east of the Ambler District (Aleinikoff et al. 1993).

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Overlaying the Schist belt to the south is the Phyllite belt, characterized in the Ambler mining district as phyllitic black carbonaceous schists of the Beaver Creek Phyllite which is assumed to underlie much of the Ambler Lowlands between the Brooks Range and the Arctic Deposit to the north and the Cosmos Hills and the Bornite Deposit to the south. The recessive weathering nature of the Beaver creek phyllite limits the exposure but is assumed to occur as a thrust sheet overlying the main Schist belt rocks.

7.1.2 REGIONAL TECTONIC SETTING

Rocks exposed along the southern Brooks Range consist of structurally bound imbricate allochtons that have experienced an intense and complex history of deformation and metamorphism. Shortening in the fold and thrust belt has been estimated by some workers to exceed 500 km (Oldow et al. 1987) based on balanced cross sections across the central Brooks Range. In general, the metamorphic grade and tectonism in the Brooks Range increases to the south and is greatest in the Schist belt. The tectonic character and metamorphic grade decreases south of the Schist belt in the overlaying Angayucham terrane.

In the late Jurassic to early Cretaceous age, the Schist belt experienced penetrative thrust-related deformation accompanied by recrystallization under high-pressure and low-temperature metamorphic conditions (Till et al. 1988). The northward directed compressional tectonics were likely related to crustal thickening caused by obduction of the Angayucham ophiolitic section over a south-facing passive margin. Thermobarometry of schists from the structurally deepest section of the northern Schist belt yield relict metamorphic temperatures of 475°C, ±35°C, and pressures from 7.6 to 9.8 kb (Gottschalk and Oldow 1988). Metamorphism in the schist belt grades from lowest greenschist facies in the southern Cosmos Hills to upper greenschist facies, locally overprinting blueschist mineral assemblages in the northern belt (Hitzman et al. 1986).

Compressional tectonics, which typically place older rocks on younger, do not adequately explain the relationship of young, low-metamorphic-grade over older and higher-grade metamorphic rocks observed in the southern Brooks Range hinterland. Mull (1982) interpreted the Schist belt as a late antiformal uplift of the basement to the fold and thrust belt. More recent models propose that the uplift of the structurally deep Schist belt occurred along duplexed, north-directed, thin-skinned thrust faults, followed by post-compressional south-dipping low angle normal faults along the south flank of the Schist belt, accommodating for an over-steepened imbricate thrust stack (Gottschalk and Oldow 1988; Moore et al. 1994). Rapid cooling and exhumation of the Schist belt began at the end of the early Cretaceous age at 105 to 103 Ma, based on Ar40/Ar39 cooling ages of hornblende and white mica near Mount Igikpak, and lasted only a few million years (Vogl et al. 2002). Additional post extension compressive events during the Paleocene age further complicate the southern Brooks Range (Mull 1985).

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7.2 AMBLER SEQUENCE GEOLOGY

Rocks that form the Ambler Sequence consist of a lithologically diverse sequence of lower Paleozoic Devonian age carbonate and siliciclastic strata with interlayered mafic lava flows and sills. The clastic strata, derived from terrigenous continental and volcanic sources, were deposited primarily by mass-gravity flow into the sub-wavebase environment of an extending marginal basin.

The Ambler Sequence underwent two periods of intense, penetrative deformation. Sustained upper greenschist-facies metamorphism with coincident formation of a penetrative schistosity and isoclinal transposition of bedding marks the first deformation period. Pervasive similar-style folds on all scales deform the transposed bedding and schistosity, defining the subsequent event. At least two later non-penetrative compressional events deform these earlier fabrics. Observations of the structural and metamorphic history of the Ambler District are consistent with current tectonic evolution models for the Schist belt, based on the work of others elsewhere in the southern Brooks Range (Gottschalk and Oldow 1988; Till et al. 1988; Vogl et al. 2002).

Figure 7.2 shows the location and geology of the Ambler mining district and the Schist belt terrane including the Anirak schist, the Kogoluktuk schist and the Ambler Sequence, the contemporaneous Bornite Carbonate Sequence in the Cosmos Hills to the south, and the allochthonous overthrust Cretaceous sedimentary rocks and Devonian Angayucham volcanic rocks.

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Figure 7.2 Geology of the Ambler Mining District

7.2.1 GENERAL STRATIGRAPHY OF THE AMBLER SEQUENCE

Though the Ambler Sequence is exposed over 110 km of strike length descriptions and comments herein will refer to an area between the Kogoluktuk River on the east and the Shungnak River on the west where NovaCopper has focused the majority of its exploration efforts over the last decade.

The local base of the Ambler Sequence consists of variably metamorphosed carbonates historically referred to as the Gnurgle Gneiss. NovaCopper interprets these strata as calc-turbidites, perhaps deposited in a sub-wavebase environment adjacent to a carbonate bank. Calcareous schists overlie the Gnurgle Gneiss and host sporadically distributed mafic sills and pillowed lavas. These fine-grained clastic strata indicate a progressively quieter depositional environment up section, and the presence of pillowed lavas indicates a rifting, basinal environment.

Overlying these basal carbonates and pillowed basalts is a section of predominantly fine-grained carbonaceous siliciclastic rocks which host a significant portion of the mineralization in the district including the Arctic Deposit. This quiescent section indicates further isolation from a terrigenous source terrain.

The section above the Arctic Deposit host stratigraphy contains voluminous reworked silicic volcanic strata with the Button Schist at its base. The Button Schist is a regionally continuous and distinctive albite porphyroblastic unit that serves as an excellent marker above the main mineralized stratigraphy. The paucity of volcanically derived strata below the Arctic Deposit host section and abundance above indicates that the basin and surrounding hinterlands underwent major tectonic reorganization during deposition of the Arctic Deposit section. Greywacke sands that NovaCopper interpret as channeled high-energy turbidites occur throughout the section but concentrate high in the local stratigraphy. Figure 7.3 shows idealized sections for several different areas in and around the Arctic Deposit.

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Several rock units show substantial change in thickness and distribution in the vicinity of the Arctic Deposit that may have resulted from the basin architecture existing at the time of deposition. Between the Arctic Ridge, geographically above the Arctic Deposit, and the Riley Ridge to the west several significant differences have been documented including:

• The Gnurgle Gneiss is thickest in exposures along the northern extension of Arctic Ridge and appears to thin to the west.

• Mafic lavas and sills thicken from east to west. They show thick occurrences in upper Subarctic Creek and to the west, but are sparsely distributed to the east.

• The quartzite section within and above the Arctic sulphide horizon does not occur in abundance east of Arctic Ridge; it is thicker and occurs voluminously to the west.

• Button Schist thickens dramatically to the west from exposures on Arctic Ridge; exposures to the east are virtually nonexistent.

• Greywacke sands do not exist east of Subarctic Creek but occur in abundance as massive, channeled accumulations to the west, centered on Riley Ridge.

These data are interpreted by NovaCopper to define a generally north-northwest-trending depocentre through the central Ambler District. Diamictite occurrences described below in concert with these formational changes suggest that the depocentre had a fault-controlled eastern margin. The basin deepened to the west; the Riley Ridge section deposited along a high-energy axis, and the COU section lies to the west-southwest distally from a depositional energy point of view. This original basin architecture appears to have controlled mineralization of the sulphide systems at Ambler and Shungnak (Dead Creek), concentrating fluid flow along structures on the eastern basin margin.

Figure 7.4 is a simplified geologic map of the area between the Kogoluktuk and the Shungnak rivers.

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Figure7.3 Ambler Sequence Stratigraphy in the Arctic Deposit Area

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Figure7.4 Generalized Geology of the Central Ambler District

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7.2.2 STRUCTURAL FRAMEWORK OF THE AMBLER DISTRICT

In addition to the underlying pre-deformational structural framework of the district suggested by the stratigraphic thickening of various facies around the Arctic Deposit, the Ambler Sequence is deformed by two penetrative deformational events that significantly complicate the distribution and spatial arrangement of the local stratigraphy.

F1DEFORMATION

The earliest penetrative deformation event is associated with greenschist metamorphism and the development of regional schistosity. True isoclinal folds are developed and fold noses typically are thickened. The most notable F1 fold is the Arctic antiform that defines the upper and lower limbs of the Arctic Deposit. The fold closes along a north-northeast- trending fold axis roughly mimicking the trace of Subarctic Creek and opening to the east. Importantly, the overturned lower limb implies that the permissive stratigraphy should be repeated on a lower synformal isocline beneath the currently explored limbs and would connect with the permissive mineralized stratigraphy to the northwest at Shungnak (Dead Creek). Figure 7.5 shows typical F1 folds developed in calcareous Gnurgle Gneiss.

Figure7.5 Typical F1 Isoclinal Folds Developed in Calcareous Gnurgle Gneiss

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F2 DEFORMATION

The earlier F1 schistosity is in turn deformed by the F2 deformational event that resulted in the local development of an axial planar cleavage. The deformational event is well defined throughout the Schist belt and results in a series of south verging open to moderately overturned folds that define a series of east-west trending folds of similar vergence across the entire Schist belt stratigraphies.

This event is likely temporarily related to the emplacement of the Devonian Angayucham volcanics, the obducted Jurassic ophiolites and Cretaceous sediments over the Schist belt stratigraphies.

In addition to the earlier penetrative deformation events, a series of poorly defined non-penetrative deformation likely as a consequence of Cretaceous extension are seen as a series of warps or arches across the district.

The interplay between the complex local stratigraphy, the isoclinal F1 event, the overturned south verging F2 event and the series of post-penetrative deformational events makes district geological interpretation often extremely difficult at a local scale.

7.3 ARCTIC DEPOSIT GEOLOGY

Previous workers at the Arctic Deposit (Russell 1995 and Schmidt 1983) describe three mineralized horizons at the Arctic Deposit: the Main Sulphide Horizon, the Upper South Horizon and the Warm Springs Horizon. The Main Sulphide Horizon was further subdivided into three zones: the southeast zone, the central zone and the northwest zone. Previous deposit modelling was grade-based resulting in numerous individual mineralized zones representing relatively thin sulphide horizons.

Recent work by NovaCopper define the Arctic Deposit as two or more discrete horizons of sulphide mineralization contained in a complexly deformed isoclinal fold with an upright upper limb and an overturned lower limb hosting the main mineral resources. Nearby drilling suggests a third limb, an upright lower limb, likely occurs beneath the currently explored stratigraphy. Figure 7.6 is a generalized geologic map of the immediate Arctic Deposit area.

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Figure7.6 Generalized Geologic Map of the Arctic Deposit

7.3.1 LITHOLOGIES AND LITHOLOGIC DOMAIN DESCRIPTIONS

Historically, five lithologic groupings have been utilized by Kennecott (URSA Engineering 1998 and Russell 1995) to describe the local stratigraphy of the deposit. These groupings include: 1) metarhyolite (Button Schist) or porphyroblastic quartz feldspar porphyry and rhyolitic volcaniclastic and tuffaceous rocks; 2) quartz mica schists composed of tuffaceous and volcaniclastic sediments; 3) graphitic schists composed of carbonaceous sedimentary rocks; 4) base metal sulphide bearing schists; and 5) talc schists composed of talc altered volcanic and sedimentary rocks.

The principal lithologic units captured in logging and mapping by NovaCopper are summarized and described in the following subsections, in broadly chronologically order from oldest to youngest.

GREENSTONE (GNST)

Greenstones are typically massive dark-green amphibole- and garnet-bearing rocks, differentiated by their low quartz content and dark green color. Textural and colour similarities along with similar garnet components and textures often cause confusion with some sedimentary greywackes within the Ambler Sequence stratigraphy. Intervals of greenstone range up to 80 m in thickness and are identified as pillowed flows, sills and dikes. Multiple ages of deposition are implied as both basal pillowed units are present as well as intrusive sill and dike-like bodies higher in the local stratigraphy.

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CHLORITE SCHIST (CHS)

This unit is likely alteration-related but has been used for rocks where more than half of the sheet silicates are composed of chlorite. In the field, some samples of chlorite schist showed a distinctive dark green to blue-green colour, but in drill core the chlorite schists commonly have lighter green colour. Some intervals of chlorite schist are associated with talc-rich units.

TALC SCHIST (TS)

Talc-bearing schists are often in contact with chlorite-rich units and reflect units which contain trace to as much as 10% talc often occurring on partings. Like the chlorite schist this unit is likely alteration related.

BLACK TO GREY SCHIST (GS)

Black or grey schists appear in many stratigraphic locations particularly higher in the stratigraphy but principally constitute the mineralized permissive stratigraphy of the Arctic Deposit lying immediately below the Button Schist (MRP). The unit is typically composed of muscovite, quartz, feldspar, graphite, and sometimes chlorite, biotite or sulphides. The texture is phyllitic, variably crenulated, well-foliated and suggests a pelitic protolith, likely deposited in a basin progressively filled with terrigenous fine sediment. This unit is host to the MS and SMS horizons that constitute the Arctic Deposit.

BUTTON SCHIST (MRP)

This rock type consists of quartz-muscovite-feldspar schists with abundant distinctive 1 to 3 cm albite porphyroblasts of metamorphic origin and occasional 0.5 to 2 cm blue quartz phenocrysts of likely igneous origin. The unit shows a commonly massive to weakly foliated texture, although locally the rocks have a well-developed foliation with elongate feldspars.

QUARTZ-MICA-(FELDSPAR) SCHIST (QMS/QFMS)

This schistose rock contains variable proportions of quartz, muscovite, and sometimes feldspar. Most contain high amounts of interstitial silica, and some have feldspar or quartz porphyroblasts. The texture of the unit shows significant variability and likely represent both altered and texturally distinct felsic tuffs and volcaniclastic lithologies.

DIAMICTITE (DIA)

This unit contains a range of unsorted, matrix supported polylithic clasts including Button Schist occurring in black to dark grey, very fine-grained graphitic schist. The unit occurs as lenses with other stratigraphies and likely represents local derived debris flows or slumps.

GREYWACKE (GW)

This unit consists of massive green rocks with quartz, chlorite, probably amphibole, feldspar, muscovite, and accessory garnet, biotite, and calcite/carbonate. Voluminous accumulations of medium-grained greywacke occur within, but generally above, the quartz mica schist and are differentiated from texturally similar greenstones by the presence of detrital quartz, fine-grained interbeds, graded bedding and flute casts.

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LITHOGEOCHEMISTRY OF IMMOBILE TRACE ELEMENTS

In 2007, work by NovaGold suggested that many of the nondescript felsic metavolcanic lithologies were simply alteration and textural variants of the felsic rock units and not adequately capturing true compositional lithological differences between units. Twelker (2008) demonstrated that the use of lithogeochemistry utilizing immobile trace elements specifically Al2O3:TiO2 (aluminium oxide:titanium dioxide) ratios could be used to effectively differentiate between different felsic volcanic and sedimentary suites of rocks at the Arctic Deposit.

Lithogeochemistry shows three major felsic rock suites in the Arctic Deposit area: a rhyolite suite; and intermediate volcanic suite and a volcaniclastic suite. These suites are partially in agreement with the logged lithology but in some instances show that alteration in texture and composition masked actual lithologic differences.

Results of the lithogeochemistry have led to a better understanding of the stratigraphic continuity of the various units and have been utilized to more accurately model the lithologic domains of the Arctic Deposit.

LITHOLOGIC DOMAINS

Though a variety of detailed lithology variations are logged during data capture, NovaCopper currently models the following domains within the deposit area itself. These domains which are in part focused on the appropriate domains of acid-generating capacities and their spatial distribution around the fold axes include the following units: both the upper and lower fold limbs of the Button Schist (MRP upper, MRP lower), a series of felsic quartz mica schists of the hangingwall in the upper limb including the Felsite Top and Felsite Upper Limb units, chlorite and talc schists of footwall Felsite Fold unit, and the carbonaceous schists of the Grey Schist unit. SMS and MS are also modelled.

7.3.2 STRUCTURE

Earlier studies (Russell 1977, 1995; Schmidt 1983) concluded mineralization at the Arctic deposit was part of a normal stratigraphic sequence striking northeast and dipping gently southwest. Subsequent reinterpretation by Kennecott in 1998 and 1999 suggested the entire Ambler Sequence at Arctic could be overturned. Proffett (1999) reviewed the Arctic geology and suggested that a folded model with mineralization as part of an isoclinal anticline opening east and closing west could account for the mapped and logged geology. His interpretation called for an F2 fold superimposed on a north-trending F1 fabric.

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Lindberg (2004) supported a folded model similar to Proffett, though he felt the main fold at Arctic is northwest closing and southeast opening. Lindberg named this feature the Arctic Antiform, and interpreted this structure to be an F1 fold.

Lindberg believes the majority of folding within the mineralized horizons occurs in the central part of the deposit within a southwest plunging “cascade zone.” The increased thicknesses of mineralized intervals in this part of the property can in part be explained by the multiple folding of two main mineralized horizons as opposed to numerous individual mineralized beds as shown in the 1995 geologic model. The cascade zone appears to be confined to the upper sulphide limbs of the Arctic Antiform.

Continuity drilling on closer spacing in 2008 across the “cascade” zone confirms the continuity of the two mineralized horizons but does not support the complexity proposed by Lindberg. Dodd et al. (2004) suggested that some of the complexity might be related to minor thrusting. Results of 2006 mapping at Arctic supported the interpretation that an F2 fold event may fold the lower Button Schist back to the north under the deposit in this area (Otto 2006). Deep drilling in 2008 just to the north of the deposit to test the concept drilled the appropriate upright stratigraphy at depth. Though the target horizon was not reached due to the drill rig limitations the hole did encountered significant mineralization below the Button Schist immediately above the sulphide-bearing permissive stratigraphy. That hole (AR07-110) intersected roughly 35 m of anomalous mineralization including 0.45 m of 1.17% copper, 0.8% lead, 5.8% zinc, 49.7 g/t silver and 0.7 g/t gold.

7.3.3 ALTERATION

Schmidt (1988) defined three main zones of hydrothermal alteration occurring at the Arctic Deposit:

• A main chloritic zone occurring within the footwall of the deposit consisting of phengite and magnesium-chlorite.

• A mixed alteration zone occurring below and lateral to sulphide mineralization consisting of phengite and phlogopite along with talc, calcite, dolomite and quartz.

• A pyritic zone overlying the sulphide mineralization.

Field observations conducted by NovaCopper in 2004 and 2005 supported by logging and short wave infrared (SWIR) spectrometry only partially support Schmidt’s observations.

Talc and magnesium chlorite are the dominant alteration products associated with the sulphide-bearing horizons. Talc alteration grades downward and outward to mixed talc-magnesium chlorite with minor phlogopite, into zones of dominantly magnesium chlorite, then into mixed magnesium chlorite-phengite with outer phengite-albite zones of alteration. Thickness of alteration zones vary with stratigraphic interpretation, but tens of metres for the outer zones is likely, as seen in phengite-albite exposures on the east side of Arctic Ridge.

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Stratigraphically above the sulphide-bearing horizons significant muscovite as paragonite is developed and results in a marked shift in sodium/magnesium (Na/Mg) ratios across the sulphide bearing horizons.

Visual and quantitative determination of many of the alteration products is difficult at best due to their light colours and the well-developed micaceous habit of many of the alteration species. Logging in general has poorly captured the alteration products and the SWIR methodology though far more effective in capturing the presence or absence of various alteration minerals adds little in any quantitative assessment.

Of particular note are the barium species including barite, cymrite (a high-pressure Ba phyllosilicate), and Ba-bearing muscovite, phlogopite and biotite. These species associated with both alteration and mineralization has also been strongly remobilized during metamorphism (Schmandt 2009). Though little has been done to document their distribution, they do have a significant impact on bulk density measurements. Additional work will be needed to establish clarity on barium distribution and its effect on bulk density in combination with the various sulphide packages.

Additional discussion of the potential impacts of barite is discussed in the SG section (Sections 11.0 and 14.0) of this report.

Talc is of particular importance at the Arctic Deposit due to its potential negative impact on flotation characteristics during processing. Though talc is recognized as a significant component of the mineralized assemblage, its quantitative distribution is poorly understood at present though logging observations and SWIR measurements suggest that the core of the antiform opening to the east in the footwall of the mineralized horizon has increased quantities.

7.4 ARCTIC DEPOSIT MINERALIZATION

Mineralization occurs as stratiform SMS to MS beds within primarily graphitic chlorite schists and fine-grained quartz sandstones. The sulphide beds average 4 m in thickness but vary from less than 1 m up to as much as 18 m in thickness. The bulk of the mineralization is within five modelled zones lying along the upper and lower limbs of the Arctic isoclinal anticline. All of the zones are within an area of roughly 1 km² with mineralization extending to a depth of approximately 250 m below the surface.

Mineralization characteristically varies from MS to SMS. Wireframes of the mineralized horizons have been modelled based on MS more than 50% sulphide minerals and SMS as 35 to 50% sulphide minerals.

Unlike more typical VMS deposits, mineralization is not characterized by steep metal zonation or massive pyritic zones. Mineralization is dominantly sheet-like zones of base metal sulphides with variable pyrite and only minor zonation usually on an extremely small scale. No stockworks or stringer zones in association with the mineralization have been observed. More importantly, the mineralization in general exhibits characteristics and textures common to replacement-style mineralization.

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Mineralization is predominately coarse-grained sulphides consisting mainly of chalcopyrite, sphalerite, galena, tetrahedrite, arsenopyrite, pyrite and pyrrhotite. Trace amounts of electrum and enargite are also present. Gangue minerals associated with the mineralized horizons include quartz, barite, white mica, black chlorite, talc, calcite, dolomite and cymrite. Figure 7.7 shows a typical massive sulphide interval. The 2 m interval grades 8.2% copper, 11.6% zinc, 1.6% lead, 103.2 g/t silver and 0.82 g/t gold.

Figure7.7 Typical Massive Sulphide Mineralization at the Arctic Deposit

7.5 GENESIS

Historic interpretation of the genesis of the Ambler Schist belt deposits have called for a syngenetic VMS origin with steep thermal gradients in and around seafloor hydrothermal vents resulting in metal deposition due to the rapid cooling of chloride bound base metals. A variety of VMS types have been well documented in the literature (Franklin et al. 2005) with the Ambler Schist belt deposits most similar to deposits associated with a bimodal mafic dominant volcanism related to incipient rifting.

The majority of field observations broadly support such a scenario at the Arctic Deposit and include: 1) the tectonic setting with Devonian volcanism in an evolving continental rift; 2) the geologic setting with bimodal volcanics including pillow basalts and limited felsic volcanic tuffs; 3) an alteration assemblage with well-defined magnesium-rich footwall alteration and sodium-rich hanging wall alteration; and 4) typical polymetallic base-metal mineralization with massive and semi-massive sulphides.

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Although the majority of field observations support a VMS genesis to the deposits of the Schist belt, a series of other observations and characteristics suggest a more direct genetic link with that of the carbonate-hosted Bornite Deposit in the Devonian Bornite Carbonate Sequence. Both deposit types have been dated at 375 to 387 Ma suggesting a clear temporal link.

Mineral deposition at the Bornite Deposit is attributed to the reduction of oxidized low temperature brines with carbonaceous and/or early framboidal pyrite. Like the Bornite Deposit, mineralization at the Arctic Deposit occurs at the first significant reductant boundary, the graphitic black schists surrounding the main ore-bearing stratigraphy immediately overlying oxidized mafic volcanics. Red beds also occur at greater depth in the stratigraphy within the Kogoluktuk Schist immediately below the Anirak Schist. Both oxidized mafic volcanics and red beds are considered prime sourcing for copper-rich chloride-rich fluids. Additionally, the replacement-like textures and the broad sheet-like zones of thin mineralization over kilometres in strike length suggest a depositional gradient more in line with a reductant control than the high temperature steep gradients related to more typical temperature dependent VMS and black smoker environments.

7.6 DEPOSITS AND PROSPECTS

In addition to the Arctic Deposit, numerous other VMS-like occurrences are present on the NovaCopper land package. The most notably of these occurrences are the Dead Creek (also known as Shungnak), Sunshine, Cliff, Horse, Cobre and the Snow prospects to the west of the Arctic Deposit and the Red, Nora, Tom-Tom and BT prospects to the east. Figure 7.9 shows the NovaCopper land package and the mineral prospects. Figure 7.9 also shows: 1) the Smucker deposit on the far west end of the Ambler Sequence which is currently controlled by Teck Inc.; 2) the Sun deposit at the eastern end of the Ambler Sequence and controlled by Andover Mining Corp., and 3) carbonate-hosted deposits and prospects in the Bornite Carbonate Sequence controlled by NovaCopper/NANA.

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Figure 7.8 Prospects of the Ambler Mining District

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8.0 DEPOSITTYPES

The mineralization at the Arctic Deposit and at several other known occurrences within the Ambler Sequence stratigraphy of the Ambler District, consists of Devonian age, polymetallic (zinc-copper-lead-silver-gold) VMS-like occurrences. VMS deposits are formed by and associated with sub-marine volcanic-related hydrothermal events. These events are related to spreading centres such as fore arc, back arc or mid-ocean ridges. VMS deposits are often stratiform accumulations of sulphide minerals that precipitate from hydrothermal fluids on or below the seafloor. These deposits are found in association with volcanic, volcaniclastic and/or siliciclastic rocks. They are classified by their depositional environment and associated proportions of mafic and/or felsic igneous rocks to sedimentary rocks. There are five general classifications (Franklin et al. 2005) based on rock type and depositional environment:

• Mafic rock dominated often with ophiolite sequences, often called Cyprus type.

• Bimodal-mafic type with up to 25% felsic volcanic rocks.

• Mafic-siliciclastic type with approximately equal parts mafic and siliciclastic rocks, which can have minor felsic rocks and are often called Besshi type.

• Felsic-siliciclastic type with abundant felsic rocks, less than 10% mafic rocks and shale rich.

• Bimodal-felsic type where felsic rocks are more abundant than mafic rocks with minor sedimentary rocks, also referred to as Kuroko type.

Prior to any subsequent deformation and/or metamorphism, these deposits are often bowl- or mound-shaped with stockworks and stringers of sulfide minerals found near vent zones. These types of deposit exhibit an idealized zoning pattern as follows:

• Pyrite and chalcopyrite near vents.

• A halo around the vents consisting of chalcopyrite, sphalerite and pyrite.

• A more distal zone of sphalerite and galena and metals such as manganese.

• Increasing manganese with oxides such as hematite and chert more distal to the vent.

Alteration halos associated with VMS deposits often contain sericite, ankerite, chlorite, hematite and magnetite close to the VMS with weak sericite, carbonate, zeolite, prehnite and chert more distal. These alteration assemblages and relationships are dependent on degree of post deposition deformation and metamorphism. A modern analog of this type of deposit is found around fumaroles or black smokers in association with rift zones.

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In the Ambler District, VMS-like mineralization occurs in the Ambler Sequence schists over a strike length of approximately 110 km. These deposits are hosted in volcaniclastic, siliciclastic and calcareous metasedimentary rocks interlayered with mafic and felsic metavolcanic rocks. Sulphide mineralization occurs above the mafic metavolcanic rocks but below the Button schist, a distinctive district wide felsic unit characterized by large albite porphyroblasts after relic phenocrysts. The presence of the mafic and felsic metavolcanic units is used as evidence to suggest formation in a rift-related environment, possibly proximal to a continental margin.

A sulphide-smoker occurrence has been tentatively identified near Dead Creek northwest of the Arctic Deposit and suggests local hydrothermal venting during deposition. However, the lack of stockworks and stringer-type mineralization at the Arctic Deposit suggest that the deposit may not be a proximal vent type VMS. Although the deposit is stratiform in nature, it exhibits characteristics and textures common to replacement-style mineralization. At least some of the mineralization may have formed as a diagenetic replacement.

At Arctic, sulphides occur as semi-massive (30 to 50% sulphide) to massive (greater than 50% sulphide) layers, typically dominated by pyrite with substantial disseminated sphalerite and chalcopyrite and trace amounts of galena and tetrahedrite. The Arctic Deposit sulphide accumulation is thought to be stratigraphically correlative to those seen at the Dead Creek and Sunshine deposits up to 12 km to the west.

There is also an occurrence of epithermal discordant vein and fracture hosted base metal (lead-zinc-copper) mineralization with significant fluorite mineralization identified at the Red prospect in the Kogoluktuk Valley, east of the Arctic Deposit. Although not yet fully understood, the genesis of this occurrence is considered to be related to the regional system that formed the VMS deposits in the Ambler District.

Although the majority of field observations support a VMS genesis to the deposits of the Ambler Sequence schists, a series of other observations and characteristics suggest a possible direct genetic link with that of the Devonian carbonate-hosted Bornite Deposit to the south of the Ambler Sequence schists. Importantly, both deposit types have been dated at 375 to 387 Ma suggesting a clear temporal link.

Mineral deposition at the Bornite Deposit is attributed to the reduction of oxidized low temperature brines by carbonaceous and/or early framboidal pyrite. Like the Bornite Deposit, mineralization at the Arctic Deposit occurs at the first significant redox boundary in the stratigraphic section, which at the Arctic Deposit, consists along the graphitic black and grey schists surrounding the main mineralization bearing stratigraphy. This reduced stratigraphy overlies oxidized greenstones and at even greater depth oxidized red beds of the Kogoluktuk schist.

The replacement-like textures and the broad sheet-like zones of thin mineralization over kilometres in strike length suggest a depositional gradient more in line with a redox boundary control rather than the high temperature steep gradients typically associated with VMS and black smoker environments. Additionally, the district-wide regional metal zonation shows high copper at the Bornite Deposit, to higher zinc at the Arctic, Dead Creek and Sunshine deposits, to lower copper and higher zinc with higher gold-silver mineralization at both Smucker and Sun on the distal ends of the district. This is similar to the broad district zonation seen in basinal brine-derived systems such as the Kupferschiefer in Poland.

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Analogs to the basinal brine model would be more in line with the Kupferschiefer or the Democratic Republic of the Congo (DRC) deposits such as Tenke and Kamoa. Though the Kupferschiefer and the Congolese deposits lack any significant volcanic component, geochemically, they appear much more similar to the Arctic Deposit than typical VMS deposits. In the DRC, deposits also show a very similar district zonation with epigenetic dolomite systems such as Kipushi (an excellent analog to Bornite) temporally similar to the adjacent basinal systems.

There are important implications for exploration philosophy and methodology between a VMS origin and a basinal brine origin for the Ambler District deposits which could impact the overall potential of the district. Basinal systems represent some of the highest grade and most laterally extensive copper deposits in the world.

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9.0 EXPLORATION

The following section summarizes and highlights work completed by NovaCopper and its predecessor company NovaGold. NovaGold began exploration of the Arctic Deposit and surrounding lands of the Schist belt in 2004 after optioning the Property from Kennecott. Previous exploration on the Property during Kennecott’s tenure is summarized in Section 6.0.

Table 9.1 summarizes the exploration work conducted by NovaGold and NovaCopper during their tenure from 2004 to the present. Field exploration was largely conducted during the period between 2004 to 2007 with associated engineering and characterization studies between 2008 and the present. Drilling related to exploration is discussed in Section 10.0.

Table 9.1 Summary of NovaCopper/NovaGold Exploration Activities Targeting VMS-style Mineralization in the Ambler Sequence Stratigraphy and the ArcticDeposit

Work Completed	Year	Details	Focus
Geological Mapping
-	2004	-	Arctic Deposit surface geology
-	2005	-	Ambler Sequence west of the Arctic Deposit
-	2006	-	COU, Dead Creek, Sunshine, Red
Geophysical Surveys
SWIR Spectrometry	2004	2004 drill holes	Alteration characterization
TDEM	2005	2 loops	Follow-up of Kennecott DIGHEM EM survey
	2006	13 loops	District targets
	2007	6 loops	Arctic extensions
Downhole EM	2007	4 drill holes	Arctic Deposit
Geochemistry
-	2005	-	Stream silts – core area prospects
-	2006	-	Soils – core area prospects
-		-	Stream silts – core area prospects
-	2007	-	Soils – Arctic Deposit area
Survey
Collar	2004 to 2011	GPS	All 2004 to 2011 NovaCopper drill holes
	2004, 2008	Resurveys	Historical Kennecott drill holes
Photography/Topography	2010	-	Photography/topography

table continues…

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Work Completed	Year	Details	Focus
Technical Studies
Geotechnical	2010	BGC	Preliminary geotechnical and hazards
ML/ARD	2011	SRK	Preliminary ML and ARD
Metallurgy	2012	SGS	Preliminary mineralogy and metallurgy
Geotechnical and Hydrology	2012	BGC	Preliminary rock mechanics and hydrology
Project Evaluation
Resource Estimation	2008	SRK	Resource estimation
PEA	2011	SRK	PEA
	2012	SRK	PEA update

Note: SWIR = short wave infrared; ML = metal leaching; BGC = BGC Engineering Inc.; SRK = SRK Consulting

9.1 GRIDS AND SURVEYS

Survey and data capture during Kennecott’s tenure on the Property was in the UTM coordinates system NAD27 Zone 24. In 2010, NovaGold converted all historical geology and topographic data for the Arctic Deposit into UTM coordinates NAD83 for consistency. At that time NovaGold contracted WHPacific to re-establish project-wide survey control and benchmarks for the Arctic Deposit. Current mineral resource and geologic models, including the updated Tetra Tech mineral resource model discussed herein, use topography completed in 2010 by PhotoSat Inc. The resolution of the satellite imagery utilized was 0.5 m and a 1 m contour domain electromagnetic (DEM) was generated.

9.2 GEOLOGICAL MAPPING

NovaGold has focused its exploration mapping efforts on an area covering approximately 18 km of strike length of the permissive Ambler Sequence rocks of the Schist belt stratigraphy. This area is centered on the Arctic Deposit and covers the thickest portion of the Ambler Sequence rocks. The area covers many of the most notable mineralized occurrences including the Red Prospect east of the Kogoluktuk River, the Arctic Deposit, and the nearby occurrences at the West Dead Creek and Dead Creek prospects, and the CS, Bud and Sunshine prospects west of the Shungnak River.

In 2004, mapping focused on the surface geology in and around the Arctic Deposit while exploration in 2005 extended the Ambler Sequence stratigraphy to the west. In 2006 with expansion of the exploration focus to encompass the immediate district and to support a major TDEM geophysical program, mapping was extended to include the area between the Sunshine prospect on the west and the Red prospect on the east. Figure 9.1 shows areas mapped by successive campaigns.

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Figure 9.1 Mapping Campaigns in and around the Arctic Deposit

The following geologists made significant contributions during the following mapping campaigns:

• Paul Lindberg (2004, 2005, and 2006)

• Doyle Albers (2004 and 2005)

• Bruce Otto (2006)

• Josh Ellis (2006)

• Nathan Chutas (2006).

Figure 9.2 shows a compilation of the mapping and the geology of the Arctic Deposit area highlighting stratigraphy within the Ambler Sequence.

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Figure 9.2 Arctic Deposit Area Geology

9.3 GEOCHEMISTRY

During NovaGold and NovaCopper’s collective tenure in the Ambler District, significant soil and silt geochemical sampling was utilized to target many of the VMS prospects in the Ambler Sequence particularly in the core area around the Arctic Deposit. Between 2005 and 2007, NovaGold collected 2,272 soils and 278 silt samples. Much of the reconnaissance soil sampling has used gridding layouts of 200 m lines and 50 m sample intervals oriented perpendicular to stratigraphy.

Soil and silt samples were submitted directly to either ALS Chemex in Fairbanks or Alaska Assay Labs in Fairbanks for sample preparation. The samples were dried and sieved to 80 mesh and forwarded to ALS Chemex for analysis. The samples were analyzed using the ME-ICP61 method and a four acid near total digestion with 27 elements measured (silver, aluminum, arsenic, barium, beryllium, bismuth, calcium, cadmium, cobalt, chromium, iron, potassium, magnesium, manganese, molybdenum, sodium, nickel, phosphorus, lead, sulphur, antimony, strontium, titanium, vanadium, tungsten, and zinc).

Figure 9.3 and Figure 9.4 illustrate typical soil sampling campaigns and sample density and shows copper and zinc distribution, respectively in silt and soil samples in the Dead Creek prospect area. Section 10.0 discusses the geochemistry and sampling of the drill core.

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Figure 9.3 Copper Distribution in Silt and Soil Samples in the Dead Creek Area

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Figure 9.4 Zinc Distribution in Silt and Soil Samples in the Arctic Deposit Area

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9.4 GEOPHYSICS

A number of different geophysical survey methods have been utilized at the Arctic Deposit during Kennecott’s tenure on the Property and are summarized in Section 6.0. During NovaCopper’s tenure, the geophysical methodology was largely focused on ground and downhole EM methods to follow-up on the 1998 DIGHEM airborne EM survey conducted by Kennecott.

From 2005 to 2007, NovaCopper conducted ongoing TDEM surveys and completed 21 different loops targeting the Arctic Deposit, extensions to the Arctic Deposit and a series of DIGHEM airborne anomalies in and around known prospects and permissive stratigraphy. Table 9.2 summarizes the TDEM loops and locations. Figure 9.5 illustrates typical TDEM loops and contoured resistivity at the Dead Creek prospect.

Frontier Geosciences of Vancouver, BC completed all of the geophysical programs using a Geonics PROTEM 37 transmitter, a TEM-57 receiver and either a single channel surface coil or a three component BH43-3D downhole probe.

Table 9.2 TDEM Loops and Locations

Area	2005	2006	2007
Arctic	1	-	6
COU	1	3	-
Dead Creek	-	4	-
Sunshine	-	2	-
Red	-	1	-
Tom Tom	-	1	-
Kogo/Pipe	-	2	-
Total	2	13	6

In addition to the TDEM surveys, Frontier Geosciences surveyed four drill holes (AR05-89, AR07-110, AR07-111, and AR07-112). All of the holes produced off-hole anomalies, notably AR07-111, which showed evidence of a strong EM conductor north of the hole. Follow-up is warranted.

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Figure 9.5 TDEM Loops and Contoured Resistivity – Dead Creek Prospect

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9.5 BULK DENSITY

Bulk density determinations are discussed in Section 11.0.

9.6 PETROLOGY, MINERALOGY AND RESEARCH STUDIES

During 2004, NovaGold completed and extensive study of the 2004 drilling utilizing an Analytical Spectral Device (ASD) shortwave infrared spectrometer to better identify alteration species within the Arctic Deposit. The results are discussed in Section 7.0.

NovaCopper supported a series of academic studies of the Arctic Deposit. In 2009, Danielle Schmandt completed an undergrad thesis entitled “Mineralogy and Origin of Zn-rich Horizons within the Arctic Volcanogenic Massive Sulfide Deposit, Ambler District, Alaska” for Smith College. The Schmandt thesis focused on a structural and depositional reconstruction of the Arctic Deposit with the goal of locating the hydrothermal vents to aid in exploration vectoring.

Bonnie Broman, a NovaCopper geologist is currently completing a Master of Science thesis for the University of Alaska focusing on the nature and distribution of the silver-bearing mineral species within the Arctic Deposit.

9.7 GEOTECHNICAL, HYDROGEOLOGICAL AND ACID BASE ACCOUNTING STUDIES

NovaCopper undertook a series of geotechnical, hydrological and ABA studies which are summarized in Table 9.1. For a review of historical geotechnical, hydrological and ABA studies undertaken by Kennecott, refer to Section 6.0.

9.7.1 GEOTECHNICAL STUDIES

In November 2010, BGC completed a preliminary geotechnical study for NovaGold. The report focused on geotechnical aspects and hazards (avalanche mitigation) associated with the construction and maintenance of road infrastructure between the Bornite and the Arctic Deposits and accessing the Arctic Deposit by developing adit access.

9.7.2 HYDROLOGICAL STUDIES

In October 2011, BGC completed a rock mechanics and hydrogeological site investigation at the Arctic Deposit. The study was based on 1,200 m of geotechnical drilling in 5 core holes. Rock core from the drill holes was oriented using the ACTII system, and core samples were submitted for geomechanical laboratory testing. Drill holes were instrumented with grouted-in vibrating wire piezometers and dataloggers, and one drill hole was instrumented with a 10-bead thermistor string.

Data obtained from the drilling program was used to develop a geotechnical model, and the rock mass of the Arctic Project has been divided into six geotechnical units in the bedrock: WRZ (weathered rock), AZ (alteration zone), MZ (mineralized zone), FZ (fault zone), PHFW (proximal hangingwall/footwall), and DHFW (distal hangingwall/footwall). The RMR of the six units is: WRZ-42; AZ–47; MZ–62; FZ–32; PHFW-56 and DHFW-59, respectively.

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Structural data was reviewed and summarized in terms of joints, faults/shears, foliation, and major structure. Structural data collected during the 2008 and 2011 geotechnical programs indicate similar trends for the pervasive southwesterly dipping foliation.

A limited amount of temperature data was collected to investigate the potential for permafrost in the Arctic Project area. Although the data set is small, only one of the four drill holes with temperature data from 2011 indicates the potential for permafrost. This suggests that the Arctic Deposit could be permafrost-free, with the possible exception of the ridge top. Permafrost likely exists elsewhere in the Arctic Project area, particularly over the valley floor in areas covered with thick organics.

Results from the drilling program were used to develop a conceptual hydrogeological model. Measured hydraulic conductivity ranges from 3 x 10^-9 to 6 x 10^-7 m/s and is interpreted to be primarily related to fracture porosity. Several tests conducted in low rock quality designation (RQD) intervals and/or in intervals where faults failed due to water bypassing the top packer or insufficient pump capacity, indicating that higher inflow rates from some fault zones should be anticipated.

Piezometric levels recorded by vibrating wire piezometers installed at the site ranged from 181 m below ground near a topographic high to 13 m below ground at the lowest elevation of the drilling program. Nested vibrating wire installations indicate downward hydraulic gradients of 0.14 and 0.67, respectively. The available hydraulic head data indicated that groundwater flow is topographically driven by recharge in the uplands and discharge in the valley bottom.

9.7.3 ACID-BASE ACCOUNTING STUDIES

In July 2011, SRK completed a preliminary ML and ARD study of the Arctic Deposit for NovaCopper.

The ARD potential varies with rock type and sulphur content. Based on the current sampling, the rock types with the highest potential for ARD are mineralized material, gray schist and felsic schist with more than 0.5% total sulphur. Felsic schist with less than 0.5% total sulphur and talc/chlorite schist with less than 1% total sulphur are predominantly potentially non-acid generating (NAG) with a few samples having uncertain ARD potential. Figure 9.6 summarizes the PAG versus NAG results by rock type for the study.

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Figure 9.6 PAG versus NAG by Rock Type at the Arctic Deposit

The study recommended humidity cells and initial leach tests to evaluate sulphide reactivity and ML. Mineralogical determinations were also recommended due to the complex sulphur and carbonate mineralogy.

9.7.4 METALLURGICAL STUDIES

In October 2012, SGS completed an in-depth metallurgical study of the Arctic Deposit in support of this report. An extensive discussion of the SGS metallurgical study and historic Kennecott metallurgical studies is presented in Section 13.0.

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10.0 DRILLING

Drilling at the Arctic Deposit has been ongoing since its initial discovery in 1965. Approximately 31,907 m of drilling in 135 drill holes have been completed at the deposit or on potential extensions in 23 campaigns spanning 45 years. All of the drill campaigns have been run under the auspices of either: 1) Kennecott and its subsidiaries, or 2) NovaGold, NovaCopper’s predecessor company.

Table 10.1 summarizes operators, campaigns, holes and metres drilled on the deposit. All drillholes listed in this table have been used in the estimation of the new resource disclosed in Section 14.0.

Table 10.1 Companies, Campaigns, Drill Holes and Metres Drilled at the Arctic Deposit

Year	Company	No. of Holes	Metres
1966	Bear Creek	1	32
1967	Bear Creek	7	774
1968	Bear Creek	17	3,782
1969	Bear Creek	3	712
1970	Bear Creek	3	831
1971	Bear Creek	2	663
1973	Bear Creek	2	557
1974	Bear Creek	3	900
1975	Bear Creek	26	4,942
1976	Bear Creek	8	479
1977	Bear Creek	3	497
1979	Bear Creek	2	371
1981	Bear Creek	1	458
1982	Bear Creek	4	494
1983	Bear Creek	1	153
1984	Bear Creek	2	253
1986	Bear Creek	1	184
1998	Kennecott	6	1,523
2004	NovaGold	11	2,996
2005	NovaGold	9	3,393
2007	NovaGold	4	2,606
2008	NovaGold	14	3,306
2011	NovaGold	5	1,193
Total	-	135	31,097

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Additional historical exploration drilling by operators other than Bear Creek/Kennecott exists in the VMS belt, however a portion is unavailable or has been lost over the years. Figure 10.1 shows drill locations of the all resource and exploration holes utilized in the Mineral Resource estimation of this report.

Figure 10.2 shows all holes utilized for metallurgical, geotechnical, hydrological and acid base accounting studies. Drill holes used for NovaCopper’s metallurgical study (SGS 2012) include AR08- 115, 117w, and 119. Drill holes utilized used geotechnical and hydrological studies (BGC 2011) include AR11-127, 128, 129, 130 and 131. Drill holes used in the ABA study (SRK 2011) include AR04-78, 83, 87, AR05-95, 97, AR08-116,120,123 and 126.

NovaCopper and its predecessor company NovaGold drilled 17,983 m in 59 different drill holes targeting the Arctic Deposit and several other prospects of the Ambler Schist belt. Table 10.2 summarizes all of the NovaCopper/NovaGold tenure drilling on the Property.

Table 10.2 Summary of NovaCopper/NovaGold Drilling

Year	Metres	No. of Drill Holes	Sequence	Purpose of Drilling
2004	2,996	11	AR04-78 to 88	Deposit scoping and verification
2005	3,030	9	AR05-89 to 97	Extensions to the Arctic Deposit
2006***	3,100	12	AR06-98 to 109	Property-wide exploration drilling
2007	2,606	4	AR07-110 to 113	Deep extensions of the Arctic Deposit
2008*	3,306	14	AR08-114 to 126	Grade continuity and metallurgy
2011**	1,193	5	AR11-127 to 131	Geotechnical studies
2012***	1,752	4	SC12-014 to 017	Exploration drilling – Sunshine

Notes:	*A total of 12 of the 14 holes drilled in 2008 were utilized in the 2012 SRK resource update. Two holes were maintained in sealed frozen storage to provide additional metallurgical samples if required.
	**Geotechnical holes drilled in 2011 are not included in the current resource estimation contained herein.
	***Drilling in 2006 and 2012 targeted exploration targets elsewhere in the VMS belt.

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Figure 10.1 Plan Map of Drill Holes Utilized in the Mineral Resource Estimation

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Figure 10.2 Drill Holes Utilized for Metallurgical, Geotechnical, Hydrological and Acid-Base Accounting Studies

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10.1 DRILL COMPANIES

Over the Arctic Project’s history, a relatively limited number of drill companies have been used by both Kennecott and NovaCopper/NovaGold at the Arctic Deposit. During Kennecott’s tenure on the Property, Sprague and Henwood, a Pennsylvania-based drilling company was the principal contractor. Tonto Drilling provided services to Kennecott during Kennecott’s short return to the district in the late 1990s. NovaCopper and NovaGold have utilized Boart Longyear as their only contractor. Table 10.3 summarizes drill companies and core sizes utilized.

Table 10.3 Drill Contractors, Drill Holes, Metreage and Core Sizes by Drill Campaign at the Arctic Deposit

Year	Company	No. of Drill Holes	Metres	Core Size	Drill Contractor
1966	Bear Creek	1	32	BX	Sprague and Henwood
1967	Bear Creek	7	774	BX	Sprague and Henwood
1968	Bear Creek	17	3,782	BX	Sprague and Henwood
1969	Bear Creek	3	712	BX	Sprague and Henwood
1970	Bear Creek	3	831	BX	Sprague and Henwood
1971	Bear Creek	2	663	BX?	Sprague and Henwood
1973	Bear Creek	2	557	BX?	Sprague and Henwood
1974	Bear Creek	3	900	NX and BX	Sprague and Henwood
1975	Bear Creek	26	4,942	NX and BX	Sprague and Henwood
1976	Bear Creek	8	479	NXWL and BXWL	Sprague and Henwood
1977	Bear Creek	3	497	NXWL and BXWL?	Sprague and Henwood
1979	Bear Creek	2	371	NXWL and BXWL?	Sprague and Henwood
1981	Bear Creek	1	458	NXWL and BXWL?	Sprague and Henwood
1982	Bear Creek	4	494	NXWL and BXWL?	Sprague and Henwood
1983	Bear Creek	1	153	NXWL and BXWL?	Sprague and Henwood
1984	Bear Creek	2	253	NXWL and BXWL?	Sprague and Henwood
1986	Bear Creek	1	184	NXWL and BXWL?	Sprague and Henwood
1998	Kennecott	6	1,523	HQ	Tonto
2004	NovaGold	11	2,996	NQ and HQ	Boart Longyear
2005	NovaGold	9	3,393	NQ and HQ	Boart Longyear
2007	NovaGold	4	2,606	NQ and HQ	Boart Longyear
2008	NovaGold	14	3,306	NQ and HQ	Boart Longyear
2011	NovaGold	5	1,193	NQ and HQ	Boart Longyear

Sprague and Henwood utilized company manufactured drill rigs during their tenure on the Property. Many of their rigs remain at the Bornite Deposit and constitute a historical inventory of 1950s and 1960s exploration artifacts. The 2004 to 2012 NovaCopper/NovaGold drill programs used a single skid-mounted LF-70 core rig, drilling HQ or NQ core. The drill was transported by skid to the various drill pads using a D-8 bulldozer located on site. The D-8 was also used in road and site preparation. Fuel, supplies and personnel were transported by helicopter.

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10.2	DRILL CORE PROCEDURES
10.2.1	KENNECOTT TENURE

There is only partial knowledge of specific drill core handling procedures used by Kennecott during their tenure at the Arctic Deposit. All of the drill data collected during the Kennecott drilling programs (1965 to 1998) was logged on paper drill logs, copies of which are stored in the Kennecott office in Salt Lake City, Utah. Electronic scanned copies of the paper logs, in PDF format, are held by NovaCopper. Drill core was cut with half core submitted to various assay labs and the remainder stored in Kennecott core storage facility at the Bornite Deposit. In 1995, Kennecott entered the drill assay data, the geologic core logs, and the downhole collar survey data into an electronic format. In 2009, NovaGold geologists verified the geologic data from the original paper logs against the Kennecott electronic format and then merged the data into a Microsoft SQL database.

Sampling of drill core prior to 2004 by Kennecott and BCMC focused primarily on the mineralized zones. During the 1998 campaign, Kennecott did sample some broad zones of alteration and weak mineralization, but much of the unaltered and unmineralized rock remains unsampled. ALS Chemex was used for analyses conducted by Kennecott. Earlier BCMC sampling was even more restricted to mineralized zones of core. Intervals of visible sulphide mineralization were selected for sampling and analyses were conducted primarily by Union Assay Office Inc. of Salt Lake City, Utah. At least six other labs were used during that time period, but mostly as check labs or for special analytical work. Numerous intervals of weak to moderate mineralization remain unsampled in the historic drill core.

10.2.2 NOVAGOLD/NOVACOPPER TENURE

Throughout NovaCopper’s tenure on the Property, the following standardized core handling procedures have been implemented. Core is slung by helicopter to either the Dahl Creek (2004 to 2008) or Bornite (2011 to 2012) camps, where core-logging facilities have been established. Upon receiving a basket of core, geologists and geotechs first mark the location of each drilling block on the core box, and then convert footages on the blocks into metres. All further data capture is then based on metric measurements. Geotechs or geologists measure the intervals (or “from/to”) for each box of core using the drilling blocks and written measurements on the boxes.

Geotechs fill out metal tags with the hole ID, box number and “from/to”, and staple them to each core box. Geotechs then measure the core to calculate percent recovery and RQD. RQD is the sum of the total length of all pieces of core over 12 cm in a run. The total length of core in each run is measured and compared to the corresponding run length to determine percent recovery.

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Geologists then mark sample intervals to capture each lithology or other geologically appropriate intervals. Sample intervals of core are typically between 1 and 3 m in length but are not to exceed 3 m in length. Occasionally if warranted by the need for better resolution of geology or mineralization, smaller sample intervals have been employed. Geologists staple sample tags on the core boxes at the start of each sample interval, and mark the core itself with a wax pencil to designate sample intervals. Sample intervals used are well within the width of the average mineralized zones in the resource area. This sampling approach is considered sound and appropriate for this style of mineralization and alteration.

Core is then logged with lithology and visual alteration features captured on observed interval breaks. Mineralization data, including total sulfide (recorded as percent), sulfide type (recorded as a relative amount), gangue and vein mineralogy are collected for each sample interval with an average interval of approximately 2 m. Structural data is collected as point data. Geotechnical data (core recovery, RQD) were collected along drill run intervals.

After logging, the core is digitally photographed and cut in half using diamond core saws. Specific attention to core orientation is maintained during core sawing to ensure the best representative sampling. One-half of the core is returned to the core box for storage on site and the other half was bagged and labeled for sample processing and analysis. Select specific gravity measurements are also taken and are further discussed in Sections 11.0 and 14.0 of this report. The remaining half core is stored on site or at NovaCopper's Fairbanks warehouse.

10.3 GEOTECHNICAL DRILL HOLE PROCEDURES

Five HQ3 diameter diamond drill holes were completed during NovaGold’s 2011 geotechnical site investigation program at the Arctic Deposit. As shown in Figure 10.2, the drill holes were drilled using an LF 70 Boart-Longyear drill and were supervised by BGC on a 24-hour basis. Oriented core measurements were obtained using the ACT II tool. Constant rate injection and falling head packer tests were completed and vibrating wire piezometers (VWPs) equipped with single channel dataloggers (RST Instruments Ltd. DT2011 model) were installed. The ACT II core orientation system was used to orient discontinuities. Geotechnical logging was completed at the drill site by BGC. Point load testing was completed by NovaCopper once the core had been flown by helicopter back to the Bornite exploration camp. Core sampling for laboratory testing was completed by both BGC and NovaCopper.

All holes received either a single or a nest of two VWPs with single channel dataloggers. The VWPs were lowered to a pre-selected depth attached to a string of polyvinyl chloride pipes, which was then used as a tremie tube to backfill the hole with cement-bentonite grout. Data from each VWP was recorded by a single channel datalogger with a storage capacity and battery life exceeding one year. Knowledge of the barometric pressure was required for accurate conversion of the vibrating wire piezometer data. A Solinst barologger was installed at AR11-0128 for this purpose. The barologger was recorded continuously and downloaded at the same time as the VWP dataloggers. A thermistor was installed at AR11-0129 to monitor ground temperatures. A datalogger was not attached to this instrument, and therefore manual reading was required. Table 10.4 lists the geotechnical holes and outlines geotechnical instrumentation installed.

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Table 10.4 Geotechnical Holes and Instrumentation

Hole ID	BGC ID	Total Depth (m)	VWP 1 Depth (m)	VWP 2 Depth (m)	Thermistor Depth (m)	Barologger on Surface
AR11-127	BGC-E	248.41	186.1	-	-	No
AR11-128	BGC-A	188.98	138.6	-	-	Yes
AR11-129	BGC-D	269.75	244.9	149.9	15.35	No
AR11-130	BGC-F	225.55	63	-	-	No
AR11-131	BGC-B	260.3	123.3	230.3	-	No

Point load testing was also completed on select intervals of core approximately every 5 m. Both axial and diametral tests were completed to investigate variation with respect to loading direction and foliation orientation.

10.4 METALLURGICAL DRILL HOLE PROCEDURES

A metallurgical test program was completed on four composite samples (Zone 1 & 2, Zone 3, Zone 5 and Zone 3 & 5) assembled from core in drill holes (AR08-115, 117w, and 119) representing material from the Arctic Deposit. The scope of the work included mineralogy, comminution and flotation test work. Figure 10.2 shows the plan location of metallurgical drill holes.

SGS Mineral Service’s Quantitative Evaluation of Minerals by Scanning Electronic Microscopy (QEMSCAM^®) methodology utilizing high definition scanning electron microscopy was used on all composites to better determine mineralogy and mineral distributions.

Limited grindability tests were conducted on five selected samples. Standard Bond grindability test (BWi) for ball mill grinding and abrasion index test were conducted.

A conventional flowsheet which produced a bulk copper/lead and zinc concentrates was developed using the four composite samples through flotation testing. The flowsheet development primarily evaluated various primary grind size, depressant dosages and the effect of reagent dosage for copper/lead and zinc circuits. Regrinding of both the copper/lead and zinc concentrates was included in the flowsheet.

A detailed discussion and review of the metallurgical results can be found in Section 13.0.

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10.5	COLLAR SURVEYS
10.5.1	KENNECOTT TENURE

Kennecott provided NovaGold with collar coordinates for all historical holes in UTM coordinates using the NAD27 datum. NovaGold re-surveyed selected historical holes in 2004 and again in 2008. The re-surveys showed little variation compared to the historical surveys.

10.5.2 NOVAGOLD/NOVACOPPER TENURE

Collar location coordinates have been determined in all NovaGold/NovaCopper drill campaigns with a two Ashtech ProMark 2 GPS units using the Riley Vertical Angle Bench Mark (611120.442E, 7453467.486N) as the base station for all surveys. Data collection times varied from 30 minutes to 2 hours. Afternoon hours provided poor satellite constellations, so all surveying was completed during the morning hours. Raw GPS data was processed with Ashtech Solutions 2.60. All surveyed data was collected in the NAD27 datum.

A 2010 survey by a WHPacific Registered Land Surveyor observed differences between the 2010 and historical coordinates used for the Riley VABM, which were of the same magnitude (0.5 m east, 0.1 m north and 1.0 m down) as other Arctic drill collars that were re-surveyed for the third time. A correction was applied to all Arctic drill holes based upon the newly established coordinates for the Riley VABM, along with converting from NAD27 to NAD83 datums. All post 2010 surveys are completed in NAD83.

During a site visit by Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM on June 20 and 21, 2013, nine collars were located using a Garmin^™ Etrex 20 GPS unit. The difference between reported and measured positions ranged between 3.4 and 7.8 m with an average discrepancy of 4.8 m. These differences are within the tolerances expected for GPS verification. The collar location surveys at the Arctic Deposit are considered to be sufficiently accurate for this study.

10.6 DOWNHOLE SURVEYS

BCMC did not perform downhole surveys prior to 1971 (AR-32). In 1971, BCMC began to survey selected (mineralized) holes using a Sperry-Sun downhole survey camera usually at 30.5 m (100 ft) intervals. They were able to re-enter and survey a few of the older holes. BCMC and later Kennecott, applied a single azimuth (49°) and uniform dip deviation every 15.24 m (50 ft) that flattens with depth to all holes collared vertically that were not surveyed.

During NovaCopper’s tenure on the Property, downhole surveys from 2004 to 2011 were collected utilizing either a Reflex EZ-shot camera or a Ranger single-shot tool with individual survey readings collected at the drill rig on roughly 50 to 60 m intervals. The downhole survey data shows a pronounced deviation of the drill holes toward an orientation more normal to the foliation.

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Tetra Tech considers the downhole survey information at the Arctic Deposit to be of sufficient quality for this study.

10.7 RECOVERY

Core recovery during NovaGold/NovaCopper’s tenure has been good to excellent, resulting in quality samples with little to no bias. There are no other known drilling and/or recovery factors that could materially impact accuracy of the samples during this period. Table 10.5 shows recoveries and RQD for each of the NovaGold/NovaCopper campaigns exclusive of the geotechnical drill holes in 2011. BGC (2012) reports a detailed and exhaustive discussion of the recoveries and RQDs of the 2011 drilling.

Table 10.5 Recovery and RQD 2004 to 2008 Arctic Drill Campaigns

Year	Metres	Recovery (%)	RQD (%)
2004	2,996	98.0	73.4
2005	3,030	96.0	74.4
2007	2,606	95.7	73.1
2008	3,306	98.0	80.1

Incomplete Kennecott data exists with regards to overall core recovery but based on 917 intervals of 10 m or less in the historical database, the average recovery was 92%. Kennecott RQD measurements in the 1998 program averaged 87.0. There has been no systematic evaluation of recovery by rock type.

Tetra Tech considers the core recovery to be adequate to support the sampling and logging for this study.

10.8 DRILL INTERCEPTS

The updated resource herein contains additional drilling not included in the previous 2012 SRK resource estimation (SRK 2012). Table 10.6 presents significant drill intercepts that have been returned since the previous Arctic Deposit mineral resource estimate (SRK 2012). Notably, several 2008 drill holes are now included in the resource estimation. These holes were maintained in frozen storage to provide additional metallurgical samples if needed. With the completion of the ongoing metallurgical studies by SGS in October 2012, these additional drill holes were assayed and added to the estimation database.

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Table 10.6 Drill Intercept Summary Table

Drill Hole ID	Easting (m)	Northing (m)	From (m)	To (m)	Length (m)	Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)
AR11-0127	613435.4	7453049.1	87.50	90.19	2.69	5.43	0.51	3.70	1.14	56.5
AR11-0128	613047.5	7453300.4	87.78	93.76	5.98	2.75	1.00	5.19	0.85	49.3
AR11-0129	613145.8	7453094.4	150.85	163.07	12.22	3.85	0.99	8.10	0.69	72.7
			185.63	195.97	10.34	1.79	0.84	3.92	0.29	37.7
			199.37	207.78	8.41	3.44	1.08	7.22	0.78	77.7
AR11-0130	612955.3	7452904.1	167.00	179.05	12.05	2.88	0.97	4.58	0.63	48.3
AR11-0131	613191.8	7453315.3	166.80	168.50	1.70	4.60	5.92	23.14	1.42	205.9

10.9 DRILLING AT OTHER PROSPECTS

In addition to the drilling focused at the Arctic Deposit, significant exploration drilling has been carried out elsewhere on the property targeting numerous occurrences along the Ambler Schist belt. Much of this exploration is historical in nature and is summarized herein.

Drill results from many of the major prospects are tabulated in Table 10.7 and Table 10.8 and show the limited amount of drilling within the main prospect areas. Figure 10.3 shows the locations of known major prospects and drill collar locations for the Ambler District including NovaCopper-controlled Ambler and Bornite sequence targets.

Table 10.7 Drill, Metreage and Average Drill Depth for NovaCopper Ambler Sequence VMS Targets

Area	Drill Holes	Metres	Average Depth
Arctic Deposit	144	31,179	217
Dead Creek/West Dead Creek	21	3,470	165
Sunshine/Bud	36	7,111	198
Snow	11	1,527	139
Horse/Cliff/DH	22	2,277	104
Red/Nora/BT	18	2,399	133
Total	252	47,964	190

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Table 10.8 Significant Drill Intercepts – NovaCopper Ambler Sequence Prospects

Prospect	Drill Hole	Length (m)	Au (ppm)	Ag (ppm)	Cu (%)	Pb (%)	Zn (%)
BT	BT-4*	2.56	N/A	33.4	1.86	0.91	2.23
	BT-6*	1.98	N/A	41.9	1.18	0.77	2.57
	BT-7*	3.2	N/A	52.3	2.92	1.56	4.39
Bud	BUD-03	2.87	0.58	67.8	1.69	0.41	1.45
	BUD-04	1.47	0.60	51.9	1.08	0.60	1.44
Cliff	CLF-01*	18.74	0.03	108.4	0.32	0.84	2.79
	CLF-02	7.32	0.04	23.7	0.44	1.15	3.50
	CLF-03	3.41	0.15	64.5	1.43	1.48	5.00
	CLF-04*	19.81	N/A	N.A.	0.39	0.67	2.46
	CLF-05	10.97	N/A	N.A.	0.23	0.64	2.50
	CLF-06	3.11	N/A	38.6	0.29	1.29	2.39
	CLF-07	5.88	N/A	31.5	0.36	0.79	1.63
	CLF-08*	9.67	N/A	50.7	0.55	1.44	2.91
	CLF-10	3.96	N/A	61.5	0.66	1.49	2.64
	CLF-11	8.05	N/A	18.3	0.68	0.70	1.54
Dead Creek	AR9803-01	2.71	0.10	22.4	0.52	0.23	1.27
	DC-01	2.28	N/A	37.7	4.47	N.A.	1.77
	DC-02*	2.59	N/A	51.7	1.66	N.A.	2.01
	DC-03*	4.26	0.12	40.6	3.13	0.07	1.05
	DC-04*	3.22	N/A	67.5	1.39	0.27	1.13
	DC-05	4.27	0.36	95.6	2.82	0.23	3.67
	DC-06	4.57	0.06	15	0.96	N.A	0.31
	DC-07*	3.97	N/A	87.7	0.70	N.A.	2.71
	DC-08	2.41	N/A	73.6	0.12	N.A>	3.68
	DC-11	1.34	0.06	64.3	0.14	1.26	3.78
DH	DH-02*	20.12	0.20	35.0	0.54	1.90	4.71
	DH-03	4.57	0.08	20.1	0.13	0.45	1.01
	DH-05	9.14	N/A	29.3	0.33	1.3	2.01
Horse	HC-02	4.72	N/A	14.1	1.41	0.47	3.57
Nora	NORA-01	1.68	0.14	17.5	1	0.58	2.94
	NORA-07	3.44	0.04	6.9	2.81	0.01	0.01
	NORA-08	2.9	0.07	0.9	1.21	0.01	0.01
Red	AR06-101	1.05	0.06	8.3	0.04	0.37	2.02
Snow	SNO-21	4.73	3.26	430.2	0.32	3.56	6.26
	SNO-23	1.22	0.07	10.3	0.02	1.70	0.97
	SNO-24	2.44	1.04	210.9	0.63	3.50	4.69

table continues…

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Prospect	Drill Hole	Length (m)	Au (ppm)	Ag (ppm)	Cu (%)	Pb (%)	Zn (%)
Sunshine	SC-01*	19.65	0.04	19.3	1.41	0.28	2.01
	SC-02*	18.04	0.07	19.8	1.14	0.38	3.06
	SC-03*	13.91	0.09	28.5	1.29	0.51	2.18
	SC-04*	14.17	0.87	33.9	0.94	0.77	3.17
	SC-05*	19.51	0.05	37.9	1.27	0.51	2.51
	SC-06*	10.67	0.06	35.1	2.16	0.68	3.46
	SC-08*	14.49	0.11	33.6	1.78	0.99	2.59
	SC-11*	6.03	0.04	8.4	0.83	0.28	1.33
West Dead Creek	WDC-04	3.81	N/A	21.9	0.44	0.48	1.21
	WDC-05*	5.64	0.12	81.8	0.19	0.39	1.78
	WDC-07	6.09	N/A	47.6	0.23	0.74	1.94
	WDC-10	2.16	0.138	43.2	0.37	0.76	2.11

Notes:	*Weighted sum or more than one interval.
	Composites based on 1.0% copper-equivalent cut-off grade, 1 m minimum composite, and up to 2 m.

Figure 10.3 Known Collar Locations and Principal Target Areas – Ambler District

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NOVACOPPER TENURE

There have been only two drill campaigns (2006 and 2012) as shown in Table 10.2 by NovaCopper during their tenure targeting additional prospects beyond Arctic in the Ambler Schist belt. Exploration in 2006 targeted a series of geophysical anomalies in the central portion of the Ambler Schist belt near to Arctic. Twelve holes totalling 3,100 m were drilled. In 2012, NovaCopper drilled an additional 4 holes totalling 1,752 m to explore the down dip extension of the Sunshine deposit. Both programs are summarized in Table 10.9 and Figure 10.4 shows the Sunshine Prospect and drill hole locations.

Table 10.9 NovaCopper Exploration Drilling – Ambler Schist Belt

Hole ID	Area	Target	UTM East	UTM North	Azimuth (°)	Dip (°)	Depth (m)
AR06-98	COU	EM Anomaly	609490	7454374	0	-90	712.6
AR06-99	98-3	EM Anomaly	610111	7458248	0	-90	420.0
AR06-100	98-3	EM Anomaly	609989	7458633	0	-90	225.6
AR06-101	Red	EM Anomaly	618083	7451673	0	-90	141.7
AR06-102	Sunshine	West Extension	601176	7457834	30	-65	97.8
AR06-103	Red	EM Anomaly	618073	7451806	0	-90	209.7
AR06-104	Red	EM Anomaly	617926	7451693	0	-90	183.2
AR06-105	Red	EM Anomaly	618074	7451537	0	-90	136.6
AR06-106	Red	EM Anomaly	618083	7451677	310	-60	185.0
AR06-107	Sunshine	West Extension	601018	7458119	30	-60	294.4
AR06-108	Dead Creek	Downdip Extension	607618	7458406	0	-90	289.0
SC12-014	Sunshine	Sunshine Extension	601948	7457759	20	-57	537.8
SC12-015	Sunshine	Sunshine Extension	601860	7457637	20	-65	477.0
SC12-016	Sunshine	Sunshine Extension	601649	7457637	45	-77	386.2
SC12-017	Sunshine	Sunshine Extension	602063	7457701	20	-60	351.1

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Figure 10.4 Sunshine Prospect and Drill Hole Locations

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11.0 SAMPLE PREPARATION,ANALYSES,ANDSECURITY

11.1	SAMPLE PREPARATION
11.1.1	CORE DRILLING SAMPLING

The data for the Arctic Deposit resource was generated over three primary drilling campaigns: 1966 to 1986 when BCMC, a subsidiary of Kennecott Copper Corporation was the primary operator, 1998 when Kennecott Minerals resumed work after a long hiatus, and 2004 to present with NovaGold Resources Inc. and now NovaCopper Inc. as the operators.

KENNECOTT AND BCMC

Sampling of drill core prior to 1998 by BCMC focused primarily on the mineralized zones; numerous intervals of weak to moderate mineralization were not sampled during this period. During the 1998 campaign, Kennecott did sample some broad zones of alteration and weak mineralization, but much of the unaltered and unmineralized drill core was left unsampled. Little documentation on historic sampling procedures is available.

NOVAGOLD AND NOVACOPPER TENURE

Between 2004 and 2006, NovaGold conducted a systematic drill core re-logging and re-sampling campaign of Kennecott and BCMC era drill holes AR-09 to AR-74. NovaGold either took 1 to 2 m samples every 10 m, or sampled entire lengths of previously unsampled core within a minimum of 1 m and a maximum or 3 m intervals. The objective of the sampling was to generate a full ICP geochemistry dataset for the Arctic Deposit and ensure continuous sampling throughout the deposit. Sample preparation procedures for NovaGold era work are described in the following subsection. Quality assurance/quality control (QA/QC) review of historic sampling is described in Section 11.4.

All drill core was transported by helicopter in secure core “baskets” to either the Dahl Creek camp or the Bornite camp for logging and sampling. Sample intervals were determined by the geologist during the geological logging process. Sample intervals were labelled with white paper tags and butter (aluminum) tags which were stapled to the core box. Each tag had a unique number which corresponded to that sample interval.

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Sample intervals were determined by the geological relationships observed in the core and limited to a 3 m maximum length and 1 m minimum length. An attempt was made to terminate sample intervals at lithological and mineralization boundaries. Sampling was generally continuous from the top to the bottom of the drill hole. When the hole was in unmineralized rock, the sample length was generally 3 m, whereas in mineralized units, the sample length was shortened to 1 to 2 m.

Geological and geotechnical parameters were recorded based on defined sample intervals and/or drill run intervals (defined by the placement of a wooden block at the end of a core run). Logged parameters were reviewed annually and slight modifications have been made between campaigns, but generally include rock type, mineral abundance, major structures, SG, point load testing, recovery and rock quality designation measurements. Drill logs were converted to a digital format and forwarded to the Database Manager, who imported them into the master database.

Core was photographed and then brought into the saw shack where it was split in half by the rock saw, divided into sample intervals, and bagged by the core cutters. Not all core was oriented; however, core that had been oriented was identified to samplers by a line drawn down the core stick. If core was not competent, it was split by using a spoon to transfer half of the core into the sample bag.

Once the core was sawed, half was sent to ALS Chemex Laboratories (ALS Chemex) in Vancouver for analysis and the other half was initially stored at the Dahl Creek camp but has been consolidated at the storage facility at the Bornite camp facilities or at NovaCopper’s warehouse in Fairbanks.

Shipment of core samples from the Dahl Creek camp occurred on a drill hole by drill hole basis. Rice bags, containing two to four poly-bagged core samples each, were marked and labelled with the ALS Chemex address, project and hole number, bag number, and sample numbers enclosed. Rice bags were secured with a pre-numbered plastic security tie and a twist wire tie and then assembled into sling loads for transport by chartered flights on a commercial airline to Fairbanks, where they were met by a contracted expeditor for deliver directly to the ALS Minerals preparation facility in Fairbanks. In addition to the core, control samples were inserted into the shipments at the approximate rate of one standard, one blank and one duplicate per 20 core samples:

• Standards: four standards were used at the Arctic Deposit. The core cutter inserted a sachet of the appropriate standard, as well as the sample tag, into the sample bag.

• Blanks: were composed of an unmineralized landscape aggregate. The core cutter inserted about 150 g of blank, as well as the sample tag, into the sample bag.

• Duplicates: the assay laboratory split the sample and ran both splits. The core cutter inserted a sample tag into an empty sample bag.

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Samples were logged into a tracking system on arrival at ALS Chemex, and weighed. Samples were then crushed, dried, and a 250 g split pulverized to greater than 85% passing 75 μm.

Gold assays were determined using fire analysis followed by an atomic absorption spectroscopy (AAS) finish. The lower detection limit was 0.005 ppm gold; the upper limit was 1,000 ppm gold. An additional 34-element suite was assayed by inductively coupled plasma-atomic emission spectroscopy (ICP-AES) methodology, following nitric acid aqua regia digestion. The copper analyses were completed by atomic absorption (AA), following a triple acid digest.

11.1.2 ACID-BASE ACCOUNTING SAMPLING

In 1998, a broad assessment of ARD at the Arctic Deposit (Robertson 1998) was conducted with a focus on characterization for surface development. Criteria used for assessing and classifying ARD potential have since been modified.

In 2010, SRK collected 148 samples from drill core based on their position relative to the massive and semi-massive sulphide mineralization (SRK 2011). Samples were targeted within, immediately adjacent to, adjacent to, and between lenses of mineralization; the sampling program focused on characterization for underground development. Samples were shipped to SGS Canada Inc., Burnaby, BC, for sample preparation and analysis. Samples were analyzed for ABA and metals. ABA tests were conducted using the Sobek method with sulphur speciation and total inorganic carbon (TIC) analysis. Metal concentrations were determined using aqua regia digestion followed by ICP-MS analysis. In addition barium and fluorine were analyzed by x-ray fluorescence (XRF) following a lithium metaborate fusion.

11.1.3 DENSITY DETERMINATIONS

Representative SG determinations conducted before 1998 for the Arctic Project are lacking. Little information regarding sample size, sample distribution and SG analytical methodology are recorded for determinations during this period.

In 1998, Kennecott collected 38 core samples from that year’s drillcore, of which 22 were from mineralized zones and 16 from non-mineralized lithologies. Mineralized samples were defined as MS (more than 50% total sulphides), SMS (less than 50% total sulphides) or lithology samples (non-mineralized country rock containing up to 10% sulphides). SG determinations were conducted by ALS Chemex and Golder and Associates and were based on short (6 to 12 cm) whole core samples and determined based on the water displacement method.

In 1999, Kennecott collected 231 samples from pre-1998 drill core for SG analysis. The samples were from NQ- and BQ-sized core and averaged 7.27 cm in length. The samples were shipped to Anchorage but were not forwarded to a lab for further analysis.

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In 2004, NovaGold forwarded the 231 samples from the pre-1998 drill campaigns, stored in Kennecott’s Anchorage warehouse, as well as 33 new samples from the 2004 drill program, to Chemex Laboratories for SG analysis.

Additionally, in 2004 NovaGold collected 127 usable field SG measurements. Samples were collected from HQ-sized core and averaged 9.05 cm in length. An Ohaus Triple Beam Balance was utilized to determine a weight-in-air value for dried core, followed by a weight-in-water value. The wet-value was determined by suspending the sample by a wire into a water-filled bucket. The SG value was then calculated using the following formula:

Weight in air
[Weight in air – Weight in water]

In 2011, NovaGold geologists stopped collecting short interval “point data” (as described above) within the mineralized zone, and instead collected “full-assay-width” determinations from existing 2008 split core and all of the sampled 2011 whole core. The samples averaged 1.69 m in length. Samples were collected continuously within mineralized zones and within a 2 to 3 m buffer adjacent to mineralized zones. Two hundred sixty-six sample pulps were also submitted to ALS Chemex for SG determination by pycnometer analysis. In total, 459 valid SG determinations were collected, ranging from 2.64 to 4.99.

11.2 SECURITY

Security measures taken during historical Kennecott and BCMC programs are unknown to NovaGold or NovaCopper. NovaCopper is not aware of any reason to suspect that any of these samples have been tampered with. The 2004 to 2011 samples were either in the custody of NovaGold personnel or the assay laboratories at all times, and the chain of custody of the samples is well documented.

11.3 ASSAYING AND ANALYTICAL PROCEDURES

The laboratories used during the various exploration, infill, and step-out drill analytical programs completed on the Arctic Project are summarized in Table 11.1.

ALS Chemex has attained International Organization for Standardization (ISO) 9001:2000 registration. In addition, the ALS Chemex laboratory in Vancouver is accredited to ISO 17025 by Standards Council of Canada for a number of specific test procedures including fire assay of gold by AA, ICP and gravimetric finish, multi-element ICP and AA assays for silver, copper, lead and zinc.

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Table 11.1 Analytical Laboratories Used by Operators of the Arctic Project

Laboratory Name	Laboratory Location	Years Used	Accreditation	Comment
Union Assay Office, Inc.	Salt Lake City, Utah	1968	Accreditations are not known.	Primary Assay Lab
Rocky Mountain Geochemical Corp.	South Midvale, Utah	1973	Accreditations are not known.	Primary and Secondary Assays
Resource Associates of Alaska, Inc.	College, Alaska	1973, 1974	Accreditations are not known.	Primary and Secondary Assays
Georesearch Laboratories, Inc.	Salt Lake City, Utah	1975, 1976	Accreditations are not known.	Primary and Secondary Assays
Bondar-Clegg & Company Ltd.	North Vancouver BC	1981, 1982	Accreditations are not known.	Primary and Secondary Assays
Acme Analytical Laboratories Ltd. (AcmeLabs)	Vancouver, BC	1998, 2012, 2013	Accreditations are not known.	2012 and 2013 Secondary Check Sample Lab
ALS Analytical Lab	Fairbanks, Alaska	1998, 2004, 2005, 2006, 2012, 2013	In 2004, ALS Chemex held ISO 9002 accreditations but changed to ISO 9001 accreditations in late 2004. ISO/International Electrotechnical Commission (IEC) 17025 accreditation was obtained in 2005.	2012 and 2013 Primary Assay Lab

11.4	QUALITY ASSURANCE/QUALITY CONTROL
11.4.1	CORE DRILLING SAMPLING QA/QC

Previous data verification campaigns were limited in scope and documentation and are described by SRK (2012).

During 2013, NovaCopper conducted a 26% audit of the NovaGold era assay database fields: sample interval, Au, Ag, Cu, Zn, and Pb. This audit is documented in a series of memos (Brown 2011; West 2013). NovaCopper staff did not identify and/or correct any transcription and/or coding errors in the database prior to resource estimation. NovaCopper also retained independent consultant Caroline Vallat, P.Geo. of GeoSpark Consulting Inc. (GeoSpark) to: 1) re-load 100% of the historical assay certificates, 2) conduct a QA/QC review of paired historical assays and NovaGold era re-assays; 3) monitor an independent check assay program for the 2004 to 2008 and 2011 drill campaigns; and 4) generate QA/QC reports for the 2004 to 2008 and 2011 drill campaigns. Below is a summary of the results and conclusions of the GeoSpark QA/QC review.

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NOVAGOLD QA/QC REVIEW ON HISTORICAL ANALYTICAL RESULTS

During 2004, NovaGold conducted a large rerun program and check sampling campaign on pre-NovaGold (pre-2004) drill core. The 2004 and 2005 ALS Chemex Laboratories primary sample results have been assigned as the primary assay results for the Arctic Project in the database, amounting to 1,287 of the total 3,186 primary samples related to pre-NovaGold drill holes.

During 2013, GeoSpark conducted a QA/QC review of available QA/QC data (20130422 – QAQC on Pre-NovaGold Arctic Assays); including sample pair data amounting to 422 data pairs which is 11% relative to the primary sample quantity. The sample pairs included original duplicates, original repeat assays, 2004 rerun assays on original sample pulps analyzed secondarily at ALS Chemex, and check samples from 2004 on original samples re-analyzed at ALS Chemex.

The review found that the available QA/QC data is related to drill holes that are spatially well distributed over the historic drill hole locations.

Figure 11.1 Spatial Availability of QA/QC Data

Review of Precision

A comparison of the original analytical results with the secondary results serves to infer the level of precision within the original results. Also, the 2004 rerun sample results and the check sample pair results from 2004 and 2005 were compared to the original assays to infer the level of repeatability or precision within the original results.

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The result of the average relative difference (AD) review on sample pairs found satisfactory to good inferred precision levels for all of the sample pairs and elements except for the 2004 rerun sample lead results. For the lead 2004 rerun sample pairs there were 66.85% of the pairs less than the 1 AD limit, inferring poor precision in the original results. Overall, the lead values were found to pass the AD criteria for the original duplicates, original repeats, and check sample reviews. More insight was made regarding the lead precision upon review of the data pairs graphically within scatter plots and Thompson-Howarth Precision Versus Concentration (THPVC) plots. The 2004 rerun sample lead values were found to infer a poor-to-moderate level of precision and an indication that the original results might be of negative bias where the original results may have been reported on average 0.2% less than their true values for grades of 0.5% lead and higher. However, the original duplicate, original repeats, and check samples inferred that there was a moderate or satisfactory level of correlation within the lead values. Furthermore, the overall inference of precision in the lead values has been defined as moderate.

The detailed review of the gold pairs inferred an overall moderate level of precision within the original analytical results.

The silver, copper, and zinc analytical pair review found overall inferred strong precision in the original analytical results.

It is GeoSpark's opinion that the detailed review of analytical pair values reported for gold, silver, copper, lead and zinc has inferred an overall acceptable level of precision within the original sample analytical results for the pre-NovaGold Arctic Project.

Review of Accuracy

The rerun sample program of 2004 included analysis of 53 QA/QC materials comprising 20 standards and 33 blanks. These standards and blanks were reviewed in order to indirectly infer the accuracy within the original sample data.

The 2004 rerun samples on original pulps also included analysis of standards and blanks with the primary samples. These results have been reviewed using control charts for review of the inferred accuracy within the 2004 rerun sample results; in addition, the inferred rerun sample accuracy is related to the accuracy of the original results in that comparison of the original results to the 2004 reruns and has been shown to be acceptable overall.

The blank results were reviewed for gold, silver, copper, lead, and zinc and it has been inferred that there is good accuracy within the results and that there was no significant issue with sample contamination or instrument calibration during the analysis.

The standard results were reviewed for gold, silver, copper, lead, and zinc. The reported control limits were available for silver, copper, lead, and zinc. The gold control limits were calculated for the review.

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In addition upon initial review, the zinc control limits were also calculated from the available data to provide a more realistic range of control values for the results. The gold, silver, and copper results were inferred to be of strong accuracy. The lead and zinc results were inferred to be of moderate accuracy overall.

It was GeoSpark’s opinion that the review for accuracy has found an acceptable level of inferred accuracy within the gold, silver, copper, lead, and zinc results reported for the 2004 rerun samples and indirectly within the original results.

Review of Bias

There were 35 check samples on original samples re-assayed at ALS Chemex during 2004. These were reviewed for an indication of bias in the original results. Additionally, the 2004 rerun sample results have been reviewed for inference of bias in the original results.

Overall, the detailed review of the check sample pair gold concentrations has found minor positive bias in the 2004 pairs and minor positive bias in the 2005 pairs. The level of bias is inferred to be at very near zero with the original being reported approximately 0.005 greater than the 2004 results reported by ALS Chemex. The 2004 rerun samples compared to the originals has inferred negligible bias in the original gold results. It is GeoSpark's opinion that these levels of inferred bias are not significant to merit concern with the overall quality of gold values reported for the pre-NovaGold Arctic Project.

The detailed review of the check sample silver pairs has found minor negative bias implied by the 2004 check sample pairs. The 2004 rerun samples have shown a negligible amount of bias in the original results. It is GeoSpark’s opinion that overall the bias in original silver concentrations is inferred to be negligible to minor negative but not significant to merit concern of the overall quality of the silver results.

The copper check samples reported in 2004 were found to have a few anomalous results that were implying significant positive bias. However, a more detailed review found that the exclusion of the anomalous pairs resulted in a minor positive bias overall. The 2004 rerun sample copper results have shown that there is a possibility for positive bias in the original copper grades at concentrations greater than 5%. Overall, it is GeoSpark’s opinion that the bias inferred within the original copper results is not significant to merit concern with the original assay quality.

The 2004 check sample review inferred overall small negative bias in the original lead results. The 2004 rerun sample data also inferred that there was a small negative bias in the original results for grades over 0.5%. Overall, it is GeoSpark’s opinion that this detailed review has inferred that the levels of inferred bias within the lead concentrations are not significant enough to merit concern over the original result quality.

The original zinc results have been inferred to be of very minor positive bias when the 2004 check sample pairs (excluding three anomalous pairs) are reviewed. The 2004 rerun sample zinc values have been shown to be very comparable with the originals and a negligible amount of bias can be inferred in the original zinc concentrations. Furthermore, this detailed bias review has inferred that there is no significant bias in the original zinc results for the pre-NovaGold Arctic Project.

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Conclusion

The pre-NovaGold Arctic Project database analytical results have been verified and updated to provide a good level of confidence in the database records.

It is GeoSpark’s opinion that with consideration of the historic nature of the Arctic Project, a sufficient amount of QA/QC data and information has been reviewed to make a statement of the overall pre-NovaGold Arctic Project analytical result quality.

It is GeoSpark’s opinion that this detailed review has inferred that the pre-NovaGold Arctic Project analytical results are of overall acceptable quality.

QA/QC REVIEW ON NOVAGOLD (2004 TO 2013) ANALYTICAL RESULTS

During 2013, GeoSpark conducted a series of QA/QC reviews on NovaGold’s 2004 to 2013 analytical results. These QA/QC reviews serve to infer the precision of the NovaGold Arctic Project analytical results through a detailed analytical and statistical review of field duplicate samples; serve to infer the accuracy of the analytical results through a review of the standards and blanks inserted throughout the NovaGold programs; and serve to define any bias in the primary sample results through a review of secondary lab checks at AcmeLabs in Vancouver, BC.

The QA/QC reviews are documented in a series of memos (Vallat 2013c, 2013d, 2013e, 2013f, 2013g, 2013h). The reviews are summarized in the following subsections by year of campaign.

2004

The 2004 exploration program at the Arctic Project included drilling and sampling related to 11 drill holes AR04-0078 through AR04-0088, amounting to 989 primary samples assayed within 61 assay certificates reported by ALS Chemex in Fairbanks, Alaska.

The field duplicate pairs were reviewed analytically using an AD guideline to gauge the inferred level of precision within the results. This review found that the gold, silver, copper, lead and zinc grades were reported with less than 0.3 AD for at least 75% of the sample pairs. This shows strong repeatability or precision throughout.

In addition, scatter plots and THPVC plots were reviewed. The scatter plots showed moderate to strong precision within the gold grades, and strong precision within the silver, copper, lead, and zinc grades reported by ALS Chemex for the 2004 Arctic Project. The THPVC review found an inferred poor level of repeatability within the gold results, but further review showed that the precision percent was exaggerated due to the low gold grades reported for the samples. It is GeoSpark’s opinion that the THPVC review of the gold is an unreliable measure of the precision due to the low grades and that the earlier analytical tests and scatter plot results are more representative of the inferred precision for the gold results.

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The THPVC review found very strong repeatability of precision within the silver, copper, lead, and zinc concentrations reported by ALS Chemex for the 2004 Arctic Project.

Overall, the precision has been inferred to be strong for the gold, silver, copper, lead, and zinc concentrations reported by ALS Chemex for the 2004 Arctic Project.

Overall, the analytical results of analysis for gold reported by ALS Chemex for the 2004 Arctic Project have been inferred to be of strong accuracy. The silver, copper, lead, and zinc values have been inferred to have moderate or satisfactory accuracy. In addition, the review has shown no significant ongoing issues with sample contamination or instrument calibration.

The check sample review has found no bias inferred within the gold and silver grades reported for the 2004 Arctic Project. A small level of positive bias was inferred within the copper, lead, and zinc results reported on high-grade samples. The copper and lead bias may be attributable to specific details of the assay methodology. The zinc bias is more likely a reflection of a lack of repeatability at high grades. It is GeoSpark’s opinion that overall the levels of bias are not significant enough to merit concern with the sample result quality.

2005

The 2005 exploration program at the Arctic Project included drilling and sampling related to nine drill holes labelled AR05-0089 through AR05-0097, amounting to 1,228 primary samples assayed within 36 assay certificates reported by ALS Chemex in Fairbanks, Alaska.

The review of field duplicates, blanks and standards, and check samples has allowed for inference of a reasonable level of precision, good accuracy, and insignificant levels of bias within the primary sample results reported by ALS Chemex related to the 2005 Arctic Project.

This detailed QA/QC review on the analytical results reported for the 2005 Arctic Project has allowed for overall confidence in the analytical result quality.

The analytical results can be inferred to be of sufficient quality to represent the Arctic Project.

2006

The 2006 exploration program at the Arctic Project included drilling and sampling related to 12 drill holes labelled AR06-98 through AR06-109, amounting to 1,175 primary samples analyzed at ALS Chemex.

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The review of field duplicates, blanks and standards, and check samples for the 2006 Arctic Project has allowed for inference of a good level of precision, good accuracy, and insignificant levels of bias within the primary sample results reported by ALS Chemex related to drill holes AR06-98 through AR06-109.

The analytical results can be inferred to be of sufficient quality to represent the Arctic Project.

2007

The 2007 exploration program at the Arctic Project included drilling and sampling related to four drill holes labelled AR07-110 through AR07-113, amounting to 950 primary samples analyzed at ALS Minerals.

The review of field duplicates, blanks and standards, and check samples for the 2007 Arctic Project has allowed for inference of a good level of precision, good accuracy, and insignificant levels of bias within the primary sample results reported by ALS Minerals related to drill holes AR07-110 through AR07-113.

The analytical results can be inferred to be of sufficient quality to represent the Arctic Project.

2008

The 2008 exploration program at the Arctic Project included drilling and sampling related to 14 drill holes labelled AR08-0114 through AR08-0126 and also drill hole AR08-0117w, amounting to 1,406 primary samples assayed within 44 assay certificates reported by ALS Chemex in Fairbanks, Alaska.

The review of field duplicates, blanks and standards, and check samples for the 2008 Arctic Project has allowed for inference of a reasonable level of precision, good accuracy, and insignificant levels of bias within the primary sample results reported by ALS Minerals related to drill holes AR08-0114 through AR08-0126.

The analytical results can be inferred to be of sufficient quality to represent the Arctic Project.

2011 (Analyzed in 2013)

For the assay certificates FA13021131, FA13021132, FA13021133, FA13021134, and FA13021135 there were six field duplicate pairs, six blank instances, and three standard instances available for review of the QA/QC of the reported results.

The duplicates for gold, silver, copper, lead, and zinc were found to correlate well with the primary sample results and it can be inferred that the primary results are of good precision.

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Each of the blank instances of analysis was returned within the control limits for the material. Issues with sample contamination and instrumentation difficulties can be ruled out. In addition the accuracy can be inferred to be strong.

The standard instances of analysis were each retuned within the acceptable range for gold, silver, copper, lead, and zinc; it is inferred that there is strong accuracy within the reported primary sample assay results.

The detailed review of secondary lab check sample results reported by ALS Chemex or the 2011 drill holes assayed in 2013 and reported within the defined analytical certificates has shown that for the gold, silver, copper, lead, and zinc results there is no need to be concerned with the overall quality of the results and any indication of bias in the results is not significant to the result quality.

The assays within the certificates reviewed by GeoSpark can be inferred to be of good quality to represent the Arctic Project.

11.4.2 ACID-BASE ACCOUNTING SAMPLING QA/QC

SRK conducted a QA/QC review of the 2010 ABA dataset for the Arctic Project in March 2011. The memo entitled “Preliminary ML/ARD Analysis Ambler District Arctic Deposit, Alaska”, located in NovaCopper’s Document Management System (DMS), discusses the results of the ABA review and documents the 33 duplicate ABA analyses on the lab certificates.

11.4.3 DENSITY DETERMINATIONS QA/QC

A QA/QC review of the SG dataset for the Arctic Project was conducted by NovaCopper staff in March 2013. The memo entitled “Arctic_Specific Gravity Review_A.West_20130326”, located in NovaCopper’s DMS, discusses the results of the QA/QC review and is summarized in the following subsections.

LAB VERSUS FIELD DETERMINATIONS

SG lab determinations conducted during 2004 produced significantly lower average SG results for the mineralized zone than the 1998 and 2004 average field determinations. In the same test, lithology samples outside the mineralized zone produced comparable values. The difference between the averaged 1998 and 2004 lab results and those from field studies may be the result of selection bias, limited population size, and sample length. Paired lab and field determinations from the 2004 program show very low variation.

In 2010, to check the validity of the wet-dry measurements on the Arctic Deposit core with respect to possible permeability of the core samples, NovaGold measured 50 unwaxed samples representing a full range of SG values for a variety of lithologies and then submitted the samples to ALS Chemex for wet-dry SG determinations after being sealed in wax. The mean difference between the NovaGold unwaxed and the ALS Chemex waxed SG determinations was 0.01 g/cm³.

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In 2011, to check the accuracy of the wet-dry measurements, the SG for 266 pulps was determined by pycnometer by ALS Chemex (ALS code OA-GRA08b). Figure 11.2 shows that the two methods compare favourably, with the wet-dry measurements displaying a very slight low bias. Generally, wet-dry measurements are considered the more acceptable method for accurate SG determinations since they are performed on whole (or split) core that more closely resembles the in-situ rock mass.

Figure 11.2 Graph Showing Good Agreement between Wet-dry Measured Specific Gravity and Pycnometer Measured Specific Gravity

STOICHIOMETRIC METHOD

Full sample length determinations can be directly compared to the assay results for copper, zinc, lead, iron, and barium that are the major constituents of the sulphide and sulphate species for the Arctic Deposit. This allows NovaCopper to check the wet-dry measurements by estimating the SG for an ideal stoichiometric distribution of the elements into sulphide and sulphate species.

Stoichiometric SG values were estimated for 279 sample intervals from 2008 drillcore that had both measured SG values and total digestion XRF barium values. Figure 11.3 compares the estimated stoichiometric SGs to the measured SGs. Overall, there is a very good correlation between the two SG populations (R2 of 0.9671), though stoichiometric estimates are slightly lower with increasing SG. Using slightly different compositional values for the assorted sulphide and sulphate species, and assuming a 1:1 ratio of weight percent iron to weight percent copper in chalcopyrite (the molar value is 1:1), the stoichiometric equation yields SGs that have an even better correlation (R2=0.9726), due to partitioning more iron into less dense chalcopyrite which leaves less iron available for more dense pyrite, essentially correcting the bias for the lack of estimated iron-bearing silicates.

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Figure 11.3 Measured versus Stoichiometric Specific Gravities

MULTIPLE REGRESSIONS METHOD

The positive comparisons/correlations of our measured SG values to the laboratory determined values and to the stoichiometric estimated values gives us high confidence in our wet-dry measurements. As a result, a multiple regression analysis can be performed using the assay data to get a best fit to the measured SGs. This may correct for the varying residencies of Fe and Ba (and also for the varying density within sphalerite due to the Zn:Fe ratio).

The best fit to the data was achieved by using the multiple regression tool in Microsoft Excel on Ba, Fe, Zn and Cu for the entire dataset (Figure 11.4). The estimate correlates very well (R2=0.9678) with observed data and has a sinusoidal pattern that fits the low and moderately high SG very well and has high bias for moderate SG values and a low bias for very high SG values. The resultant SG formula is as follows:

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SG(Regression)= 2.567 + 0.0048*Cu (wt%) + 0.045*Fe (wt%) + 0.032*Ba (wt%) + 0.023%*Zn (wt%)

Figure 11.4 Scatter Plot Showing the Measured Specific Gravity versus Multiple (Copper, Iron, Zinc, Barium) Regression Estimate

DENSITY DETERMINATIONS PERFORMANCE

The SG of a field sample interval can be reproduced in the lab or estimated from assay values using either a stoichiometric method which assumes a fixed metal residency in certain sulphide and sulphates or by a multiple regression method that empirically fits measured data. Overall, what this QA/QC analysis suggests is that the measured SG values can be replicated by various methods, thus supporting the quality of the measured SG data.

11.5 AUTHOR’S OPINION

Tetra Tech believes the database meets or exceeds industry standards of data quality and integrity. Tetra Tech believes the sample preparation, security and analytical procedures are adequate to support resource estimation and completion of a PEA.

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12.0 DATA VERIFICATION

Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM, and one of the QPs of this NI 43-101 PEA, conducted a site visit from June 19 to 21, 2013 in support of this technical report. Verifications are based on data collected during the site visit, and a desktop study review.

12.1 DRILL HOLE COLLAR VERIFICATION

Nine drill hole collars (AR-03, AR-04, AR-10, AR-44, AR-47, AR-64, AR05-0094, AR05-0097 and AR-40) were located using a Garmin Etrex 20 GPS unit. The offset distances between the collar coordinates reflected in the drill hole database provided by NovaCopper and the measured positions range from 3.4 to 7.8 m with an average offset of 4.8 m. This range is within the tolerance to be expected from GPS measurements and Tetra Tech considers the collar positions to be adequately located to form the basis of resource estimation work.

12.2 TOPOGRAPHY VERIFICATION

Tetra Tech conducted two foot traverses over representative areas of the Arctic Deposit. Continuous GPS measurements were compiled during these traverses. The averages of these 724 spot height measurements within 10 m² by 10 m² areas were compared to the corresponding digital terrain model (DTM) survey points (Figure 12.1).

Figure 12.1 Distribution of the Differences Between GPS Elevations and the DTM

For the traverse data, 90% confidence limits are -0.73 m and +0.09 m. The vertical dimension of the estimation model block size is 5 m, which Tetra Tech considers to be acceptable for the block model estimation process.

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12.3 CORE LOGGING VERIFICATION

Tetra Tech visited the NovaCopper core storage facility in Fairbanks and reviewed three drill holes for lithology, mineraization and the quality of storage.

Core boxes were found to be in good condition and intervals were easily retrieved for the following drill holes:

• AR05-0092 (129 to 147 m)

• AR08-0117 (128 to 216 m)

• AR08-0126 (144 to 211 m).

Logged descriptions of massive and semi-massive sulphide mineralization and general sampling results corresponded to the appearance of the core for selected intervals. Tetra Tech considers that the logging style and data is acceptable for the geological modelling and resource estimation processes.

12.4 DATABASE VERIFICATION

Assay data in the database was verified against the corresponding ALS Chemex assay certificates (Table 12.1) by observation and independent spreadsheet comparison. Nine hundred and ten composite samples (analyzed for all five elements of interest: copper, zinc, lead, gold, and silver) were reviewed.

The NovaCopper drill database has been reviewed, and no significant concerns were noted. The results of previous data verification by an external QP (SRK 2012), completed for NovaCopper, were also reviewed. This previous data verification exercise included an extensive review of all NovaGold drilling as well as drilling completed by previous operators. Based on this review, Tetra Tech believes that the data verification completed on the NovaCopper dataset is sufficiently robust to support resource estimation.

Table 12.1 Assay Certificates Reviewed

	No. of
Certificate	Samples
FA04064633	20
FA04064636	20
FA04056401	20
FA04058933	20
FA04060274	20
FA04061994	20
FA04063778	20
FA04058925	20
EL08107514	40

table continues…

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Certificate	No. of Samples
EL08107517	40
EL08107519	40
FA04061933	20
FA04061934	20
FA04061935	20
FA04061993	20
FA04062661	20
FA04064634	20
FA04064635	20
FA04064637	20
FA04064639	20
FA04073763	20
FA04073764	20
FA04073765	20
FA04073768	20
FA05011929	50
FA05073602	40
FA05074433	40
FA05075793	40
FA05075795	40
FA05077735	40
FA05079117	40
FA05079118	40
FA13021133	40

12.5 QA/QC REVIEW

Standards, blanks, duplicates and check samples have been regularly submitted at a combined level of 20% of sampling submissions for all NovaGold/NovaCopper era campaigns. During the spring of 2013, GeoSpark conducted a thorough QA/QC review of pre-2004, 2004, 2005, 2006, 2007, 2008 and 2011 sampling campaigns which included review for accuracy, precision and bias (see Section 11.0). In addition to the QA/QC review, GeoSpark has been retained to provide ongoing database maintenance and QA/QC support.

Tetra Tech has reviewed the QA/QC dataset and reports and found the sample insertion rate and the timeliness of results analysis meets or exceeds industry best practices. The QA/QC results indicate that the assay results collected by NovaCopper, and previously by NovaGold, are reliable and suitable for the purpose of this study.

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12.6 QP OPINION

It is Tetra Tech’s opinion that the drill database for the Arctic Deposit is reliable and sufficient to support the purpose of this technical report and a current mineral resource estimate.

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13.0	MINERAL PROCESSING AND
	METALLURGICAL TESTING

13.1	METALLURGICAL TEST WORK REVIEW
13.1.1	INTRODUCTION

The Arctic Deposit is a stratiform polymetallic VMS deposit comprised of semi-massive and massive sulphides deposited in a highly variable metasedimentary and metavolcanic stratigraphy. Hydrothermal alteration has resulted in the development of footwall magnesium-rich alteration characterized by abundant chlorite and talc and hanging wall sodium-rich alteration characterized by paragonite. In the mineralized zone, the principal economic minerals are chalcopyrite, sphalerite, galena, and minor tetrahedrite and bornite. Metallurgical studies have spanned over 30 years with metallurgical test work campaigns undertaken at the Kennecott Research Center (KRC), Lakefield Research Ltd. (Lakefield) and SGS. Tetra Tech reviewed the test programs conducted on the deposit, which are listed in Table 13.1; the test results are summarized in the following sections.

Table 13.1 Metallurgical Test Work Programs

Year	Laboratory	Mineralogy	Grindability	Flotation
2012	SGS	√	√	Cu/Pb and Zn batch rougher and cleaner, Cu/Pb separation and locked cycle tests
1999	Lakefield	-	-	Cu/Pb and Zn batch rougher and cleaner, Cu/Pb separation
1976	KRC	-	√	Cu, Pb, Zn and Ag batch rougher flotation (selective flotation procedure)
1975	KRC	√	-	-
1972	KRC	-	-	Cu/Pb and Zn batch rougher and cleaner, Cu/Pb separation
1970	KRC	√	-	-

13.1.2 MINERALOGICAL AND METALLURGICAL TEST WORK – SGS 2012

INTRODUCTION

In 2012, SGS conducted a test program on the samples produced from mineralization zones 1, 2, 3, and 5 of the Arctic Deposit (Section 14.0). To the extent known, the samples are representative of the styles and types of mineralization and the mineral deposit as a whole. Drill core samples were composited from each of the zones into four different samples for the SGS test work which included process mineralogical examination, grindability parameter determination, and flotation tests.

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SGS used QEMSCAN^™, a quantitative mineralogical technique utilizing scanning electron microscopy to determine mineral species, species liberation and mineral associations in order to develop grade limiting/recovery relationships for the composites.

Standard Bond grindability tests were also conducted on five selected samples to determine the BWi and Ai.

The flotation test work investigated the effect of various process conditions on copper, lead and zinc recovery using copper-lead bulk flotation and zinc flotation followed by copper and lead separation.

TEST SAMPLES

The 2012 test program used 90 individual drill core sample intervals totaling 1,100 kg from the Arctic Deposit. Individual samples were combined into four composites representing different zones and labelled as Composites Zone 1 & 2, Zone 3, Zone 5, and Zone 3 & 5. The sample materials used in the 2012 test program at SGS were specifically obtained for metallurgical test purposes. The drill cores were stored in a freezer to ensure sample degradation and oxidation of sulphide minerals did not occur.

The head grades of the composites are shown in Table 13.2.

Table 13.2 Head Grades – Composite Samples – 2012

Sample ID	Cu (%)	Pb (%)	Zn (%)	Fe (%)	S (%)	Au (g/t)	Ag (g/t)	MgO (%)
Zone 1 & 2 A	2.66	0.93	3.48	7.92	8.53	0.79	57.1	5.77
Zone 1 & 2 B	2.60	0.96	3.38	7.54	8.18	0.78	58.0	5.79
Average	2.63	0.95	3.43	7.73	8.36	0.79	57.6	5.78
Zone 3 A	3.55	1.73	8.47	17.4	25.4	0.72	80.4	1.95
Zone 3 B	3.57	1.72	8.69	17.6	26.1	0.62	80.3	1.93
Average	3.56	1.73	8.58	17.5	25.8	0.67	80.4	1.94
Zone 3 & 5 A	4.45	1.64	7.81	16.8	23.6	1.01	81.7	3.86
Zone 3 & 5 B	4.37	1.55	7.7	16.5	23.4	0.93	82.2	4.05
Average	4.41	1.60	7.76	16.7	23.5	0.97	82.0	3.96
Zone 5 A	2.56	1.34	5.64	15.5	21.5	1.54	65.1	0.92
Zone 5 B	2.55	1.32	5.72	16.1	20.9	0.77	60.8	0.88
Average	2.56	1.33	5.68	15.8	21.2	1.16	63.0	0.90

MINERALOGICAL INVESTIGATION

SGS used QEMSCAN^™ to complete a detailed mineralogical study on each composite to identify mineral liberations and associations, and to develop grade/recovery limiting relationships for the samples. Head assays indicate that all four composite samples contain a considerable amount of magnesium oxide, implying the potential for significant talc which could impact flotation.

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The mineral modal abundance for the composites is shown in Table 13.3.

Table 13.3 Mineral Modal Abundance for Composite Samples – 2012

	Mass (%)
Mineral	Zone 1 & 2	Zone 3	Zone 3 & 5	Zone 5
Chalcopyrite	9.2	9.4	12.2	6.4
Bornite	0.02	0.01	0.03	0.4
Tetrahedrite	0.1	0.4	0.2	0.2
Antimony	0.03	0.2	0.005	0.3
Galena	1.3	2.1	2.1	2.1
Sphalerite	7.2	14.6	14.3	11.3
Pyrite	6.7	30.4	23.8	27.8
Pyrrhotite	2.2	0.2	0.2	1.4
Arsenopyrite	0.5	0.1	0.6	2.2
Other Sulphides	0.1	0.1	0.2	0.1
Quartz	30.2	8.6	9.0	16.6
Feldspar	0.9	0.2	0.4	0.3
Magnesium-Chlorite	11.9	3.4	2.8	1.1
Talc	2.0	0.8	6.3	0.1
Micas	14.2	1.9	7.0	9.4
Cymrite	3.5	3.9	1.8	1.9
Clays	0.6	0.05	0.2	0.1
IronOxides	0.3	0.3	0.5	0.3
Carbonates	3.4	1.3	4.2	2.0
Barite	3.0	21.8	13.4	14.5
Fluorite	1.7	0.1	0.4	1.2
Other	1.1	0.3	0.4	0.4
Total	100.0	100.0	100.0	100.0

The mineralogical study showed that the mineralogy of all four composites was similar. Each composite was composed mainly of pyrite, quartz, and carbonates. However, Composite Zone 1 & 2 contains approximately 30% quartz, compared to 8.6% for Composite Zone 3, and 16.6% for Composite Zone 5. The study also showed that Composite Zone 1 & 2 had the lowest pyrite content (6.7%) while Composites Zone 3 and Zone 5 contained approximately 30.4% and 27.8% pyrite, respectively.

In all four samples, the major floatable gangue minerals were talc and pyrite. Chalcopyrite was the main copper carrier. Combined bornite, tetrahedrite, and other sulphides (mainly enargite) accounted for less than 5% of the copper minerals in the Zone 1 & 2, Zone 3, and Zone 3 & 5 composites. In the Zone 5 sample, a slightly higher amount of bornite accounted for approximately 9% of the copper minerals. Galena was the main lead mineral (1.3% in the Zone 1 & 2 composite, and 2.1% in the other three composites) and sphalerite was the main zinc mineral (7.2% in Zone 1 & 2 composite and 11 to 14% in the other three composites).

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All the composites contained a significant amount of talc, which may have the potential to consume reagents and dilute final concentrates. Therefore, SGS recommended that talc removal using flotation be employed prior to base metal flotation.

At a grind size of approximately 90% passing 150 µm (ranging from 94.5 to 89% passing 150 µm), chalcopyrite liberation ranged from approximately 80 to 87% (free and liberated combined) for all composites. The chalcopyrite is mostly free, with 7 to 10% associated with pyrite. For all composites, galena liberation ranged from 54 to 68% (free and liberated combined). Sphalerite liberation varied between 81 to 89%. Sphalerite is mostly free with about 7 to 10% associated with pyrite.

In general, SGS indicated that the liberation of galena and chalcopyrite was adequate, and acceptable copper and lead metallurgical performance was expected within the rougher circuit. Sphalerite was well liberated at the grind size.

COMMINUTION TEST WORK

SGS conducted a comminution study on five selected samples during the test program. The tests included the standard BWi test and Ai test.

Table 13.4 shows the results of the tests. The BWi values range from 6.5 to 11 kWh/t for the materials sampled. The data indicates that the samples are not resistant to ball mill grinding. The Ai ranged from 0.017 to 0.072 g, which indicates that the samples are not abrasive.

Table 13.4 Bond Ball Mill Grindability and Abrasion Index Test Results

Sample	Mesh of Grind Size	P80 (µm)	BWi (kWh/t)	Ai (g)
MET – 1105341	150	88	6.7	0.032
MET – 1106043	150	88	6.5	0.019
MET – 1105868	150	85	7.4	0.030
MET – 1106033	150	87	9.3	0.072
MET – 1105853	150	89	11.1	0.017

FLOTATION TEST WORK

In 2012, SGS conducted bench-scale flotation test work to investigate the recovery of copper, lead, zinc, and associated precious metals using bulk copper-lead flotation and zinc flotation, followed by copper and lead separation. The four composite samples were tested for rougher flotation kinetics, cleaner efficiency, and copper and lead separation flotation efficiency. SGS also conducted locked cycle flotation tests on each composite.

The tests produced similar metallurgical performances among the samples tested, although the Zone 1 & 2 composite showed slightly inferior performance compared to the Zone 3 composite and Zone 5 composite.

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Open Circuit Flotation Test Work

The initial flotation tests evaluated rougher kinetics by investigating the effect of various reagent regimes on the flotation kinetics of copper, lead, and zinc minerals.

Cytec 3418A promoter and sodium isopropyl xanthate (SIPX) were used as collectors in the copper and lead flotation circuits. Methyl isobutyl carbinol (MIBC) was used as the frother to maintain a stable froth in the flotation stages. Hydrated lime was used as the pH regulator. Zinc cyanide, a mixture of zinc sulphate and sodium cyanide, or zinc sulphate alone, was used to suppress zinc minerals that might report to the copper and lead bulk concentrate.

Zinc was floated after the copper-lead bulk flotation using the traditional reagent regime, including SIPX as the collector and copper sulphate as the sphalerite activator at an elevated pH.

The feed material was ground to 80% passing 70 µm prior to talc pre-flotation. The talc flotation tailings were sent for copper-lead bulk flotation. The bulk copper-lead flotation tailings were conditioned with copper sulphate to activate sphalerite prior to zinc rougher flotation.

Regrinding was included in the flowsheet for both the copper-lead bulk concentrate and the zinc concentrate. The target regrind sizes were 80% passing 24 µm for the copper-lead bulk concentrate and 40 µm for the zinc concentrate.

The reground bulk copper-lead concentrate was cleaned to further reject sphalerite, pyrite, and other gangues. The reground zinc rougher concentrate was cleaned to produce the final zinc concentrate.

The testing indicated that a primary grind size of 80% passing 70 µm was adequate for the optimum copper-lead bulk rougher flotation and zinc rougher flotation. Copper grade and recovery to the bulk copper/lead rougher concentrate ranged from 16 to 21% and from 86 to 94%, respectively. The bulk concentrate also recovered between 89 and 94% lead, grading at 6.8 to 8.4%.

Gold and silver reported preferentially to the bulk copper-lead rougher concentrate. Gold recovery ranged from 54 to 80% to the bulk copper and lead cleaner concentrate, while silver recovery to the concentrate was in the range of between 68 and 84%.

Approximately 250 g/t of zinc cyanide was required to effectively depress the zinc minerals during flotation of the copper and lead minerals. Although zinc sulphate could be used as an alternative for zinc cyanide, approximately 1,500 g/t of zinc sulphate would be required, which is much higher than the zinc cyanide dosage. SGS recommended further tests to optimize the reagent regimes for zinc mineral suppression.

The cleaner flotation tests showed that regrinding was required to upgrade the bulk concentrates prior to separation of copper and lead minerals. The regrind size had not been optimized. It appeared that a regrind size of 80% passing approximately 30 µm would provide sufficient liberation for the bulk concentrate upgrading and copper-lead separation. Concentrate regrinding was incorporated into all locked cycle tests and open circuit cleaning tests.

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In the batch cleaner tests, lead was separated from the bulk copper and lead concentrate using a procedure to float lead minerals and suppress copper minerals. With one stage of lead rougher flotation and two stages of cleaner flotation, approximately 50 to 75% of the lead was recovered to the lead concentrate containing 41 to 60% lead. A high-grade copper concentrate was produced, ranging between 29 and 31% copper. The concentrate recovered between 75% and 91% of the copper from the bulk concentrates produced from the four composites.

Locked Cycle Test

SGS conducted six locked cycle tests to simulate bulk copper-lead flotation and zinc flotation in closed circuit. The bulk copper and lead concentrates produced were tested for copper and lead separation in an open circuit. The average locked cycle test results are shown in Table 13.5.

The copper recoveries to the bulk copper-lead concentrates produced from the locked cycle tests were as follows:

• 89 to 92% for the Zone 3 & 5 composite

• 93% for the Zone 3 composite

• 86 to 91% for the Zone 5 composite

• 84% for the Zone 1 & 2 composite.

The Zone 1 & 2 composite produced a low copper recovery. This could be due to insufficient sample for developing the optimized conditions for the test.

The copper grades of the copper concentrate produced ranged from 24 to 28%.

Approximately 88 to 94% of the lead was recovered to the bulk copper-lead concentrates, which contained 9 to 13% lead.

Three of the four composites demonstrated good zinc recovery in the locked cycle tests, excluding the Zone 1 & 2 composite sample.

The zinc recoveries to the final zinc concentrates produced from the locked cycle tests were as follows:

• 92% for the Zone 3 & 5 composite

• 93% for the Zone 3 composite

• 91% for the Zone 5 composite

• 84% for the Zone 1 & 2 composite.

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On average, the zinc grades of the concentrates produced were higher than 55%, excluding the concentrate generated from Composite Zone 1 & 2, which contained only 44.5% zinc.

Gold and silver were predominantly recovered into the bulk copper-lead concentrates. Gold recoveries to this concentrate ranged from 65 to 80%, and silver recoveries ranged from 80 to 86%.

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Table 13.5 Locked Cycle Metallurgical Test Results

Test No.	Product	Regrind Size 80% Passing	Weight %	Assays						Distribution (%)
				Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)	S (%)	Cu	Pb	Zn	Au	Ag	S
Zone 3 & 5 LCT-1	Talc Concentrate	Cu/Pb Rougher Concentrate: 52 µm; Zn Rougher Concentrate: 53 µm	7.3	0.66	0.35	1.25	0.09	15.7	2.56	0.4	0.5	0.5	0.3	0.5	0.3
	Cu/Pb Cleaner 2 Concentrate		14.8	27.6	10.2	1.96	4.05	405	30.0	89.4	91.7	4.6	70.3	84.0	20.0
	Zn Cleaner 2 Concentrate		10.3	3.11	0.62	57.2	0.67	60.9	32.8	7.0	3.9	92.8	8.1	8.8	15.3
	Zn Cleaner 1 Scavenger Tailings		3.9	1.24	0.42	1.46	2.85	41.7	28.9	1.1	1.0	0.9	13.1	2.3	5.1
	Zn Rougher Tailings		63.6	0.15	0.07	0.13	0.11	4.90	20.8	2.1	2.9	1.3	8.2	4.4	59.4
	Feed		100.0	4.42	1.59	6.17	0.83	69.1	21.5	100.0	100.0	100.0	100.0	100.0	100.0
Zone 3 LCT-2	Talc Concentrate	Cu/Pb Rougher Concentrate: 43 µm; Zn Rougher Concentrate: 41 µm	1.6	2.39	2.44	4.05	0.51	105.0	9.97	0.4	0.8	0.3	0.4	0.8	0.2
	Cu/Pb Cleaner 2 Concentrate		12.9	24.7	12.4	3.61	4.73	506	30.5	92.5	92.6	5.5	77.6	85.9	15.4
	Zn Cleaner 2 Concentrate		12.9	1.02	0.38	61.4	0.40	41.7	32.9	3.8	2.8	93.0	6.5	7.1	16.5
	Zn Cleaner 1 Scavenger		5.9	0.85	0.33	0.86	0.97	35.0	38.7	1.5	1.1	0.6	7.3	2.7	9.0
	Tailings
	Zn Rougher Tailings		66.7	0.10	0.07	0.09	0.10	4.01	22.5	1.9	2.7	0.7	8.3	3.5	58.9
	Feed		100.0	3.42	1.71	8.43	0.78	75.3	25.4	100.0	100.0	100.0	100.0	100.0	100.0
Zone 5 LCT-3	Talc Concentrate	Cu/Pb Rougher Concentrate: 36 µm; Zn Rougher Concentrate: 35 µm	1.3	7.15	3.71	2.46	1.22	187.0	13.7	1.2	1.2	0.2	0.3	1.4	0.3
	Cu/Pb Cleaner 2 Concentrate		9.9	23.8	12.9	5.04	11.2	499	31.5	91.3	92.0	9.1	70.9	84.2	14.7
	Zn Cleaner 2 Concentrate		8.3	0.91	0.56	59.1	0.55	46.4	30.5	2.9	3.4	89.3	2.9	6.6	11.9
	Zn Cleaner 1 Scavenger Tailings		7.1	0.80	0.28	0.56	4.55	30.0	32.4	2.2	1.4	0.7	20.5	3.6	10.7
	Zn Rougher Tailings		73.4	0.09	0.04	0.05	0.11	3.38	18.1	2.4	2.0	0.7	5.3	4.2	62.4
	Feed		100.0	2.56	1.37	5.47	1.55	58.2	21.1	100.0	100.0	100.0	100.0	100.0	100.0

table continues…

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Test No.	Product	Regrind Size 80% Passing	Weight %	Assays						Distribution (%)
				Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)	S (%)	Cu	Pb	Zn	Au	Ag	S
Zone 3 & 5 LCT 4	Talc Concentrate	Cu/Pb Rougher Concentrate: 45 µm; Zn Rougher Concentrate: 23 µm	7.3	0.72	0.38	1.37	0.11	17.6	3.01	0.4	0.6	0.4	0.3	0.6	0.3
	Cu/Pb Cleaner 2 Concentrate		16.0	25.3	9.25	3.13	4.28	408	29.4	91.7	92.3	6.4	73.8	85.0	21.3
	Zn Cleaner 2 Concentrate		11.8	1.78	0.39	60.9	0.48	50.7	32.5	4.8	2.9	91.6	6.1	7.8	17.4
	Zn Cleaner 1 Scavenger Tailings		4.8	1.15	0.38	1.09	2.6	39.8	27.1	1.3	1.1	0.7	13.5	2.5	5.9
	Zn Rougher Tailings		60.2	0.14	0.08	0.13	0.1	5.19	20.2	1.9	3.2	1	6.3	4.1	55.1
	Feed		100.0	4.41	1.6	7.85	0.93	76.6	22.1	100	100	100	100	100	100
Zone 5 LCT-5	Talc Concentrate	Cu/Pb Rougher Concentrate: 32 µm; Zn Rougher Concentrate: 24 µm	1.1	8.29	3.33	2.21	1.31	229	12.5	1.2	0.9	0.1	0.4	1.4	0.2
	Cu/Pb Cleaner 2 Concentrate		8.9	24.3	13.2	4.09	8.93	507	29.7	85.7	88.3	6.4	62.5	76.3	12.5
	Zn Cleaner 2 Concentrate		9.6	2.01	0.83	54.7	0.64	75.5	32.9	7.6	5.9	91.8	4.8	12.2	14.8
	Zn Cleaner 1 Scavenger Tailings		11.4	0.55	0.23	0.31	2.76	22.9	38.8	2.5	2.0	0.6	24.8	4.4	20.9
	Zn Rougher Tailings		69	0.11	0.06	0.09	0.14	4.97	15.8	3.1	3.0	1.0	7.5	5.8	51.5
	Feed		100.0	2.54	1.34	5.69	1.28	59.4	21.2	100	100	100	100	100	100
Zone 1 & 2 LCT-6	Talc Concentrate	Cu/Pb Rougher Concentrate: 62 µm; Zn Rougher Concentrate: 55 µm	4.8	0.67	0.34	0.90	0.40	13.9	1.88	0.4	0.6	0.4	0.8	0.4	0.3
	Cu/Pb Cleaner 2 Concentrate		9.5	23.7	9.54	5.12	6.65	481	30.2	84.2	94.0	14.3	79.7	84.2	32.5
	Zn Cleaner 2 Concentrate		6.4	5.84	0.49	44.5	0.91	101.5	32.8	14.0	3.2	83.7	7.4	12.0	23.9
	Zn Cleaner 1 Scavenger Tailings		7.4	0.22	0.06	0.17	0.91	12.3	19.6	0.6	0.5	0.4	8.4	1.7	16.4
	Zn Rougher Tailings		71.8	0.03	0.02	0.06	0.04	1.34	3.30	0.8	1.7	1.2	3.7	1.8	26.8
	Feed		100.0	2.69	0.97	3.42	0.80	54.6	8.8	100.0	100.0	100.0	100.0	100.0	100.0

Note: LCT = locked cycle test

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Copper/Lead Separation

SGS performed copper and lead separation tests on the bulk copper-lead concentrates produced from the locked cycle tests. Sodium cyanide was used to suppress copper minerals; 3418A was used as the lead collector and lime added to adjust the pulp pH to 10. Table 13.6 summarizes the separation test results.

The copper concentrates that were produced assayed at:

• 31% copper from Composite Zone 3 & 5

• 31% copper from Composite Zone 3

• 30% copper from Composite Zone 5

• 28 to 29% copper from Composite Zone 1 & 2.

The lead second cleaner concentrates that were produced contained:

• 41% lead from Composite Zone 3 & 5

• 59% lead from Composite Zone 3

• 67% lead from Composite Zone 5

• 55% lead from Composite Zone 1 & 2.

On average, the lead concentrates that were produced from the Zone 1 & 2, Zone 3, and Zone 5 composites contained approximately 2.2% copper while the copper content of the concentrate from the Zone 3 & 5 composite was higher, grading at 5%. There is a substantial reduction in lead recovery when the lead first cleaner concentrate was further upgraded.

Gold was predominantly recovered into the copper concentrate. However, silver recovered to the lead second cleaner concentrates varied from sample to sample. It appears that the silver in the bulk copper-lead concentrate that was produced from Composite Zone 3 is more likely to report to the lead concentrate. However, the copper concentrate recovered more silver than lead concentrate for the Zone 1 & 2 and Zone 5 composites.

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Table 13.6 Copper and Lead Separation Test Results

Test	Product	Weight %	Assays						Distribution (%)
			Cu (%)	Pb (%)	Zn (%)	Ag (g/t)	Au (g/t)	S (%)
									Cu	Pb	Zn	Ag	Au	S
Zone 3 & 5 Cu/Pb Separation Feed from LCT-4 (Cycle 2)	Pb 2ndCleaner Concentrate	8.2	5.99	41.0	2.02	2,330	18.9	13.1	1.9	37.0	6.0	44.7	35.9	3.4
	Pb 1stCleaner Concentrate	22	6.87	37.5	4.34	1,665	13.6	20.6	5.9	90.8	34.8	85.7	69.5	14.3
	Pb Rougher Concentrate	37.7	16.4	23.0	3.43	1,033	9.17	26.2	24.1	95.5	47.4	91.3	80.3	31.4
	Pb Rougher Tailings (Cu Concentrate)	62.3	31.3	0.65	2.31	59	1.36	34.7	75.9	4.5	52.6	8.7	19.7	68.6
	Cu/Pb 2ndCleaner Concentrate (Head)	-	25.7	9.07	2.73	4.27	4.31	31.5	-	-	-	-	-	-
Zone 3 & 5 Cu/Pb Separation Feed from LCT-4 (Cycle 3)	Pb 2ndCleaner Concentrate	10.8	4.09	41.2	2.75	1,970	1.87	12.3	1.7	49.7	8.9	53.4	4.9	4.2
	Pb 1stCleaner Concentrate	20.3	5.47	38.1	4.76	1,618	1.38	19.3	4.3	86.6	28.9	82.6	6.8	12.5
	Pb Rougher Concentrate	28.9	11.6	28.4	4.16	1,206	1.19	23.4	13	92.0	36.0	87.6	8.4	21.6
	Pb Rougher Tailings (Cu Concentrate)	71.1	31.6	1.0	3.01	69	5.29	34.7	87	8.0	64.0	12.4	91.6	78.4
	Cu/Pb 2ndCleaner Concentrate(Head)	-	25.8	8.93	3.34	398	4.11	31.4	100.0	100.0	100.0	100.0	100.0	100.0
Zone 3 Cu/Pb Separation from Open Circuit Test (Test F25)	Pb 2ndCleaner Concentrate	2.1	2.22	58.8	5.58	1,622	0.3	20.8	1.4	74.9	1.4	44.2	1.0	1.8
	Pb 1stCleaner Concentrate	2.9	4.51	48.3	6.94	1,369	0.5	24.1	3.8	83.8	2.4	50.9	2.0	2.8
	Pb Rougher Concentrate	4.3	12.4	33.6	6.54	1,026	1.05	26.9	15.3	86.0	3.3	56.3	6.6	4.6
	Pb Rougher Tailings (Cu Concentrate)	8.3	31.5	0.29	4.33	231	5.24	33.3	75.1	1.4	4.2	24.5	63.9	11.0
	Cu/Pb 2ndCleaner Concentrate (Head)	12.6	25.0	11.6	5.08	502	3.81	31.1	90.4	87.4	7.5	80.8	70.5	15.5
Zone 5 Cu/Pb Separation Feed from LCT-5 (Cycle 2)	Pb 2ndCleaner Concentrate	6.6	2.42	69.0	2.68	1,230	1.27	15.8	0.6	41.1	3	17.2	1.8	3.3
	Pb 1stCleaner Concentrate	15.2	3.78	57.6	4.18	993	1.92	20.5	2.3	78.8	11.5	31.9	6.1	9.8
	Pb Rougher Concentrate	25.5	10.3	40.3	4.82	778	6.31	25.1	10.5	92.4	22.1	41.9	33.6	20.1
	Pb Rougher Tailings (Cu Concentrate)	74.5	30.0	1.13	5.79	369	4.26	34.1	89.5	7.58	77.9	58.1	66.4	79.9
	Cu/Pb 2ndCleaner Concentrate (Head)	-	25.0	11.1	5.54	473	4.78	31.8	100.0	100.0	100.0	100.0	100.0	100.0
Zone 5 Cu/Pb Separation Feed from LCT 5 (Cycle 3)	Pb 2ndCleaner Concentrate	5.2	2.09	65.4	3.72	1,180	1.98	17.8	0.4	28.0	4.7	12.5	1.6	2.9
	Pb 1stCleaner Concentrate	17.5	3.54	54	4.09	900	1.24	21.9	2.5	77.9	17.5	32.2	3.4	12.1
	Pb Rougher Concentrate	27.3	8.5	40	4.27	760	7.83	25.9	9.5	90	28.5	42.4	33	22.2
	Pb Rougher Tailings (Cu Concentrate)	72.7	30.4	1.67	4.01	388	5.97	34	90.5	10	71.5	57.6	67	77.8
	Cu/Pb 2ndCleaner Concentrate (Head)	-	24.4	12.1	4.08	489	6.48	31.8	100.0	100.0	100.0	100.0	100.0	100.0

table continues…

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Test	Product	Weight %	Assays						Distribution (%)
			Cu (%)	Pb (%)	Zn (%)	Ag (g/t)	Au (g/t)	S (%)	Cu	Pb	Zn	Ag	Au	S
Zone 1 & 2 Cu/Pb Separation Feed from LCT-6 (Cycle 2)	Pb 2ndCleaner Concentrate	7.59	2.4	57.3	5.59	0.54	1,313	15.1	0.76	47.1	8.1	0.7	20.1	3.78
	Pb 1stCleaner Concentrate	16.4	4.38	45.3	7.96	0.77	1,038	19.9	2.98	80.5	24.9	2.2	34.4	10.8
	Pb Rougher Concentrate	23.6	9.6	34.3	7.19	1.13	849	22.9	9.4	87.7	32.3	4.6	40.4	17.8
	Pb Rougher Tailings (Cu Concentrate)	76.4	28.6	1.49	4.64	7.14	386	32.6	90.6	12.34	67.7	95.4	59.6	82.2
	Cu/Pb 2ndCleaner Concentrate (Head)	-	24.1	9.23	5.24	5.72	495	30.3	100.0	100.0	100.0	100.0	100.0	100.0
Zone 1 & 2 Cu/Pb Separation Feed from LCT-6 (Cycle 3)	Pb 2ndCleaner Concentrate	4.74	1.8	53.2	3.86	0.77	1,373	11.8	0.36	28.4	4.07	0.7	14.1	1.87
	Pb 1stCleaner Concentrate	13.2	3.31	48.3	6.37	0.74	1,155	16.6	1.84	72.2	18.8	1.8	33.2	7.3
	Pb Rougher Concentrate	22	8.7	34.6	6.24	1.13	874	20.9	7.99	85.7	30.5	4.5	41.7	15.3
	Pb Rougher Tailings (Cu Concentrate)	78	28.1	1.62	4.01	6.72	344	32.4	92	14.3	69.5	95.5	58.3	84.7
	Cu/Pb 2ndCleaner Concentrate (Head)	-	23.8	8.9	4.5	5.49	461	29.9	100.0	100.0	100.0	100.0	100.0	100.0

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Concentrate Multi-element Assay

ICP assays were conducted on the copper and lead concentrates produced from the open circuit tests and the zinc concentrate from the locked cycle tests. The main impurity elements are shown in Table 13.7 to Table 13.9.

The results indicated that on average, the arsenic, antimony, and lead content of the copper concentrates may be higher than the penalty thresholds outlined by most copper concentrate smelters. The mercury content of the concentrate should be lower than the penalty threshold excluding the concentrate produced from Composite Zone 5, which may exceed the threshold.

Similar to the copper concentrates, the lead concentrate may receive penalties for the high arsenic and antimony concentrations.

Silicon dioxide and fluoride assays should be conducted on the concentrates to determine whether or not they are higher than the penalty thresholds.

Table 13.7 Multi-element Assay Results – Copper Concentrate

Composite	Zone 3 & 5	Zone 5	Zone 1 & 2
Mercury (ppm)	1.9	9.8	4.2
Arsenic (ppm)	1,114	4,040	82
Antimony (ppm)	123	>1,000	1,370
Cadmium	130	259	224
Lead	0.8	1.4	1.6
Zinc	2.7	4.9	4.3

Table 13.8 Multi-element Assay Results – Lead Concentrate

Composite	Zone 3 & 5	Zone 5	Zone 1 & 2
Mercury (ppm)	4.5	4.1	2.7
Arsenic (ppm)	11,640	3,165	114
Antimony (ppm)	>1,000	>1,000	1,376
Cadmium	169	195	249
Copper	5.0	2.3	2.1
Zinc	2.4	3.2	4.7

In the zinc concentrates produced from the locked cycle tests, the cadmium content generally ranges from 2,100 to 3,400 ppm, which will likely be higher than the penalty thresholds outlined by most zinc concentrate smelters. The arsenic content may be higher than the penalty mark in the concentrate produced from Composite Zone 5. However, the mineralization from Zone 5 is not expected to be mined separately, on average; therefore, the arsenic in the zinc concentrate should not attract a penalty.

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Table 13.9 Multi-element Assay Results – Zinc Concentrate

Test	LCT 1	LCT 2	LCT 3	LCT 4	LCT 5	LCT 6
Composite	Zone 3 & 5	Zone 3	Zone 5	Zone 3 & 5	Zone 5	Zone 1 & 2
Mercury (ppm)	No Data	No Data	No Data	No Data	No Data	No Data
Arsenic (ppm)	688	89	1310	706	1,020	754
Antimony (ppm)	436	184	418	211	584	550
Cadmium	3,010	3,390	3,290	3,440	2,910	2,110
Copper	3.1	1.0	0.9	1.8	2.0	5.8
Lead	0.6	0.4	0.6	0.4	0.8	0.5

13.1.3 HISTORICAL TEST WORK REVIEW

METALLURGICAL TESTING (1968 TO 1976)

Mineralogy

Between 1970 and 1976, KRC conducted two initial mineralogical studies to evaluate and identify the potential beneficiation or metallurgical treatment of concentrates of the samples from the deposit.

Kennecott Research Center – 1970

In the 1970 mineralogy investigation, KRC reported that the host rock of the mineralization is generally muscovite, chlorite, or talc schist. Principal economic minerals in the deposit were identified as chalcopyrite, sphalerite, and argentiferous galena. Table 13.10 presents a complete list of metallic minerals identified in the Arctic Project samples.

Table 13.10Metallic Mineral Identified in Arctic Project Samples

	Mineral		Mineral
Mineral	Abundance	Mineral	Abundance
Chalcopyrite	Very Abundant	Tennantite	Minor
Sphalerite	Very Abundant	Digenite	Minor
Galena	Common	Bornite	Minor
Pyrite	Common	Covellite	Trace
Sphene	Common	Carrollite	Trace
Rutile	Common	Glaucodot	Trace
Pyrrhotite	Minor	Stromeyerite	Trace
Marcasite	Minor	Electrum	Trace
Arsenopyrite	Minor	Unidentified	Trace

Source: KRC 1976

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The sizes of sulphide mineral particles in the mineralization sample ranged from submicron to a maximum of several centimetres; most of the sulfide particles were relatively large (coarser than 74 µm). KRC noted that the target sulphide minerals should be liberated from gangue at a primary grind size of 100% passing 100 mesh.

It should be possible to obtain a zinc concentrate that is low in iron and contains most of the cadmium that occurs in the mineralization. There was a close association between chalcopyrite and sphalerite, including some chalcopyrite exsolution particles within the sphalerite grains. Because of this association, some copper was expected to report to the zinc concentrate.

The copper-lead concentrate would contain most of the silver, gold, nickel, and cobalt that is recovered from the mineralization. A major portion of the silver in the mineralization occurs in galena. In addition, some silver minerals were physically attached to galena particles. Because of these associations, the silver will tend to go with the lead in any further concentration of lead from the copper-lead concentrate. Nickel and cobalt recovered in the flotation concentrates were expected to follow the copper minerals.

Kennecott Research Center – 1975

The objective of the 1975 test program was to identify potential problems that might influence beneficiation of the mineralization.

A detailed mineralogical examination was conducted on 88 drill core samples. The mineralogical observations are summarized as follows:

• Large variations in mineralogy occur both vertically and laterally within the deposit.

• A significant portion of the chalcopyrite is severely interlocked with either sphalerite or galena.

• Pyrite contains abundant base metal sulphide inclusions.

• Silver is present in galena and in tetrahedrite.

• Arsenic and antimony can be expected in the concentrates due to the presence of arsenopyrite and tetrahedrite/tennantite.

• Trace quantities of nickel and bismuth sulphides were observed.

The important sulphide minerals are pyrite, sphalerite, chalcopyrite, galena, pyrrhotite and arsenopyrite.

The following potential problems were identified:

• It may be difficult to liberate chalcopyrite from sphalerite.

• Abundant base metal sulphide inclusions in pyrite may make it difficult to reject this mineral by flotation.

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• It may be difficult to liberate galena from chalcopyrite.

• Silver values are present in both tetrahedrite and galena.

• Flotation of arsenopyrite and tetrahedrite-tennantite may cause elevated arsenic and antimony in the concentrates.

• Trace quantities of nickel and bismuth minerals were observed in the mineralization.

Comminution Test Work

In 1976, KRC conducted preliminary comminution test work using the standard Bond Work Index determination procedure.

Table 13.11 shows the results of the BWi tests. Mineralization from the Arctic Deposit is relatively soft, with a Bond Work Index in the range of 5.7 to 12.0 kWh/t.

Table 13.11 Bond Ball Mill Work Index

Hole No.	Work Index (kWh/t)	Talc (%)
11B	11.96	90
34B	8.33	50
34B	5.71	5
34B	11.3	Mainly Talc
34C	9.98	Nil
48A	10.5	Mainly Sulphide
48B	9.60	20

Source: KRC 1976

Various observations from the grindability tests conducted during the KRC 1976 test program are summarized as follows:

• Wet bulk material is expected to be quite sticky and would require special consideration with regard to screen blinding and clogging conveyor belts and chutes.

• Arctic samples that contain talc may cause some difficulty during grinding because the talc may flatten into flakes rather than breaking, which may cause grinding and classification problems.

• The sample that contained mainly talc did not respond in the normal manner to the standard Bond Work Index laboratory determination.

Flotation Test Work

Between 1968 and 1976, KRC carried out initial amenability testing. The focus was on selective flotation to provide separate copper, lead, and zinc concentrates for conventional smelting. In 1968, initial amenability testing was conducted on core composites from eight diamond drill holes (which is not available to review). Other tests were conducted in 1972 on four composites from three additional diamond core holes. The laboratory-scale tests conducted between 1968 and 1976 included the conventional selective flotation approach to produce separate lead, copper and zinc concentrates.

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The major problem encountered for the tests by KRC was the separation between lead and copper minerals, and the reduction of zinc deportment to the copper and lead concentrates. The copper concentrates produced from open circuit tests contained 30 to 32.4% copper, 0.45 to 3.48% zinc and 0.15% to 1.31% lead. The copper recoveries were less than 80.7%. The lead concentrate grades were low, ranging from 17.1 to 36.5%.

Sphalerite flotation was generally efficient, producing zinc flotation concentrates grading approximately 55% zinc. Because of the low gold content of the test samples, no appraisal was made of gold recoveries.

From 1975 and 1976, large diameter cores from 14 drill holes were used for more detailed testing. Two composites labelled as Composite No. 1 (Eastern Zone) and Composite No. 2 (Western Zone), were prepared. The test program included bench-scale testing of various process parameters for sequential flotation, including locked cycle tests. A talc flotation step prior to sulfide flotation was considered to be necessary, as previously established. It was determined that chalcopyrite and sphalerite could be recovered into separate commercial grade copper and zinc concentrates. However, the production of a selective high grade lead concentrate was not successful.

Using zinc sulphate and sodium bisulphate to suppress galena and sphalerate, 90% of the copper was recovered into a concentrate containing 26% copper, 1.5% lead, and 6% zinc. KRC indicated that because of close interlocking of chalcopyrite and sphalerite, the zinc content of the copper concentrate could not be reduced to below 6% without sacrifice of copper recovery.

Only low-grade silver-bearing lead concentrates were obtained. Under the best test conditions, approximately 65% of the silver reported to the low-grade lead concentrate. Some of the silver in the mineralization occurred as tetrahedrite, which was recovered to the copper concentrate.

It appeared that zinc minerals responded well to the test procedure.

METALLURGICAL TEST WORK (1998 TO 1999)

In 1999, Lakefield conducted a metallurgical test program to confirm and improve upon the results from the 1970s KRC test work program. The Lakefield work was carried out on test composites from the Arctic Deposit prepared from three separate drill holes. The test composite from the upper portion of AR-72 was identified as being low in talc content; however, composites from the lower portion of AR-72 were high in talc content, as were AR-74 and AR-75. The head analyses for the respective resulting test composites are summarized in Table 13.12.

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Table 13.12 Head Analyses

Composite	Talc	Cu (%)	Zn (%)	Pb (%)	Fe (%)	ST (%)	Au (g/t)	Ag (g/t)
Hole #72 – Upper	Low	5.28	7.16	1.86	15.6	1.14	72.3	23.4
Hole #72 – Lower	High	2.68	5.85	1.34	13.0	1.60	75.9	16.9
Hole #74	High	2.46	4.43	0.90	17.0	1.55	45.1	23.7
Hole #75	High	2.35	8.36	1.95	15.7	1.23	77.3	21.8

Note:	ST= total sulphur
Source:	Lakefield 1999

Low Talc Composite Flotation

Lakefield conducted a series of five tests on the low talc mineralized composite. The following parameters were used for all tests:

• MIBC was used in the talc pre-float.

• Sulphur dioxide was used in the copper-lead flotation circuit.

• A grind size of approximately 80% passing 53 µm was used.

• Bulk copper-lead flotation was included, followed by zinc flotation.

The bulk copper-lead rougher concentrate was reground and subjected to two stages of cleaner flotation and one stage of copper and lead separation, using zinc oxide and sodium cyanide to depress the copper while floating the lead. The resulting lead rougher concentrate was upgraded with two stages of cleaner flotation to produce the final lead concentrate. The lead rougher flotation tailings were the final copper concentrate.

The zinc rougher concentrate was reground and upgraded with two stages of cleaner flotation. The results of the best open circuit flotation test for the low talc composite are summarized in Table 13.13. The test results showed that:

• Copper concentrate produced contained 29% copper. 86.8% of the copper was recovered to the concentrate.

• The lead concentrate recovered 68% of the lead.

The zinc concentrate that was produced from the open circuit test contained 59.1% zinc.

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Table 13.13 Flotation Test on Ambler Low Talc Composite

Item	Weight (%)	Assays					Distribution (%)
		Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)	Cu	Pb	Zn	Au	Ag
Lead Concentrate	2.22	6.5	58.8	3.43	38.9	1,703	2.7	68.1	1.1	48.7	47.3
Copper Concentrate*	15.76	29.1	1.2	2.61	1.23	73.5	86.8	9.8	5.7	10.9	14.5
Zinc Concentrate	9.91	0.44	0.36	59.1	0.65	14.7	0.8	1.9	81.1	3.6	1.8
Zinc Tailings**	61.6	0.11	0.13	0.22	0.4	3.47	1.2	4.3	1.9	13.7	2.7
Head (Calculation)	100.0	5.28	1.92	7.21	1.78	80.1	100.0	100.0	100.0	100.0	100.0

Source:	Lakefield 1999.
Notes:	*Pb Rougher Tailings
	**Does not include intermediate cleaner tailings

High Talc Composite Flotation

Lakefield also conducted flotation tests on each of the high talc composites using a test procedure similar to the one used for the low talc composite, with the exception that carboxymethyl cellulose (CMC) was added as a depressant for talc. The results of these tests showed that the presence of talc had a significant negative impact on the copper and lead mineral recoveries. Lakefield also used talc pre-flotation prior to sulphide flotation in an effort to reduce talc effect on base metal flotation. It appears that the talc pre-flotation improved copper and lead metallurgical performances. However, the test results showed that an elevated talc content had a significant effect in copper and lead flotation response.

In the test report, Lakefield also concluded that:

• A grind particle size as coarse as approximately 80% passing 74 µm provided good results.

• Copper-lead separation was difficult using a cyanide compound with the talc mineralization due to the talc and perhaps soluble copper as well.

13.2 RECOMMENDED TEST WORK

In general, the flowsheet developed in the 2012 test program is feasible for the Arctic mineralization. Tetra Tech recommends further metallurgical test work on representative samples to optimize flowsheet and process conditions and determine the design related parameters required for further studies. The recommended test work is shown in Section 26.3.

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14.0 MINERALRESOURCEESTIMATES

14.1 INTRODUCTION

The Mineral Resource Statement presented herein represents the latest in a series of mineral resource evaluations filed by Kennecott, NovaGold and NovaCopper for the Arctic Project, starting from 1990 (Randolf 1990), 2008 (SRK 2008), 2011 (SRK 2011) and most recently 2012 (SRK 2012). The mineral resource estimate which formed the basis of this PEA was completed by Mike O’Brien, Pr.SciNat (SACNASP#400295/87); the overall effective date of this resource estimate is July 30, 2013.

The mineral resource model prepared by Tetra Tech considers diamond drill holes drilled by different operators during the period 1965 to 2011. The majority of the assaying has been completed in recent years by NovaCopper and its previous parent company NovaGold. The mineral resource for the Arctic Project is supported by 43 core holes (approximately 13,500 m) drilled by NovaGold and 92 core holes (approximately 17,600 m) drilled by previous owners Kennecott, and/or a Kennecott subsidiary.

The geological and assay database used to estimate the Arctic Project mineral resources have been reviewed and audited by Tetra Tech. It is Tetra Tech’s opinion that the current drilling information is sufficiently reliable to interpret with confidence the boundaries for VMS mineralization and that the assay data are sufficiently reliable to support mineral resource estimation.

Leapfrog^™ software (version 2.5.1) was used to review and verify the resource estimation domains, prior to being imported into Isatis^™ software (version 2012.1) to prepare assay data for geostatistical analysis, variography, block model construction, metal grade estimation and mineral resource tabulation. Mineral Resources were estimated into five MS and six SMS lenses, and then combined with waste for an overall grade for the 10 m by 10 m by 5 m block. Extreme lead and gold assays were capped prior to compositing. OK and ID² estimates were run, with OK used for resource reporting and ID² used for validation. Search parameters were constrained within each mineralized domain and required an optimum number of 15 composites, minimum number of 5 composites, minimum number of 2 drill holes, and maximum search distance range of 200 m. In general, blocks categorized as Indicated were supported by at least two drill holes within a 75 m search radii, and blocks categorized as Inferred were supported by at least 2 drill holes within a 150 m search radii.

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This section describes the resource estimation methodology and summarizes the key assumptions considered by Tetra Tech. In Tetra Tech’s opinion, the resource evaluation reported herein is a sound representation of the copper/zinc/lead/gold/silver mineral resources found on the Arctic Project at the current level of sampling. The mineral resources have been estimated in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines.

14.2 PREVIOUS RESOURCE ESTIMATES

Differences between the previously reported mineral resource estimate (SRK 2012) and the current resource estimate are primarily related to additional drilling, updated geological interpretation, additional specific gravity determinations, and reporting of grades and tonnes within an open pit designed to support the requirements for reasonable prospects for economic extraction.

Previous resource estimates for the Arctic Deposit, which includes work completed by Randolf (1990), and SRK (2008, 2011 and 2012), were intended to support an underground mining concept. Consequently the resource estimates focussed on the high grade portions of the zones, and did not consider the potential amenability of the mineral resources to open pit mining methods.

14.2.1 KENNECOTT (1990) RESOURCE ESTIMATE

The historic resource estimates are considered relevant but not necessarily reliable. In each case, a significant amount of resampling and verification is required to upgrade or verify the historical estimate as current mineral resources or mineral reserves. Please note that a QP has not done sufficient work to classify any of the historical estimates as current mineral resources or mineral reserves and the issuer is not treating the historical estimates as current mineral resources or mineral reserves.

In 1990, Kennecott completed a resource estimate for the Arctic Deposit based on 70 core holes. This resource estimate is summarized in Table 14.1 and considered to be an Inferred Resource. This historical resource estimate pre-dates the development of NI 43-101 reporting guidelines and was not estimated in compliance with NI 43-101 procedures and should not be relied upon (Randolf 1990).

Table 14.1 Historical Resource Estimate

Classification	Tonnes (kt)	Cu (%)	Zn (%)	Pb (%)	Ag (ppm)	Au (ppm)
Inferred	36,300	4.0	5.5	0.8	54.9	0.7

Source: Randolf (1990)

14.2.2 SRK (2008) RESOURCE ESTIMATION

In February 2008, NovaGold filed an NI 43-101 compliant resource authored by SRK for the Arctic Deposit. The estimation utilized 119 core holes, of which, 4,808 intervals were sampled representing 9,128 m of sampled drilling. Sample lengths varied from 0.1 to 12 m, and averaged about 1.9 m. Each interval contained assays for copper, zinc, lead, gold and silver, as well as codes for lithology and mineralized zone. Assays were capped based on the drill hole assay statistics presented in Table 14.2.

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Table 14.2 Drill Hole Assay Statistics – 2008 SRK Resource Estimation

Category	Length (m)	Cu (%)		Zn (%)		Pb (%)		Au (ppm)		Ag (ppm)
		Uncut	Cap 15	Uncut	Cap 18	Uncut	Cap 4	Uncut	Cap 7	Uncut	Cap 190
Host	19,703.45	0.07	0.07	0.06	0.06	0.01	0.01	0.01	0.01	1.09	1.09
Sulphide	1,077.11	3.77	3.76	6.05	6.04	0.94	0.93	0.83	0.75	56.86	55.81
Total	20,780.56	0.26	0.26	0.37	0.37	0.06	0.06	0.06	0.05	3.98	3.93
Number Capped			4		8		9		3		5

Downhole surveys were recorded at 15 to 45 m intervals for the majority of the drill holes. A standard “typical deviation” was applied to 40 holes, which were unsurveyed.

Due to the large discrepancy between previous Kennecott SG measurements, a more extensive field SG program was implemented by NovaGold in 2004, the results of which are tabulated along with the 1998 Kennecott data in Table 14.3.

Table 14.3 Historical SG Data Statistics – Arctic Deposit: 1998–2004

Program	MS >50% Sulphide (Average g/cm3)	No. of Samples	SMS <50% Sulphide (Average g/cm3)	No. of Samples	Lithologies (Average g/cm3)	No. of Samples
1998 Lab (Chemex, Golder)	4.37	15	4.02	7	2.84	16
2004 Field	4.40	35	3.84	19	2.83	73
2004 Lab (Chemex)	4.06	121	3.36	77	2.85	66
All Programs	4.16	171	3.49	103	2.84	155

Note: MS = massive sulphide; SMS = semi-massive sulphide

For the purpose of the 2008 resource estimate, non-rejected SG measurements were categorized by rock type and vary from 2.62 to 4.87 with an average of 4.2 for MS and SMS. Actual values within each zone were used to interpolate SG into the block model using ID² but, where SG sample density was too sparse, a default value of 4.2 was used in the mineralized zones. A default of 2.9 was used for all host rock, which was the average SG of non-rejected quartz mica schist samples. SG measurements are shown by rock type in Table 14.4.

Table 14.4 SG Measurements Categorized by Rock Type

Rock Category	Count	Average	Max	Min
MS+SMS	77	4.2	4.87	2.84
Non-MS/SMS	93	2.9	4.26	2.62

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Five zones representing MS lenses were modelled and composites were created at 1 m down hole intervals, broken at changes in zone codes. General directional variograms were generated for each element and, due to the drill spacing and orientation, the best variograms are in the orientation of azimuth 150°, plunge 30°, with ranges of 40 to 50 m in all elements except gold, which had a range of 25 m.

Grades were estimated using OK and a spherical search restricted within the zones. Multiple passes were used (50, 100, and 150 m) to fill as many blocks as possible within the zones. The first search pass used a minimum of two samples, with no more than three from any one drill hole; the second two passes did not have a number per drill hole restriction.

After the metal grades were estimated, a simplified Gross Metal Value (GMV) was calculated based on metal prices applied to each individual grade. The GMV was equal to the sum of each grade multiplied by the value of the metal unit. A $100 GMV cut-off was selected for the resource cut-off grade based on the assumption of underground mining methods.

Resources in the mineral zones, which were estimated by the first (50 m) search, were classified as Indicated resources, roughly based on a distance that was twice the variogram range and within one cross section distance inside a modelled shape. All blocks outside of the mineral zones, and all other estimated blocks too distant from the samples for the first pass, were classified as Inferred resources. The resource statement is shown in Table 14.5 at $100 GMV cut-off.

Table 14.5 Arctic Deposit Resources at $100 GMV Cut-off – 2008 SRK

Resource		Grade					Contained Metal
Zone	kt	Cu (%)	Zn (%)	Pb (%)	Ag (g/t)	Au (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Ag (koz)	Au (koz)
Indicated
Zone 1	5,294	4.56	6.45	1.05	62.8	0.956	533	752	122	10,684	163
Zone 2	2,982	4.36	5.82	0.80	45.8	0.521	287	383	53	4,387	50
Zone 3	1,957	3.66	6.00	0.93	51.2	0.522	158	259	40	3,220	33
Zone 4	6,092	3.82	6.00	0.98	68.7	1.008	513	805	131	13,451	197
Zone 11	517	4.16	3.32	0.34	32.9	0.254	47	38	4	546	4
Total Indicated	16,841	4.14	6.03	0.94	59.6	0.826	1,538	2,237	350	32,289	447
Inferred
Zone 0	1,162	2.21	2.27	0.69	4.2	0.333	57	58	18	156	12
Zone 1	3,163	3.92	5.75	0.93	55.0	0.760	273	401	65	5,596	77
Zone 2	1,559	4.06	5.60	0.74	43.4	0.433	139	193	25	2,176	22
Zone 3	1,307	3.83	5.13	0.63	48.1	0.438	110	148	18	2,021	18
Zone 4	4,382	3.34	5.03	0.84	58.4	0.891	323	486	81	8,224	126
Zone 11	370	4.27	3.32	0.36	33.8	0.293	35	27	3	402	3
Total Inferred	11,944	3.56	4.99	0.80	48.4	0.674	937	1,313	210	18,575	259

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Notes:	1.	g/t=ppm.
	2.	The effective date of this mineral resource estimate is January 31, 2008.
	3.	Inferred resources have a great amount of uncertainty as to their existence and as to whether they can be mined legally or economically. It cannot be assumed that all or any part of Inferred resources will ever be upgraded to a higher category.
	4.	Mineral resources that are not mineral reserves do not have demonstrated economic viability.
	5.	This mineral resource estimate assumes metal prices of $2.25/lb copper, $525/oz gold, $9.50/oz silver, $0.55/lb lead and $1.05/lb zinc.

14.2.3 SRK (2011) RESOURCE ESTIMATION UPDATE

In 2011, as part of a PEA of the Arctic Project as a potential underground mining operation, SRK re-stated the 2008 resource for NovaGold using slightly different classification criteria based on a NSR calculation utilizing updated operating costs and then current metal pricing. Though the block model estimation did not change, reported resources were updated. Table 14.6 shows the mineral resource estimate as reported May 9, 2011.

Table 14.6 Arctic Deposit Resources at $75 NSR Cut-off – 2011 SRK

Resource		Grade					Contained Metal
Zone	kt	Cu (%)	Zn (%)	Pb (%)	Ag (g/t)	Au (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Ag (koz)	Au (koz)
Indicated
Zone 1	5,293	4.56	6.45	1.05	62.77	0.96	533	752	122	10,683	163
Zone 2	2,982	4.36	5.82	0.80	45.76	0.52	287	384	53	4,387	50
Zone 3	1,964	3.66	5.98	0.93	51.02	0.52	158	259	40	3,222	33
Zone 4	6,089	3.82	6.00	0.98	68.71	1.01	513	805	131	13,451	197
Zone 11	517	4.16	3.32	0.34	32.86	0.25	47	38	4	546	4
Total Indicated	16,845	4.14	6.02	0.94	59.62	0.83	1,538	2,237	350	32,289	447
Inferred
Zone 0	1,191	2.18	2.24	0.70	4.17	0.34	57	59	18	159	13
Zone 1	3,166	3.91	5.74	0.93	54.98	0.76	273	401	65	5,596	77
Zone 2	1,559	4.06	5.60	0.74	43.40	0.43	139	193	25	2,175	22
Zone 3	1,307	3.83	5.13	0.63	48.08	0.44	110	148	18	2,020	18
Zone 4	4,492	3.28	4.95	0.83	57.56	0.87	325	490	82	8,312	126
Zone 11	373	4.25	3.30	0.35	33.65	0.29	35	27	3	404	3
Total Inferred	12,087	3.56	4.94	0.79	48.04	0.67	940	1,317	212	18,667	260

Notes:	1.	g/t=ppm.
	2.	Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resources will be converted to Mineral Reserves.
	3.	Resources stated as contained within potentially economically minable underground shapes above a US$75.00/t NSR cut-off.
	4.	NSR calculation is based on assumed metal prices of $2.50/lb for copper, $1,000/oz for gold, $16.00/oz for silver, $1.00/lb for zinc and $1.00/lb for lead. A mining cost of $45.00/t and combined processing and G&A costs of $31.00 were assumed to form the basis for the resource NSR cut-off determination.
	5.	Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbers may not add due to rounding.

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14.2.4 SRK (2012) RESOURCE ESTIMATION UPDATE

In 2012, as a result of the spinoff of NovaCopper from NovaGold, NovaCopper commissioned SRK to further update the 2011 Arctic Project PEA.

The 2012 mineral resource was developed from a drill hole database consisting of 131 core holes, 108 of which intercepted significant mineralization. Of the 28,350 m drilled within the Arctic Deposit, 6,220 intervals were sampled, representing 12,434 m of sampled drilling. No other changes were made to the 2008 SRK estimation methodology and criteria as previously outlined. There were no changes made to the classification criteria, and no adjustments were made to metal pricing or operating costs related to the 2011 PEA. Table 14.7 shows the mineral resource estimate as reported March 9, 2012.

Table 14.7Arctic Deposit Resources at $75 NSR Cut-off – 2012 SRK

Resource		Grade					Contained Metal
Zone	kt	Cu (%)	Zn (%)	Pb (%)	Ag (g/t)	Au (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Ag (koz)	Au (koz)
Indicated
Zone 1	5,667	4.50	6.15	1.06	63.39	0.91	562	768	132	11,549	165
Zone 2	3,792	4.55	6.05	0.97	50.79	0.52	380	505	81	6,193	63
Zone 3	2,448	3.56	5.56	0.91	53.69	0.67	192	300	49	4,226	53
Zone 4	7,020	3.57	65.68	0.96	65.18	0.96	553	880	149	14,711	216
Zone 11	517	4.16	3.32	0.34	32.86	0.25	47	38	4	546	4
Total Indicated	19,445	4.05	5.81	0.97	59.55	0.80	1,735	2,491	415	37,226	501
Inferred
Zone 0	1,242	2.16	2.19	0.70	4.14	0.35	59	60	19	165	14
Zone 1	2,918	3.82	5.53	0.92	53.83	0.70	246	356	59	5,050	66
Zone 2	1,386	4.16	5.90	0.79	45.43	0.39	127	180	24	2,025	18
Zone 3	1,177	3.99	5.04	0.61	48.45	0.47	104	131	16	1,833	18
Zone 4	4,313	3.18	4.88	0.83	55.33	0.84	302	464	79	7,672	116
Zone 11	373	4.25	3.30	0.35	33.66	0.29	35	27	3	404	3
Total Inferred	11,409	3.47	4.84	0.80	46.75	0.64	873	1,217	201	17,149	235

Notes:	1.	g/t=ppm.
	2.	Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resources will be converted to Mineral Reserves.
	3.	Resources stated as contained within potentially economically minable underground shapes above a US$75.00/t NSR cut-off.
	4.	NSR calculation is based on assumed metal prices of $2.50/lb for copper, $1,000/oz for gold, $16.00/oz for silver, $1.00/lb for zinc and $1.00/lb for lead. A mining cost of $45.00/t and combined processing and G&A costs of $31.00 were assumed to form the basis for the resource NSR cut-off determination.
	5.	Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbers may not add due to rounding.

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14.3 RESOURCE ESTIMATION PROCEDURES

The resource evaluation methodology involved the following procedures:

•	database compilation and verification
•	review and construction of wireframe models
•	definition of resource domains
•	geostatistical analysis and variography
•	block modeling and grade interpolation
•	SG modelling and ABA assignment
•	resource classification and validation
•	assessment of “reasonable prospects for economic extraction” and selection of
	appropriate cut-off grades
•	preparation of the mineral resource statement.

14.4 DATABASE

The Arctic database was provided to Tetra Tech in a Microsoft Excel format (.csv). These .csv files contain collar, survey, assay, lithology, mineralization, specific gravity, and geotechnical data collected by NovaGold/NovaCopper and confirmed by Tetra Tech (see Section 12.1). The current drill hole database within the resource area consists of almost 10,323 samples from 135 drill holes. Table 14.8 provides a summary of the database used for the Arctic resource estimation.

Table 14.8 Exploration Data within the Resource Area

Company	Year	No. of Drill Holes	No. of Samples	Total Sample Length (m)
Kennecott (BCMC)	1967 to 1985	86	3,306	4,175
Kennecott	1998	6	548	964
NovaGold	2004 to 2008, 2011	43	6,469	16,564

Review of the database led Tetra Tech to the modification of the database for the resource estimation, as shown in Table 14.9.

Table 14.9 Modifications to Database for Resource Estimation

Action	Records Affected
Missing assay values, set to zero	57

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SG measurements have been collected by Kennecott (1998 and 1999) and NovaGold (2004 and 2011). By far the largest data set has been collected by NovaGold. Basic statistical analysis indicates there are no significant differences between the paired Kennecott and recently collected NovaGold SG values, or between the various measurement techniques.

14.5 GEOLOGICAL MODELLING

Geological modelling focused on the distribution of mineralization (massive and semi-massive sulphide), specific gravity and lithogeochemical models. NovaCopper supplied wireframes (.dxf files) of the topography and the geological models to Tetra Tech, who imported and verified the wireframes in Leapfrog^™ software (version 2.5.1) prior to importing into Isatis^™ software (version 2012.1) for use as estimation domain boundaries. The final estimation domain models represent a combination of these models with a specific focus on the mineralization domains for grade interpolation.

14.5.1 MINERALIZATION MODEL

In 2013, NovaCopper geologists updated the mineralization models, representing massive and semi-massive VMS-style mineralization. Geometrically, the mineralization is confined to six lenticular mineralized zones concentrated along an isoclinal fold hinge. Five of the six SMS zones contain a core of MS material (Figure 14.1).

14.5.2 SPECIFIC GRAVITY MODEL

In 2013, NovaCopper geologists recognized the effect of barium alteration on SG determinations. Using the geochemical and SG database, a total of four high-density (more than 3.5) SG domains were defined and modeled for the 2013 resource update (Figure 14.1)

14.5.3 LITHOGEOCHEMICAL MODEL

In 2013, NovaCopper geologists updated the lithogeochemical model (Twelker 2008) to include the 2011 drill results. Lithogeochemistry shows three major felsic rock suites in the Arctic Deposit area: a rhyolite suite; and intermediate volcanic suite and a volcanoclastic suite. NovaCopper currently models the following seven domains within the resource area: Undifferentiated Upper Plate, Felsite Top, Felsite Upper Limb, Felsite Fold, MRP Upper Limb, MRP Lower Limb and Grey Schist (Figure 14.1).

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Figure 14.1 Cross-section Illustrating the Arctic Deposit Geological Model

14.5.4 ESTIMATION DOMAINS

The final domain model, based on the mineralization, SG, and lithogeochemical models comprises 11 grade estimation domains, 11 SG estimation domains, and 9 ABA domains. Grade estimation domains are based on mineralization models. SG estimation domains are based on a combination of SG and mineralization models. ABA domains are based on a combination of lithogeochemical and mineralization models.

To test the validity of these models, and to determine the ideal method for treating the wireframe boundaries, contact profiles were generated and are discussed in Section 14.6.

The description of each domain and the equivalent coded value is listed in Table 14.10 to Table 14.12.

Table 14.10 Arctic Grade Estimation Domains

Name	Type	Domain	Volume (m3)	Thickness (m)
Amb1	SMS	10	2,097,100	-
Amb1	MS	11	1,155,500	-
Amb1	Total	-	3,252,600	6.5
Amb2	SMS	20	233,601	-

table continues…

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Name	Type	Domain	Volume (m3)	Thickness (m)
Amb2	MS	21	68,129	-
Amb2	Total	-	301,730	4.2
Amb3	SMS	30	1,954,481	-
Amb3	MS	31	865,319	-
Amb3	Total	-	2,819,800	6.5
Amb4	SMS	40	372,150	-
Amb4	MS	41	160,380	-
Amb4	Total	-	532,530	7.0
Amb5	SMS	50	3,225,515	-
Amb5	MS	51	41,285	-
Amb5	Total	-	3,266,800	5.6
Amb5a	SMS	60	11,479	-
Amb5a	MS	-	-	-
Amb5a	Total	-	11,479	0.7
Total	SMS	-	7,894,326	-
Total	MS	-	2,302,092	-
Total	Total	-	10,196,418	6.0

Note: MS = massive sulphide; SMS = semi-massive sulphide

Table 14.11 Arctic SG Estimation Domains

Name	Type	Domain	Volume (m3)
SG0	waste	0	Country Rock Volume
SG1	SMS	1	7,100,500
SG2	MS+HighSG	2	3,208,500

Table 14.12 Arctic ABA Domains

Wireframe	Description	Material Code
U Plate 2013	Undif. Upper Plate 50/50 split between MRP and Upper Felsite	1
Fels1-Top 2013	Felsite Top	2
MRP-Upper 2013	MRP Upper Limb	3
Fels1_Upper 2013	Felsite Upper Limb	4
GS 2013	Waste Grey Schist	5
Fels1-fold core 2013	Felsite Fold	6
MRP-Lower 2013	MRP Lower Limb – PAG/NAG Upper Limb ratios assumed	7
	If Material Code = 5 and Waste % < 100 then PAG	8

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14.6	EXPLORATORY DATA ANALYSIS
14.6.1	CONTACT PROFILES

Contact profiles were generated to test the validity of the wireframe models and to determine the ideal method for treating wireframe boundaries. Despite the narrow domains, the mineralization domains display hard boundaries with the wallrock for all metals of interest (as shown in Figure 14.2). Contact profiles also display hard boundaries between the MS and SMS domains (as shown in Figure 14.3).

Figure 14.2Contact Profile between Zone 1 SMS and Country Rock for Copper

Note: Numbers on graph are numbers of samples.

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Figure 14.3 Contact Profile between Zone 1 SMS and MS for Zinc

Note: Numbers on graph are numbers of samples.

14.6.2 RAW ASSAY DATA AND STATISTICS

Table 14.13 and Table 14.14 summarizes the raw drill hole assay and length statistics for the Arctic grade domains. They are tabulated by assayed metals copper, zinc, lead, gold and silver as well as by mineralization domain.

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Table 14.13 Raw Metal Sample Statistics (Length Weighted)

Mineralization Domain	Metal	Count	Minimum Grade	Maximum Grade	Mean Grade	Standard Deviation of Grade
SMS	Cu (%)	1,965	0.00	19.60	2.78	2.94
	Zn (%)	1,938	0.00	20.80	3.51	4.41
	Pb (%)	1,894	0.00	6.11	0.62	0.86
	Au (g/t)	1,846	0.00	32.80	0.74	2.31
	Ag (g/t)	1,942	0.17	669.00	42.32	54.79
MS	Cu (%)	1,045	0.01	18.35	4.49	3.27
	Zn (%)	1,036	0.02	25.60	7.56	5.31
	Pb (%)	1,019	0.00	8.04	1.26	1.16
	Au (g/t)	985	0.00	17.00	0.90	0.99
	Ag (g/t)	1,033	0.17	599.99	76.12	61.56

Table 14.14 Sample Length Statistics

Mineralization Domain	Count	Minimum (m)	Maximum (m)	Mean (m)	Standard Deviation	Variance
SMS	1966	0.01	3.23	0.62	0.60	0.36
MS	1045	0.01	3.10	0.64	0.60	0.35

14.7 COMPOSITING

Inspection of the raw assay data indicates that a common composite length of 2 m accommodates most sample lengths as the majority of sample lengths are less than 2 m in length. For resource estimation, the sample data was composited into a 2 m standard composite length for the deposit for the eleven individual mineralized semi-massive and massive sulphide zones.

14.7.1 OUTLIER MANAGEMENT AND CAPPING STRATEGY

When dealing with skewed populations, as well as outliers to the distribution, it is common practice in the industry to restrict the influence of high grade assays through “top-cutting” or “capping”. Capping was implemented on the 2 m composite assay data after sample length compositing. Capping limits were chosen for gold (six composites) and lead (one composite) as a function of the continuity-discontinuity of the high-grade “tail” of the metals histograms. Silver, copper and zinc were not capped.

The Arctic Deposit lead histogram high-grade “tail” shows poor coherence above 6% (Figure 14.4) and a grade cap of 6% was applied to the lead composites.

Similarly, the gold histogram high-grade “tail” shows poor coherence above 10 g/t (Figure 14.5). A cap of 10 g/t was applied for gold.

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The cap values were applied within both MS and to SMS mineralized zones.

Figure 14.4Lead Histogram

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Figure 14.5 Gold Histogram

14.7.2 COMPOSITE STATISTICS

The drill hole “capped” composite statistics by mineralization domain are shown in Table 14.15.

Table 14.15Arctic Deposit Drill Hole Composite Statistics

Mineralization Domain	Metal	Count	Minimum Grade	Maximum Grade	Mean Grade	Standard Deviation	Variance
SMS	Cu (%)	641	0.00	16.35	2.75	2.44	5.94
	Zn (%)	633	0.01	20.80	3.45	3.76	14.14
	Pb (%)	623	0.00	5.15	0.62	0.75	0.57
	Au (g/t)	618	0.00	28.52	0.72	1.96	3.85
	Ag (g/t)	635	0.17	416.24	41.70	44.80	2,007.25
MS	Cu (%)	351	0.01	18.00	4.52	3.02	9.11
	Zn (%)	350	0.02	24.10	7.64	4.66	21.70
	Pb (%)	347	0.00	6.70	1.26	1.03	1.07
	Au (g/t)	340	0.00	8.06	0.88	0.80	0.64
	Ag (g/t)	349	0.17	501.00	76.40	54.86	3,009.34

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14.8	SPECIFIC GRAVITY ANALYSIS
14.8.1	SPECIFIC GRAVITY STATISTICS AND SPATIAL ANALYSIS

The SG database for the Arctic Project contains 1,947 values, of which 878 are valid and in the resource area. Tetra Tech considers the data amenable for use in estimation of SG within most resource domains, due to the favourable spatial distribution of the samples. SG determination protocols and QA/QC review are discussed in detail in Section 11.0.

NovaCopper provided a set of SG wireframes delineating volumes where the specific gravity was observed to be greater than 3.5 and barite alteration was recorded. Tetra Tech combined these SG domains with the MS domains to produce a “high-density” domain for use in SG estimation. Outside of the “high-density domain” and within the SMS domain, a “medium-density” domain was defined. Material located outside the SG and mineralization models is defined by the “low-density” domain. The three density domains correlate to SG values “>3.5”, “<3.5” and “waste” respectively. Summary statistics for these measurements are provided in Table 14.16.

Table 14.16Arctic SG Determinations Summary Statistics

Domain	SG	Count	Minimum (m)	Maximum (m)	Mean (m)	Standard Deviation	Variance
High Density	> 3.5	128	2.83	4.64	3.80	0.48	0.23
Medium Density	< 3.5	54	2.72	4.26	3.06	0.37	0.14
Low Density	Waste	696	1.47	4.30	2.80	0.20	0.04
Total	-	878	1.47	4.64	2.96	0.44	0.19

14.8.2 SPECIFIC GRAVITY INTERPOLATION PLAN

SG was estimated into the block model (cells 10 m by 10 m by 5 m) by ID²methodology.

Results for density estimation are summarized in Table 14.17.

Table 14.17Arctic Density Interpolation Summary Statistics

SG Domain	Count	Minimum (m)	Maximum (m)	Mean (m)	Standard Deviation	Variance
High Density	6,417	2.68	4.58	3.42	0.49	0.24
Medium Density	14,201	2.97	4.15	3.11	0.28	0.08
Low Density	1,275,065	1.47	3.80	3.00	0.08	0.01
Total	1,295,683	1.47	4.58	3.01	0.10	0.01

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14.9 ACID BASE ACCOUNTING ANALYSIS

The current ABA database for the Arctic Project is limited to 148 samples (SRK 2010); however, some historical ABA sampling has been described by Kennecott. The 2010 program was designed to characterize waste based on an underground mine plan, thus current sampling is spatially concentrated around the semi-massive and massive sulphide lenses. Due to the limited spatial distribution of the ABA samples within the waste material, and based on a review of the ABA dataset, SRK recommended to use total sulphur percent as a proxy for acid potential (AP) and calcium percent as a proxy for neutralization potential (NP) to better characterize ABA in an open pit mine plan scenario. A total of 3,087 separate assay records totaling 6,073 m of samples within the resource area were tagged by lithogeochemical domain, and PAG/NAG ratios were determined for each domain based on the following steps:

• Acid Potential = 31.25 x Total S%

• Neutralization Potential = Ca% was converted to Ca kg CaCO3/t via the equation (Ca%x1000/40)

• Regression equations based on the Table 3 of the SRK July 2011 Preliminary ML/ARD Analysis, Ambler Project Arctic Deposit, Alaska Report were then used to calculate the PAG/NAG of each individual sample. ARD Classification criteria:

- If AP is less than 8 kg CaCO3/t the sample is NAG regardless of NP/AP (this is equivalent to 0.25% S).

- If AP is greater than 8 kg CaCO3/t, then NP/AP>3 defines NAG and NP/AP<3 defines PAG.

PAG/NAG ratios were assigned to the block model based on the lithogeochemical and mineralization domains (Table 14.18).

Table 14.18 PAG and NAG Ratio Assignment for the ABA Domains

Wireframe	Description	Material Code	Ratio PAG	Ratio NAG
U Plate 2013	Undif. Upper Plate 50/50 split between MRP and Upper Felsite	1	0.43	0.57
Fels1-Top 2013	Felsite Top	2	0.17	0.83
MRP-Upper 2013	MRP Upper Limb	3	0.55	0.45
Fels1_Upper 2013	Felsite Upper Limb	4	0.80	0.20
GS 2013	Waste Grey Schist	5	0.62	0.38
Fels1-fold core 2013	Felsite Fold	6	0.51	0.49
MRP-Lower 2013	MRP Lower Limb – PAG/NAG Upper Limb ratios assumed	7	0.55	0.45
	If Material Code = 5 and Waste % < 100 then PAG	8	1.00	0.00

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14.10 VARIOGRAPHY AND SPATIAL ANALYSIS

Variography was modelled utilizing Isatis^™ software. Variography (both downhole and spatial) was performed on composited drill hole data for semi-massive and massive sulphide domains. Representative variograms are shown in Figure 14.6 to Figure 14.9.

Table 14.19 shows variogram models for copper, zinc, lead, gold and silver within the two major semi-massive and massive domain types.

Table 14.19 Variogram Models

Domain	Metal	Nugget (Sill 0)	Sill 1	Range 1	Sill 2	Range 2
SMS	Copper	3.3	2.0	40; 40; 15	50; 50; 50	0.6
SMS	Zinc	8.3	2.1	30 ;30; 10	67; 67; 45	2.5
SMS	Lead	0.3	0.1	45; 45; 5	-	-
SMS	Gold	2.4	1.71	30; 30; 5	-	-
SMS	Silver	1,000	750	45; 45; 15	-	-
MS	Copper	3.0	5.9	45; 45; 14	-	-
MS	Zinc	8.0	13.2	45; 45; 16.5	-	-
MS	Lead	0.35	0.22	30; 30; 10	0.67	70; 70; 26
MS	Gold	0.33	0.2	45; 45; 20	-	-
MS	Silver	900	800	30; 30; 12	250	75; 75; 45

Figure 14.6 Copper: Variography in SMS Domains

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Figure 14.7 Zinc: Variography in SMS Domains

Figure 14.8 Copper: Variography in MS Domains

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Figure 14.9 Zinc: Variography in MS Domains

14.11 RESOURCE ESTIMATION METHODOLOGY

Resource estimation was completed using Isatis^™ (version 2012.1) software within an area encompassing all of the Arctic deposit with block model geometry and extents as presented in Table 14.20. As discussed in other sections, resource estimation methodology was based on the following:

• Assays were composited to 2 m and coded within each of 11 mineralization domains, and SG determinations were coded within each of the SG domains.

• Composites were capped for extreme outlier grades.

• During estimation, all of the estimation domains were treated as hard boundaries, preventing the sharing of composites across boundaries.

• Copper, zinc, lead, gold and silver were estimated using OK methodology.

• Whenever possible, SG was estimated using ID²methodology. All un-estimated blocks were assigned average SG values assessed separately for each domain.

• NAG and PAG ratios were assigned to the block model based on lithogeochemical and mineralization domains.

• Block model grades for the mineralized material portion of the block were calculated based on the proportion of MS and SMS within each block.

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• Whole block model grades were calculated based on the weighted value of the estimated grade item based on the proportion of the mineralization domain (i.e. (%Cu_MS x %MS) + (%Cu_SMS x %SMS) + (%Cu_Waste x % Waste). Whole block model SG values were calculated based on the weighted value of the SG based on the proportion of the high-density, medium -density and waste domains within the block. Whole blocks, including internal dilution, were used in pit optimization.

The selection of the search radii and rotations of search ellipsoids were guided by modelled ranges of continuity and were guided by modeled directions of continuity from variograms. In addition, the search radii parameters were established by conducting repeated test resource estimates and reviewing the results as a series of plan views and sections.

Hard estimation domain boundaries were applied to the model, which has the effect of constraining the data for estimation inside the narrow lenticular domain wireframes; therefore, effectively constraining the estimation without the need to apply extreme local anisotropy. Search distances were limited to a maximum of 300 m. The contribution to the estimates by data from distances further than twice the range of the variograms is very limited as the weighting placed on such samples is extremely low. Quantitative Kriging Neighbourhood Analysis (QKNA) was applied to optimize the search.

Variography was modeled separately for semi-massive and massive sulphide mineralization domains. In all cases, spherical variogram structures were applied. The orientation of the plane of greatest continuity (the flattening plane of the mineralized zones, variograms and search ellipsoid) was modelled on a plane striking at 80° and inclined at 15° to the south.

Search parameters were constrained within each mineralized domain and required an optimum number of 15 composites, minimum number of 5 composites, minimum number of 2 drillholes, and maximum search range of 200 m.

Table 14.20 Search Ellipse Orientation and Dimensions for Mineralization Domains

Domain	Min Samples per Estimate	Max Samples per Octant	Min Octants with Samples Octant	Max Search Perpendicular to Mineralization	Discretization
All	15	15	200	50	5x5x3

The SG ID² estimates were constructed in two steps. A flattened search ellipsoid with a range of 200 m in the mineralization plane and 40 m at right angles to the mineralization plane was applied with a minimum of 3 and maximum of 20 SG determinations required to estimate a block. The boundaries between the high-density, medium-density and waste domains were treated as hard boundaries as the change in SG across these boundaries is quite substantial. Final SG block values were calculated based on the estimated values and the weighted proportion of each domain in a block.

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14.12	RESOURCE BLOCK MODEL
14.12.1	CONFIGURATION

The block model is oriented with the model origin at the lower left hand corner (bottom southwest). All dimensions are in metres. The model is not rotated. Cell size and configuration was designed to facilitate best fit modelling of the narrow-vein geology of the Arctic mineralization. The block model configuration is recorded in Table 14.21. Proportions were used to define the portion of each block that lies below the topographic surface or inside the relevant mineralization model zones.

14.12.2 CELL ATTRIBUTES

Categorical, calculated and deterministic variables for the cells of the block model are tabulated in Table 14.21. Categorical variables include domains (10, 11, 20, 21, 30, 31, 40, 41, 50, 51 and 60) and resource classifications (Measured, Indicated and Inferred). Deterministic variables include, but are not limited to, gold grade, indicator probability (zero to one), density, and Kriging variance.

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Table 14.21 Block Model Cell Attributes

Value	Description
BM_10x10x5.csv X0,Y0,Z0 center of bottom leftmost block	-
NX=30; X0= 612400.00m; DX=10.00m	-
NY=105; Y0=7452600.00m; DY=10.00m	-
NZ=180; Z0=350.00m; DZ=5.00m	-
Xmin	612395
Xmax	613695
Ymin	7452595
Ymax	7453645
Zmin	347.5
Zmax	1247.5
Rotation: No rotation	-
Variables	-
SN,	Sequence number in block model
X,Y,Z	Block centroid
ID2_Ag,ID2_Au,ID2_Cu,ID2_Pb,ID2_Zn	ID2estimates for mineralized portions of each block
OK_Ag,OK_Au,OK_Cu,OK_Pb,OK_Zn	OK estimates for mineralized portions of each block – USED IN RESOURCE STATEMENT
Ore_percent	Percent of mineralized material in each 10x10x5 block
ID2_SG	SG estimated by ID2
CLASS	Resource classification 0= nothing, 1= inferred, 2 = Indicated (provisional, as resource cut-off will be based on economic analysis)
z_number	Orebody identifier, first digit 1=amb1, 2=amb2,5=amb5,6=amb5a; second digit, 0=semi- massive, 1=massive sulphide. This 21=amb2, massive sulphide portion
ID Ag BLK,ID Au BLK,ID Cu BLK,ID Pb BLK,ID Zn BLK	ID2estimates for entire block (including waste and ore)
OK Ag BLK,OK Au BLK,OK Cu BLK,OK Pb BLK,OK Zn BLK	OK estimates for entire block (including waste and ore) – USED IN GRADE SENSITIVITY ANALYSIS
MM_nn_Xx	Number of samples used to estimate blocks (MM=method, Xx=element)
MM_meandist_Xx	To estimating samples

table continues…

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Value	Description
Xx_sM semi-massive and XxMS Massive Sulphide estimates in overlap zones	-
Prop_sM	Proportion of semi massive sulphide in mixed blocks
Prop_MS	Proportion of massive sulphide in mixed blocks
SG_1	SG ID2estimates into (0) waste material (1) semi massive zones outside high specific gravity zones and (2) massive sulphide zones united with high SG zones
SG_blk	SG ID2estimates adjusted for margins around MS and semi massive units – USED IN RESOURCE STATEMENT & GRADE SENSITIVITY ANALYSIS

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14.13 MODEL VALIDATION

The reported resource relies on the OK method as the best non-biased interpolator of grade. ID² was employed for model validation purposes. The block model was validated using the following tools:

• Visual comparisons of drill hole composites with resource block grades.

• Statistical comparisons between block and composite data.

• Swath plot comparisons (drift analysis) comparing the OK model with the ID² model.

• Scatter plot comparisons of the OK estimates with the ID²estimates.

• Grade and tonnage comparisons of the OK and ID²estimates over a range of cut-off grades.

14.13.1 VISUAL VALIDATION

Figure 14.10 through Figure 14.14 shows representative west-east cross sections for the Arctic deposit, for copper, zinc, lead, gold and silver. Visual inspection of these sections conforms with observed drill hole data and interpreted geology.

Figure 14.10 Cross-section for Copper

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The centre cores of the zones, contain MS concentrations where higher grade drill intercepts were obtained. The colour-coding in Figure 14.11 to Figure 14.14 shows higher estimated block grades in similar positions reflecting the zonation observed from the drill intercepts.

Figure 14.11Cross-section for Zinc

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Figure 14.12 Cross-section for Lead

Figure 14.13 Cross-section for Gold

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Figure 14.14Cross-section for Silver

14.13.2 STATISTICAL COMPARISONS BETWEEN BLOCK AND COMPOSITE DATA

SWATH PLOTS

Figure 14.15 through Figure 14.18 display representative swath plots for the Arctic Deposit block model OK and ID² block estimates.

The algorithms used for OK and ID² estimation are different, yet the similar grade patterns for both methods across the mineralized zones as shown in Figure 14.15 to Figure 14.18, demonstrates that the independent estimation methods provide similar outcomes which provides a degree of assurance for the estimation process.

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Figure 14.15 Swath Plot of Block Copper Grade Values by Northing for the Arctic Deposit

Figure 14.16 Swath Plot of Block Copper Grade Values by Easting for the Arctic Deposit

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Figure 14.17 Swath Plot of Block Zinc Grade Values by Northing for the Arctic Deposit

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Figure 14.18 Swath Plot of Block Zinc Grade Values by Easting for the Arctic Deposit

SCATTER PLOTS

Figure 14.19 to Figure 14.23, show scatter plots comparing the OK estimates and the ID² estimates within the mineralized domains. The regression coefficients and regression lines are shown on each graph with a better than 90% correlation between the two estimation methodologies. A 90% correlation is sufficient to validate the estimates.

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Figure 14.19 Copper: Scatter Plot of OK versus ID Estimates

Figure 14.20 Zinc: Scatterplot of OK versus ID Estimates

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Figure 14.21 Lead: Scatterplot of OK versus ID Estimates

Figure 14.22 Gold: Scatterplot of OK versus ID Estimates

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Figure 14.23 Silver: Scatterplot of OK versus ID Estimates

GRADE AND TONNAGE COMPARISONS

Grade tonnage curves are graphs showing the progressive changes in tonnage and grade as the cut-off grade is increased for a particular mineralized deposit. These graphs provide a measure of the sensitivity of the mineralization in question to changes in cut-off grade which may be dictated by other factors such as mineral price and mining costs. The steeper the curve, the more sensitive is the grade or tonnage to changes in the cutoff grade. The graphs also provide a convenient measure of comparison between different estimation methodologies. The more different the shapes of the grade or tonnage curves, the more different the estimation of the blocks. Grade tonnage comparisons for OK and ID² estimations are shown in Figure 14.24 to Figure 14.28 for the copper, zinc, lead, gold and silver respectively. In all five cases, the similarities in the graphs provide comfort that the block estimation methods have generated similar results. The marginally higher grade curves and marginally lower tonnage curves for the ID² estimates, suggests that the ID² estimation provides more weighting to the composites closer to the estimated blocks.

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Figure 14.24 Block Model Grade Tonnage Comparison: Copper OK (red line) and Copper ID² (broken line)

Figure 14.25 Block Model Grade Tonnage Comparison: Zinc OK (red line) and Zinc ID² (broken line)

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Figure 14.26 Block Model Grade Tonnage Comparison: Lead OK (red line) and Lead ID² (broken line)

Figure 14.27 Block Model Grade Tonnage Comparison: Gold OK (red line) and Gold ID² (broken line)

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Figure 14.28Block Model Grade Tonnage Comparison: Silver OK (red line) and Silver ID²(broken line)

14.14 MINERAL RESOURCE CLASSIFICATION

Block model quantities and grade estimates for the Arctic Project were classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves (adopted by the CIM Council on December 11, 2005) and incorporated by reference in NI 43-101 by Mike O’Brien, Pr.SciNat (SACNASP#400295/87).

A “Measured Mineral Resource” is that part of the total resource for which the physical characteristics are well established for use in production planning and economic evaluation; data is sufficient enough to confirm both geological and grade continuity. An “Indicated Mineral Resource” is that part of the total resource for which the physical characteristics are well established for use in production planning and economic evaluation; data is sufficient enough to reasonably assume, but cannot verify, geological and grade continuity. An “Inferred Mineral Resource” is that part of the total resource for which the quantity and grade can be estimated; data is sufficient enough to reasonably assume both geological and grade continuity.

Mineral resource classification is typically a subjective concept; industry best practices suggest that resource classification should consider, but not be limited to, drill and sample spacing, the deposit-type and mineralization continuity, surface and/or underground mineralization exposure, variography, kriging efficiency (KE), theoretical slope of regression (ZZ*) and/or prior mining experience. With respect to this resource classification, model cells are assigned classification on the basis of interpolation search and drill data density. Solid wireframes or strings were used to define the Inferred and Indicated resources.

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14.14.1	MEASURED RESOURCE CLASSIFICATION
	No resource is classified as Measured as the drill intercepts are rarely at shorter distances than the variograms ranges and the samples taken within the relatively narrow mineralized zones exhibit a relatively high nugget effect.
14.14.2	INDICATED RESOURCE CLASSIFICATION
	Estimated blocks which have been interpolated within a distance of 75 m from at least two drill holes and within the modelled wireframes of the mineralized zones, are classified as Indicated. The observed geometric continuity of the higher grade zones which can be visually identified by the presence of sulphides provides additional comfort for the interpolation between drill holes.
14.14.3	INFERRED RESOURCE CLASSIFICATION
	Estimated blocks which have been interpolated within a distance of 200 m from at least two drill holes and within the modeled wireframes of the mineralized zones, are classified as an Inferred Resource with the exception of the Indicated Resource.
14.15	MINERAL RESOURCE ESTIMATE
	CIM Definition Standards for Mineral Resources and Mineral Reserves (November 2010) defines a mineral resource as:

“(A) concentration or occurrence of diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals in or on the Earth’s crust in such a form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge”.

The “reasonable prospects for economic extraction” requirement generally implies that the quantity and grade estimates meet certain economic thresholds and that the mineral resources are reported at appropriate cut-off grade taking into account extraction scenarios and processing recoveries. In order to meet this requirement, Tetra Tech considers that major portions of the Arctic Project are amenable for open pit extraction.

In order to determine the quantities of material offering “reasonable prospects for economic extraction” by an open pit, Tetra Tech used a pit optimizer and reasonable mining assumptions to evaluate the proportions of the block model (Indicated and Inferred blocks) that could be “reasonably expected” to be mined from an open pit.

The optimization parameters were selected based on experience and benchmarking against similar projects (Table 16.2 in Section 16.0). The reader is cautioned that the results from the pit optimization are used solely for the purpose of testing the “reasonable prospects for economic extraction” by an open pit and do not represent an attempt to estimate mineral reserves. The results are used as a guide to assist in the preparation of a mineral resource statement and to select an appropriate resource reporting cut-off grade.

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The pit optimization, economic, technical and operational parameters and as well as the metallurgical recoveries used for pit and mine schedule optimizations are provided in Table 16.1 and Table 16.2 in Section 16.0.

Table 14.22 Indicated Mineral Resource Table Restated by Mineralization Zone, Arctic Project, Alaska, Tetra Tech (July 30, 2013)

Indicated	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Blb)	Zn (Blb)	Pb (Blb)	Au (Moz)	Ag (Moz)
amb1	8.053	2.53	3.04	0.49	0.68	50.3	450	540	87.1	0.18	13.0
amb2	0.930	3.20	7.17	1.30	0.77	68.3	66	147	26.7	0.02	2.0
amb3	5.955	3.21	3.82	0.64	0.42	38.1	449	501	84.4	0.08	7.3
amb4	1.393	4.42	9.12	1.58	0.79	89.8	136	280	48.6	0.04	4.0
amb5	7.517	3.70	5.25	0.93	0.95	59.5	613	870	153.9	0.23	14.4
Total	23.848	3.26	4.45	0.76	0.71	53.2	1,713	2,338	400.9	0.55	40.8

Table 14.23 Inferred Mineral Resource Table Restated by Mineralization Zone, Arctic Project, Alaska, Tetra Tech (July 30, 2013)

Inferred	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Blb)	Zn (Blb)	Pb (Blb)	Au (Moz)	Ag (Moz)
amb1	1.120	3.07	4.25	0.69	0.85	57.2	76	105	17.1	0.03	2.1
amb2	0.038	3.19	5.50	0.92	1.06	66.4	3	5	0.8	0.00	0.1
amb3	1.354	3.18	3.27	0.41	0.31	27.5	95	98	12.4	0.01	1.2
amb4	0.000	0.00	0.00	0.00	0.00	0.0	0	0	0.0	0.00	0.0
amb5	0.851	3.48	4.03	0.67	0.68	41.7	65	76	12.6	0.02	1.1
Total	3.363	3.22	3.84	0.58	0.59	41.5	239	285	43.2	0.06	4.5

The following notes apply to Table 14.22 and Table 14.23:

1.	These resource estimates have been prepared in accordance with NI 43-101 and the CIM Definition Standards. Mineral resources that are not mineral reserves do not have demonstrated economic viability. Inferred resources have a great amount of uncertainty as to their existence and whether they can be mined legally or economically. It cannot be assumed that all or any part of the Inferred resources will ever be upgraded to a higher category.
2.	Mineral Resources are reported within mineralization wireframes, contained within an Indicated and Inferred pit design using an assumed copper price of $2.90/lb, zinc price of $0.85/lb, lead price of $0.90/lb, silver price of $22.70/oz, and gold price of $1,300/oz.
3.	Appropriate mining costs, processing costs, metal recoveries and inter ramp pit slope angles were used to generate the pit design.
4.	The $35.01/t milled cut-off is calculated based on a process operating cost of $19.03/t, G&A of $7.22/t and site services of $8.76/t. NSR equals payable metal values, based on the metal prices outlined in Note 2 above, less applicable treatment, smelting, refining costs, penalties, concentrate transportation costs, insurance and losses and royalties.
5.	The LOM strip ratio is 8.39.
6.	Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content.
7.	Tonnage and grade measurements are in metric units. Contained copper, zinc and lead pounds are reported as imperial pounds, contained silver and gold ounces as troy ounces.

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14.16 GRADE SENSITIVITY ANALYSIS

The mineral resources of the Arctic Project are sensitive to the selection of the mining methodology. To illustrate this sensitivity, the global block model quantities and grade estimates within the mineralization domains and within the whole blocks (including internal dilution within a mineable block size of 10 m by 10 m by 5 m) are presented in Table 14.24 for Indicated resources and in Table 14.25 for Inferred resources, at an $35/t NSR cut-off value. The reader is cautioned that the values presented in these tables should not be misconstrued with a Mineral Resource Statement. The tables are only presented to show the sensitivity of the block model estimates to the mining block internal dilution.

Table 14.24 Material Comprising the Indicated Mineral Resource (Table 14.22) Diluted to Full Blocks, Restated by Mineralization Zone, Arctic Project, Alaska

Indicated	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Blb)	Zn (Blb)	Pb (Blb)	Au (Moz)	Ag (Moz)
Mineralized Only	23.848	3.26	4.45	0.76	0.71	53.2	1,713	2,338	400.9	0.55	40.8
With Internal Block Dilution	32.808	2.37	3.23	0.55	0.52	38.7	1,713	2,338	400.9	0.55	40.8

Note: Based on a cut-off calculated from NSR plus infrastructure cost of $35/mineralized tonne. Details are shown in Section 16.0.

Table 14.25 Material Comprising the Inferred Mineral Resource (Table 14.23) Diluted to Full Blocks, Restated by Mineralization Zone, Arctic Project, Alaska

Inferred	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Blb)	Zn (Blb)	Pb (Blb)	Au (Moz)	Ag (Moz)
Mineralized Only	3.363	3.22	3.84	0.58	0.59	41.5	239	285	43.2	0.06	4.5
With Internal Block Dilution	5.102	2.12	2.53	0.38	0.39	27.3	239	285	43.2	0.06	4.5

Note: Based on a cut-off calculated from NSR plus infrastructure cost of $35/mineralized tonne. Details are shown in Section 16.0.

The sensitivity of the mineral resources of the Arctic Project to the economic cut-off is represented by selecting material above different NSR values. The global block model quantities and grade estimates within the mineralization domains and within the whole blocks (including internal dilution within a mineable block size of 10 m by 10 m by 5 m) are presented in Table 14.26 for Indicated resources and in Table 14.27 for Inferred resources, at an $35/t (the current base case) and at elevated $40/t and $45/t NSR cut-off values. The reader is cautioned that the values presented in these tables should not be misconstrued with a Mineral Resource Statement. The tables are only presented to show the economic sensitivity of the block model estimates.

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Table 14.26 Indicated Mineral Resource Table Sensitivity to NSR Within Whole Blocks

Indicated	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Au (koz)	Ag (Moz)
NSR 35	32.808	2.37	3.23	0.55	0.52	38.7	1,713	2,338	400.9	5.5	40.8
NSR 40	31.937	2.37	3.30	0.57	0.53	39.4	1,667	2,324	398.5	5.4	40.5
NSR 45	31.074	2.42	3.36	0.58	0.54	40.2	1,658	2,302	397.3	5.4	40.1

Table 14.27 Inferred Mineral Resource Table Sensitivity to NSR, Arctic Project Within Whole Blocks

Inferred	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Au (koz)	Ag (Moz)
NSR 35	5.102	2.12	2.53	0.38	0.39	27.3	239	285	43.2	0.6	4.5
NSR 40	4.918	2.17	2.58	0.39	0.40	28.0	235	280	42.3	0.6	4.4
NSR 45	4.724	2.23	2.64	0.40	0.41	28.7	232	275	41.7	0.6	4.4

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15.0 MINERALRESERVEESTIMATES

In this technical report, the terms “mineral reserve”, “probable mineral reserve” and “proven mineral reserve” have the meanings ascribed to those terms by the CIM, as the CIM Definition Standards on Mineral Resources and Mineral Reserves adopted by CIM Council, as amended.

A “mineral reserve” has not been estimated for the Arctic Project as part of this PEA.

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16.0 MININGMETHODS

16.1	INTRODUCTION
	Tetra Tech has prepared an open pit mining study to support development of an updated PEA for the Arctic Project. This section outlines the input data, procedures and results of this PEA-level open pit optimization, design, mine production scheduling and associated costing.
16.2	PIT OPTIMIZATION
	Tetra Tech performed the open pit optimizations using Gemcom Whittle™software, which is based on the Lerchs -Grossmann (LG) optimization algorithm. Pit optimization parameters were prepared from first principle estimates, technical studies and experience from other projects.
16.2.1	BLOCK MODEL
	Mineral Resources for the Arctic Project were prepared by Tetra Tech. This work is documented in Section 14.0, and forms the basis of the open pit optimizations. A block model was provided to the mining team as a MineSight™file with 10 m by 10 m by 5 m blocks.
16.2.2	PIT SLOPE ANGLE
	An overall pit slope angle of 43° was used in the open pit optimizations. This pit slope angle is based on limited available geotechnical data, conservative estimates from previous experience and widely accepted empirical evidence (EBA 2013).
16.2.3	SURFACE TOPOGRAPHY
	NovaCopper provided digital topographical drawings to Tetra Tech in NAD 83 for the Arctic Project area from 2010. The topography can be seen in Figure 16.1.
16.2.4	PIT OPTIMIZATION PARAMETERS
	The economic, technical and operational parameters as well as the metallurgical recoveries used for pit and mine schedule optimizations are provided in Table 16.1 and Table 16.2.

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Figure 16.1 Arctic Project Topography

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Table 16.1 Pit Optimization Parameters

Item		Units	Value	Cu Concentrate	Pb Concentrate	Zn Concentrate
Production Rate	Mineralized material processed	t/day	10,000	-	-	-
Metal Prices	Gold	$/oz	1,300.00	-	-	-
	Copper	$/lb	2.90	-	-	-
	Silver	$/oz	22.70	-	-	-
	Lead	$/lb	0.90	-	-	-
	Zinc	$/lb	0.85	-	-	-
Royalty	-	% of NSR	1.0	-	-	-
Concentrate	Concentrate grade	% Cu/Pb/Zn	-	29.0	50.0	56.0
	Gold in concentrate	g/t	-	Variable	Variable	Variable
	Silver in concentrate	g/t	-	Variable	Variable	Variable
	Moisture content	%	-	9.0	9.0	9.0
Concentrate Transportation	Trucking	$/wmt	-	66.96	66.96	66.96
	Port	$/wmt	-	20.00	20.00	20.00
	Ocean freight	$/wmt	-	63.00	63.00	63.00
	Total	$/wmt	-	149.96	149.96	149.96
Smelter Deductions	Concentrate unit deduction	Units of concentrate grade (%)	-	1.00	3.00	8.50
	Gold, deductions	%	-	5.00	5.00	No payment
	Silver, deductions	%	-	10.00	10.00	No payment
Penalties	Element 1 – Arsenic	$/dmt	-	-	4.46	-
	Element 2 – Antimony	$/dmt	-	-	1.50	-
	Element 3 – Mercury	$/dmt	-	0.01	-	-
	Element 4 – Lead	$/dmt	-	0.45	-	-
	Element 5 – Zinc	$/dmt	-	1.95	-	-
	Total	$/dmt	-	2.41	5.96	0.00

table continues…

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Item		Units	Value	Cu Concentrate	Pb Concentrate	Zn Concentrate
Others	Concentrate losses (during transport and rehandling)	%	-	0.42	0.42	0.42
	Insurance	% NIV	-	0.15	0.15	0.15
	Representation/marketing	$/wmt	-	2.50	2.50	2.50
Smelting/Refining Terms	Smelting	$/dmt	-	70.00	180.00	300.00
	Refining		-	-	-	-
	- copper	$/lb	-	0.07	-	-
	- gold	$/oz	-	10.00	10.00	10.00
	- silver	$/oz	-	0.60	0.60	0.60
	- lead	$/lb	-	-	-	-
	- zinc	$/lb	-	-	-	-
Operating Cost	Mining (mineralized material or waste)	$/t mined	3.00	-	-	-
	Added haul cost/bench	$/t mined	0.02	-	-	-
	Processing	$/t milled	19.03	-	-	-
	Surface services	$/t milled	8.76	-	-	-
	G&A	$/t milled	7.22	-	-	-
Block Model	Size	m	10 by 10 by 5	-	-	-
Density	Mineralized material and waste rock	t/m3	Variable	-	-	-
	Default SG (mineralized material)	t/m3	3.40	-	-	-
	Default SG (waste)	t/m3	2.80	-	-	-
Mining Dilution	-	%	5	-	-	-
Mining Recovery	-	%	95	-	-	-
Pit Slope Angle	-	degrees	43	-	-	-

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Table 16.2 Metallurgical Recovery

Head Grade	Recovery
Head, % Cu	Recovery, % Cu
>5.0	92.00
2.0-5.0	4.1209 x ln(head grade) + 83.652
0.5-2.0	-3.6363 x (head grade)2 + 16.08 x (head grade) + 68.792
0.3-0.5	50.00
0.1-0.3	20.00
<0.1	0.00
Head, % Pb	Recovery, % Pb
>2.0	85.00
0.5-2.0	3.7128 x ln(head grade) + 76.499
0.3-0.5	8.0157 x ln(head grade) + 79.365
0.2-0.3	50.00
<0.2	0.00
Head, % Zn	Recovery, % Zn
>7.0	92.00
3.0-7.0	5.159 x ln(head grade) + 81.06
1.5-3.0	8.940 x ln(head grade) + 76.80
1.0-1.5	75.00
0.5-1.0	50.00
<0.5	0.00
Head, g/t Ag	Recovery, % Ag
>100.0	88.00
40.0-100.0	6.102 x ln(head grade) + 58.8
20.0-40.0	-0.012 x (head grade)2 + 1.197 x (head grade) + 52.54
10.0-20.0	50.00
5.0-10.0	25.00
<5.0	0.00
Head, g/t Au	Recovery, % Au
>1.5	85.00
0.51-1.5	12.303 x ln(head grade) + 74.072
0.1-0.51	26.003 x ln(head grade) + 83.069
<0.1	0.00

16.2.5 PIT OPTIMIZATION RESULTS

Using the pit optimization parameters in Table 16.1 and the process recoveries in Table 16.2, 49 pit shells were generated using Gemcom Whittle^™software, corresponding to price factors ranging between 0.292 and 1.2. Optimizations were run using Indicated and Inferred mineral resources. The discounted value of each pit was estimated using a discount rate of 8%. All operating costs and smelting/refining terms in Table 16.1 were considered in the discounted values. No capital costs were considered in generating these values. The pit optimization results are summarized in Table 16.3. Based on the discounted value, pit 27 was selected to be the final pit. This pit corresponds to a price factor of 0.76.

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Table 16.3 Pit Optimization Results

Pit	Price Factor	Discounted Value ($000)	Total Tonnes Mined (t)	Mineralized Material Mined (t)	Cu Grade (%)	Pb Grade (%)	Zn Grade (%)	Au Grade (g/t)	Ag Grade (g/t)
1	0.292	1,241,029	91,506,034	11,073,401	2.75	0.68	4.02	0.53	42.38
2	0.300	1,354,602	99,811,925	12,347,334	2.70	0.69	4.04	0.54	41.97
3	0.320	1,587,243	118,832,054	15,646,729	2.59	0.65	3.82	0.55	41.36
4	0.340	1,644,715	123,679,698	16,726,285	2.55	0.64	3.74	0.54	40.77
5	0.360	1,716,797	131,398,298	18,199,897	2.50	0.62	3.66	0.53	40.05
6	0.380	1,752,065	135,449,322	18,952,025	2.48	0.62	3.61	0.53	39.71
7	0.400	2,042,751	181,898,688	24,958,036	2.39	0.58	3.41	0.51	38.07
8	0.420	2,137,932	200,210,781	27,287,278	2.36	0.57	3.36	0.50	37.65
9	0.440	2,202,656	217,809,344	28,987,367	2.35	0.56	3.32	0.50	37.47
10	0.460	2,223,852	223,877,773	29,639,952	2.34	0.56	3.30	0.50	37.33
11	0.480	2,253,074	232,939,309	30,526,639	2.34	0.55	3.27	0.50	37.15
12	0.500	2,269,410	238,625,192	31,067,030	2.33	0.55	3.26	0.50	37.04
13	0.520	2,283,094	244,218,775	31,660,921	2.32	0.55	3.24	0.50	36.90
14	0.540	2,287,084	245,881,130	31,896,102	2.32	0.55	3.23	0.49	36.82
15	0.560	2,297,763	251,390,824	32,459,969	2.31	0.54	3.20	0.49	36.67
16	0.580	2,302,132	254,173,635	32,767,578	2.30	0.54	3.19	0.49	36.57
17	0.591	2,305,059	256,019,634	32,958,694	2.30	0.54	3.18	0.49	36.52
18	0.592	2,348,788	282,491,248	34,885,340	2.26	0.54	3.16	0.49	36.47
19	0.600	2,352,851	285,511,537	35,149,503	2.26	0.54	3.15	0.49	36.42
20	0.620	2,356,262	288,574,043	35,554,707	2.25	0.53	3.14	0.49	36.29
21	0.640	2,359,157	291,915,359	35,936,186	2.24	0.53	3.12	0.49	36.17
22	0.660	2,365,531	299,599,350	36,452,825	2.24	0.53	3.11	0.49	36.11
23	0.680	2,367,999	302,094,218	36,675,567	2.23	0.53	3.10	0.49	36.06
24	0.700	2,369,087	303,522,442	36,836,105	2.23	0.53	3.09	0.49	36.00
25	0.720	2,373,382	310,689,985	37,311,249	2.22	0.52	3.08	0.49	35.91
26	0.740	2,374,653	314,359,773	37,619,157	2.22	0.52	3.06	0.49	35.83
27	0.760	2,377,134	321,537,575	38,015,718	2.21	0.52	3.05	0.49	35.78
28	0.780	2,377,092	322,246,524	38,153,413	2.21	0.52	3.05	0.48	35.73
29	0.800	2,376,939	324,893,126	38,353,791	2.20	0.52	3.04	0.48	35.69
30	0.820	2,376,783	325,186,003	38,421,686	2.20	0.52	3.03	0.48	35.65
31	0.840	2,376,499	329,107,590	38,614,089	2.20	0.52	3.03	0.48	35.61
32	0.860	2,375,843	332,799,860	38,847,858	2.20	0.51	3.02	0.48	35.55
33	0.880	2,375,457	333,844,770	38,925,548	2.19	0.51	3.01	0.48	35.53
34	0.900	2,374,381	337,373,626	39,168,160	2.19	0.51	3.01	0.48	35.46
35	0.920	2,373,503	338,709,274	39,276,488	2.19	0.51	3.00	0.48	35.42
36	0.940	2,373,079	339,144,938	39,346,029	2.18	0.51	3.00	0.48	35.39

table continues…

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Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Pit	Price Factor	Discounted Value ($000)	Total Tonnes Mined (t)	Mineralized Material Mined (t)	Cu Grade (%)	Pb Grade (%)	Zn Grade (%)	Au Grade (g/t)	Ag Grade (g/t)
37	0.960	2,372,004	342,213,188	39,546,123	2.18	0.51	2.99	0.48	35.34
38	0.980	2,371,592	342,927,941	39,615,040	2.18	0.51	2.99	0.48	35.31
39	1.000	2,370,194	344,840,391	39,751,273	2.17	0.51	2.98	0.48	35.27
40	1.020	2,369,568	345,313,884	39,805,458	2.17	0.51	2.98	0.48	35.25
41	1.040	2,368,753	346,263,522	39,891,956	2.17	0.51	2.98	0.48	35.21
42	1.060	2,367,822	347,119,812	39,981,247	2.17	0.51	2.97	0.48	35.17
43	1.080	2,365,527	351,324,684	40,151,750	2.16	0.51	2.97	0.48	35.15
44	1.100	2,364,682	352,623,409	40,206,627	2.16	0.51	2.97	0.48	35.13
45	1.120	2,363,807	354,571,214	40,320,322	2.16	0.51	2.96	0.48	35.10
46	1.140	2,362,792	356,748,523	40,408,292	2.16	0.51	2.96	0.48	35.07
47	1.160	2,362,653	356,972,620	40,419,045	2.16	0.51	2.96	0.48	35.07
48	1.180	2,362,343	357,476,805	40,443,693	2.16	0.51	2.96	0.48	35.06
49	1.200	2,361,902	358,442,346	40,464,900	2.16	0.51	2.96	0.48	35.06

16.3	MINE DESIGN
16.3.1	BENCH HEIGHT AND PIT WALL SLOPE

Limited geotechnical data is available for the Arctic Project. Current open pit design is based on the recommendations by EBA (2013). The pit design incorporates a bench height of 5 m and a 45° inter-ramp angle. After adding the ramps, the overall slope angle will be within 43°, as recommended by EBA (2013). Pit excavation will be triple benched as shown in Figure 16.2.

Figure 16.2 Pit Wall Slope

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Preliminary Economic Assessment Report on the Arctic
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16.3.2 MINIMUM WORKING AREA

Benches have been designed to accommodate an 11.0 m³ diesel hydraulic shovel and a 90 t truck. At the pit bottom, there are areas where mining will be reduced to 40 m width. Tetra Tech recommends the use of a front-end loader and 43.5 t articulated trucks for such areas where available space restricts production.

16.3.3 HAUL ROAD

Main haul roads for the project were designed to accommodate 90 t haul trucks with two-way traffic and one-way traffic for the three lowest benches of the pit. Haul road design details can be seen in Table 16.4, Figure 16.3 and Figure 16.4. Ramps are designed with a maximum grade of 10%.

Table 16.4 Haul Road Width

Haul Road Width
Traffic	One-way (m)	Two way (m)
Running Surface	13.0	22.7
Safety Berm	5.3	5.3
Ditch	2.0	2.0
Total	20.3	30.0

Figure 16.3 One-way Haul Road

NovaCopper Inc.	16-8	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Figure 16.4 Two-way Haul Road

16.3.4 PIT HYDROLOGY/DEWATERING

No pit hydrology/dewatering were included in this PEA study. However, an allowance is included in the mining operating cost to account for pit dewatering.

16.3.5 PIT DESIGN RESULTS

After applying the design criteria, the final pit includes 35.7 Mt of mineral resource with a LOM strip ratio of 8.39. A material summary from the final pit can be seen in Table 16.5. Figure 16.5 shows a general view of the final pit.

Table 16.5 Pit Design Results

Material	Mass (Mt)	Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)
Mineralized Material	35.7	2.2792	0.5306	3.1259	0.4978	36.9145
Waste – NAG	134.6	-	-	-	-	-
Waste – PAG	164.8	-	-	-	-	-

NovaCopper Inc.	16-9	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Figure 16.5 Pit Design

NovaCopper Inc.	16-10	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

16.4 PRODUCTION SCHEDULE

The mining schedule was developed based on a maximum mill capacity of 10,000 t/d. The Arctic Project’s total mine life is 13 years, including 1 year of pre-stripping followed by 12 years of production. The pit uses four pushbacks and a minimum mining width of 40 m. As shown in Table 16.6, over the 13-year life, the pit is producing 35.7 Mt of mineralized material and 299.4 Mt of waste rock. The LOM stripping ratio is 8.39 and the stripping ratio excluding the pre-stripping waste rock is 7.94. The mining schedule does not currently consider a low-grade stockpiling option but this can be assessed in more detail in future studies. The cut-off grade was optimized to maximize the discounted value of the Arctic Project. Due to operational and geometric constraints, the mine will require a ramp-up period of three years before achieving full mill capacity. Figure 16.6 shows the production schedule indicating the total mined waste and total mined mineralized material fed to the process plant. Figure 16.7 through Figure 16.10 show the status of mining activity at the end of the pre-production year (-1) and years 4, 8, and 12 (end of mine life).

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Preliminary Economic Assessment Report on the Arctic
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Table 16. 6 Production Schedule

Year	Mineralized Material Processed (t)			Waste Production (t)			Strip Ratio	Total Material Mined (t)	Head Grades
	Inferred	Indicated	Total	PAG	NAG	Total			Cu (%)	Pb (%)	Zn (%)	Au (g/t)	Ag (g/t)
-1	0	0	0	7,187,581	8,812,419	16,000,000	-	16,000,000	0.00	0.00	0.00	0.00	0.00
1	352,156	648,937	1,001,092	16,228,292	14,770,616	30,998,908	30.97	32,000,000	2.42	0.41	2.71	0.45	28.53
2	32,439	1,905,170	1,937,609	15,775,430	14,286,961	30,062,391	15.52	32,000,000	2.38	0.41	2.86	0.36	35.49
3	25,245	2,682,237	2,707,483	15,579,203	13,713,314	29,292,517	10.82	32,000,000	2.14	0.54	3.13	0.46	34.94
4	31,759	3,188,537	3,220,296	15,312,509	13,467,195	28,779,704	8.94	32,000,000	2.15	0.61	3.57	0.46	34.32
5	22,219	3,627,781	3,650,000	16,034,444	12,307,138	28,341,582	7.76	31,991,582	2.58	0.69	3.97	0.57	41.43
6	110,237	3,539,763	3,650,000	12,660,879	9,307,401	21,968,279	6.02	25,618,279	2.46	0.52	2.94	0.57	39.79
7	298,623	3,351,377	3,650,000	12,660,879	9,307,401	21,968,279	6.02	25,618,279	2.44	0.55	3.36	0.50	39.39
8	724,295	2,507,152	3,231,448	12,660,879	9,307,401	21,968,279	6.8	25,199,727	2.38	0.58	3.42	0.45	37.97
9	784,841	2,579,931	3,364,772	12,660,879	9,307,401	21,968,279	6.53	25,333,051	2.05	0.41	2.57	0.43	30.41
10	493,525	3,156,475	3,650,000	12,660,879	9,307,401	21,968,279	6.02	25,618,279	2.38	0.51	3.20	0.50	38.43
11	1,048,358	2,601,642	3,650,000	12,660,879	9,307,401	21,968,279	6.02	25,618,279	1.90	0.49	2.61	0.55	35.35
12	687,886	1,280,633	1,968,519	2,704,145	1,365,556	4,069,701	2.07	6,038,220	2.07	0.49	2.53	0.60	40.73
Total	4,611,583	31,069,635	35,681,219	164,786,875	134,567,602	299,354,477	8.39	335,035,696	2.28	0.53	3.13	0.50	36.91

NovaCopper Inc.	16-12	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

Figure 16. 6 Production Schedule

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

Figure 16.7 Pre-production Year Mine Status Map

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

Figure 16.8 Year 4 Mine Status Map

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

Figure 16.9 Year 8 Mine Status Map

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

Figure 16.10 Year 12 Mine Status Map

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
District, Northwest Alaska

16.5 MINE WASTE ROCK MANAGEMENT

Over the LOM, the pit produces 134.6 Mt of NAG waste rock and 164.8 Mt of PAG waste rock to a total of 299.4 Mt of waste rock. A total of 67.6 Mt of the NAG material will be utilized in building and raising the tailing dam and the remaining 67 Mt of NAG material will be disposed in the waste dump located at the west of the pit. The 164.8 Mt PAG material will be hauled and buried in the tailings facility.

16.6	MINING EQUIPMENT
16.6.1	MINE EQUIPMENT FLEET
	Mining equipment selection was based on the material movement schedule, geometry of the open pit, and integration with the rest of the fleet.
16.6.2	OPERATING HOURS

Mining will operate 365 d/a with 2 shifts per day and 12 hours per shift for the duration of the Arctic Project. The anticipated delays per shift are shown in Table 16.7. Based on these delays, the overall availability of mining is 18.3 hours per day.

Table 16.7 Operational Delays per Shift

Delay	Time (min)
Weather	34
Breaks	60
Shift Change	15
Blasting	30
Communication	2
Training	1
Fuel, Equipment Moves, Other	30
Total	172

16.6.3 PRIMARY EQUIPMENT

Primary equipment was selected to match scheduled mine operations. Four 11.0 m³ diesel hydraulic shovels and twenty-four 90 t trucks were selected as the primary loading and haulage equipment. Haul truck cycle times were estimated using Caterpillar Fleet Production and Cost software, estimated travel times can be seen in Table 16.8.

Blasthole drilling will be performed using up to seven diesel rotary drills as primary drilling equipment. Secondary drilling will be performed using a percussion drill. Blasting will be performed using ammonium nitrate/fuel oil (ANFO) and emulsion with mix proportions of 0.85 and 0.15, respectively.

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Preliminary Economic Assessment Report on the Arctic
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Primary equipment requirements for the LOM are listed in Table 16.9.

Table 16.8 Haulage Cycle Times

Cycle Time (min)
Destination	Year -1	Year 1 to 4	Year 5 to 8	Year 9 to 12
Crusher	15.18	12.47	11.64	13.09
NAG Dump	22.23	21.89	23.82	25.75
Dam	32.46	28.35	24.20	28.33
PAG Dump	30.92	23.80	17.02	10.46

Table 16.9 Primary Equipment Requirements

	Year
Equipment	-1	1	2	3	4	5	6	7	8	9	10	11	12
Hydraulic Shovels 14.4 yd3	2	4	4	4	4	4	3	3	3	3	3	3	1
Haul Truck (100 ton)	16	24	24	24	24	20	16	16	16	14	14	14	3
Articulated Truck (43.5 ton)	2	4	4	4	4	3	2	2	2	2	2	2	1
Wheel Loader 15 yd3	1	1	1	1	1	1	1	1	1	1	1	1	1
Wheel Loader 9 yd3	1	1	1	1	1	1	1	1	1	1	1	1	1
Service Wheel Loader	1	1	1	1	1	1	1	1	1	1	1	1	1
Diesel Drill (4.5 to 8.5 in)	4	7	7	7	7	7	6	6	6	6	6	6	2
Secondary Drill (4.5 to 5.5 in)	1	1	1	1	1	1	1	1	1	1	1	1	1

16.6.4 SUPPORT AND ANCILLARY EQUIPMENT

Four 433 kW dozers will be required for waste dump management along with two 372 kW wheel dozers for shovel cleanup. Two 16 ft (4.9 m) graders will be used to maintain roads during the pre-stripping year, increasing to three graders during the first year of production and decreasing to one grader in the last year of production. Table 16.10 and Table 16.11 list support and ancillary equipment requirements for the duration of the mine.

Table 16.10 Support Equipment Requirements

		Year
Equipment	-1	1	2	3	4	5	6	7	8	9	10	11	12
Track Dozer (433 kW)	4	4	4	4	4	4	4	4	4	4	4	4	4
Wheel Dozer (372 kW)	1	2	2	2	2	2	2	2	2	2	2	2	1
Grader (16 ft)	2	3	3	3	3	3	3	3	3	3	3	3	1
Water Truck (10,000 gal)	1	1	1	1	1	1	1	1	1	1	1	1	1
Vibratory Compactor	1	1	1	1	1	1	1	1	1	1	1	1	1
Integrated Tool Carrier	1	1	1	1	1	1	1	1	1	1	1	1	1
Excavator	3	3	3	3	3	3	3	3	3	3	3	3	3

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Preliminary Economic Assessment Report on the Arctic
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Table 16.11 Ancillary Equipment Requirements

	Year
Equipment	- 1	1	2	3	4	5	6	7	8	9	10	11	12
Flatbed Truck	1	1	1	1	1	1	1	1	1	1	1	1	1
Fuel/Lube Truck	2	2	2	2	2	2	2	2	2	2	2	2	2
Mechanics Service Truck	1	1	1	1	1	1	1	1	1	1	1	1	1
Welder Truck	1	1	1	1	1	1	1	1	1	1	1	1	1
Tire Service Truck	1	1	1	1	1	1	1	1	1	1	1	1	1
Snow Plow/Sanding Truck	1	1	1	1	1	1	1	1	1	1	1	1	1
Pick-up Truck	8	15	15	15	15	15	15	15	15	15	15	15	6
Truck Mounted Hydraulic Crane	1	1	1	1	1	1	1	1	1	1	1	1	1
Rough Terrain Forklift	1	1	1	1	1	1	1	1	1	1	1	1	1
Shop Forklift	1	1	1	1	1	1	1	1	1	1	1	1	1
Light Plant/Towers	8	8	8	8	8	8	8	8	8	8	8	8	8
Mobile Radios	50	100	100	100	100	100	81	81	81	81	81	81	30
Safety Equipment	1	1	1	1	1	1	1	1	1	1	1	1	1
Engineering/Geology Equipment (computers, software, licenses)	1	1	1	1	1	1	1	1	1	1	1	1	1
Maintenance Management System	1	1	1	1	1	1	1	1	1	1	1	1	1
Explosives Plant (facilities and buildings)	1	1	1	1	1	1	1	1	1	1	1	1	1
Surveying	1	1	1	1	1	1	1	1	1	1	1	1	1

16.7 MINING LABOUR

Mining labour requirements were estimated based on a 12-hour long shift, 2 shifts per day and a 2-week-on/2-weeks-off rotation schedule.

Operator and maintenance staff requirements are estimated based on the scheduled hours. Operational management labour and staff numbers were estimated from experience with existing mines and anticipated operating conditions for the Arctic Project.

Staff and hourly operating rates are based on the base rates and burdens. A benefit package of 40% was applied to both salaried staff and the hourly labour base rates. The labour burden consists of vacation, statutory holidays, medical and health insurance, employment insurance, long-term disability insurance, overtime, shift differential and other factors.

The LOM maximum number of mine management staff on payroll is 67. The mining operator and maintenance labour on payroll is shown in Table 16.12.

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Preliminary Economic Assessment Report on the Arctic
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Table 16.12 Operator and Maintenance Staff on Payroll

	Year
Staff	-1	1	2	3	4	5	6	7	8	9	10	11	12
Operators	130	200	199	197	197	180	154	154	153	147	148	148	58
Maintenance	62	97	97	96	96	90	79	79	78	76	76	76	32
Total	192	297	296	293	293	270	233	233	231	223	224	224	90

NovaCopper Inc.	16-21	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

17.0 RECOVERYMETHODS

17.1 MINERAL PROCESSING

A 10,000 t/d process plant has been designed to process the massive and semi-massive sulphide mineralization of the Arctic Property. The main economic elements found in the deposit are copper, zinc, lead, and associated gold and silver. The process plant will operate two shifts per day, 365 d/a with an overall plant availability of 92%. The process plant will produce three concentrates: 1) copper concentrate, 2) zinc concentrate, and 3) lead concentrate. Gold and silver are expected to be payable at a smelter and are recovered in both the copper and lead concentrates. The process plant feed will be supplied from the Arctic open pit mine.

The mill feed will be hauled from the open pit to a primary crushing facility where the material will be crushed by a jaw crusher to a particle size of 80% passing 125 mm.

The crushed material will be ground by two stages of grinding, consisting of one SAG mill and one ball mill in closed circuit with hydrocyclones (SAB circuit). The hydrocyclone overflow with a grind size of approximately 80% passing 70 µm will first undergo pre-talc flotation, and then be processed by conventional bulk flotation (to recover copper, lead, and associated gold and silver), followed by zinc flotation. The rougher bulk concentrate will be cleaned and followed by copper and lead separation to produce a lead concentrate and a copper concentrate. The final tailings from the zinc flotation circuit will be pumped to the TSF. Copper, lead, and zinc concentrates will be thickened and pressure-filtered before being transported by truck to a port and shipped to smelters.

The average annual dry concentrate production is estimated as follows:

• copper concentrate: 203,400 t/a

• lead concentrate: 23,400 t/a

• zinc concentrate: 144,100 t/a.

17.1.1 FLOWSHEET DEVELOPMENT

The process flowsheet is based on the 2012 SGS laboratory test work results conducted on the sample materials from the Arctic Deposit.

The process plant will consist of the following unit operations:

•	crushing:
	- primary crushing by a jaw crusher

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Preliminary Economic Assessment Report on the Arctic
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- associated conveying and dust suppression systems

• grinding and flotation:

- primary grinding by a SAG mill and secondary grinding by a ball mill

- talc pre-flotation

- bulk copper and lead rougher/cleaner flotation

- copper and lead separation flotation

- zinc rougher/cleaner flotation

- concentrate dewatering and load out, including water treatment for the water generated from copper and lead concentrate thickeners

- tailings disposal to the TSF.

The simplified process flowsheet is illustrated in Figure 17.1.

NovaCopper Inc.	17-2	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
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Figure 17.1 Simplified Process Flow Diagram

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
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MAJOR DESIGN CRITERIA

The process plant is designed to process 10,000 t/d, equivalent to 3,650,000 t/a. The major design criteria developed for the Arctic Project are outlined in Table 17.1.

Table 17.1 Major Design Criteria

Design Criteria	Unit	Value
Daily Processing Rate	t/d	10,000
Operating Days per Year	d/a	365
Operation Schedule	-	two shifts/day; 12 hours/shift
Crushing Availability	%	65
Grinding/Flotation Availability	%	92
Abrasion Index	g	0.046
Bond Ball Mill Work Index	kWh/t	11.1
Crushing
Nominal Processing Rate	t/h	641.0
Crusher Feed Particle Size	mm	less than 1,000
Primary Crushing Product Particle	mm	125
Size, 80% Passing
Grinding
Nominal Processing Rate	t/h	452.9
Primary Grind Size, 80% Passing	µm	800-900
Secondary Grind Size, 80% Passing	µm	70
Grinding Recirculating Load	%	300
Flotation
Flowsheet	-	Conventional flotation, including: i) talc pre-flotation, ii) copper-lead bulk flotation, iii) zinc flotation, and iv) copper/lead separation flotation

OPERATING SCHEDULE AND AVAILABILITY

The process plant is designed to operate on the basis of two shifts per day, 365 d/a. The crushing circuit overall availability will be 65%. The grinding and flotation availability will be 92%. These design availabilities will allow for scheduled and unscheduled maintenance as well as potential weather interruptions.

17.1.2 PROCESS PLANT DESCRIPTION

CRUSHING PLANT

Run-of-mine (ROM) material will be trucked to a 200 t receiving dump bin equipped with a stationary grizzly screen with an opening of 1,000 mm. The undersize of the screen will be fed to a primary jaw crusher driven by a 250 kW motor. The nominal crushing rate designed is 641 t/h. A rock breaker will be provided to break oversize rocks retained on the screen.

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Preliminary Economic Assessment Report on the Arctic
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ROM material will be crushed by the jaw crusher to 80% passing 125 mm. The crusher product will be conveyed to a 10,000 t live capacity stockpile.

A dust collection system will be provided to control fugitive dust generated during crushing and transport of the materials. A magnet will be provided over the crusher discharge conveyor to remove any steel pieces, which will protect the downstream conveyor.

The main equipment in the crushing area will include:

• one 1,000 mm stationary grizzly

• one hydraulic rock breaker

• one primary grizzly feeder, 1,600 mm wide by 6,100 mm long

• one jaw crusher, 1,200 mm by 1,600 mm; driven by a 250 kW motor

• one belt scale

• one belt magnet

• one jaw crusher discharge conveyor

• one stockpile feed conveyor

• one baghouse/dust collector.

17.1.3 COARSE MATERIAL STORAGE

The coarse material stockpile will have a live capacity of 10,000 t which is equal to 1 day of mill feed at the nominal mill feed rate, or a dead capacity of approximately 40,000 t. The coarse material stockpile will be a covered facility to effectively control the dust spreading and to mitigate freezing of the stockpiled material. The coarse material stockpile will be equipped with sufficient building access to allow for the operation of mobile equipment to work the pile as required. The stockpiled material will be reclaimed from the stockpile by three 900 mm wide by 5.1 m long apron feeders at a nominal rate of 226 t/h per feeder. Reclaimed material from the apron feeders will be discharged onto one 900 mm wide by 135 m long SAG mill feed conveyor.

The stockpile reclaim area will also be equipped with a dust collection system to minimize the spread of dust generated during transport of the crushed material.

17.1.4 GRINDING AND CLASSIFICATION

PRIMARY GRINDING AND CLASSIFICATION

The primary grinding circuit will consist of one 7.3 m diameter by 3.8 m long SAG mill equipped with a trommel screen. The SAG mill will be driven by one 3,000 kW motor. Crushed material will be conveyed from the stockpile to the SAG mill feed chute and then to the SAG mill. The SAG mill will be equipped with 50 mm pebble ports for the removal of undersize material. Mill discharge from the SAG mill will be screened by the SAG mill trommel. The trommel undersize material will flow by gravity to a sump where the ball mill discharge will report to as well. The blended slurry will be pumped to the hydrocyclone cluster. Oversized material from the trommel will be sent to a surge bin and then trucked to the TSF for disposal. As required, steel balls will be added into the SAG mill to maintain grinding efficiency. The estimated primary grind size is 80% passing 800 to 900 µm.

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Preliminary Economic Assessment Report on the Arctic
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SECONDARY GRINDING AND CLASSIFICATION

The secondary grinding circuit will include one ball mill in closed circuit with one hydrocyclone cluster. The SAG mill product will be pumped to the hydrocyclone cluster along with the ball mill discharge. The hydrocyclone underflow will gravity-flow to the ball mill while the hydrocyclone overflow, with a solid density of 33% w/w, will be sent downstream to the copper and lead bulk flotation circuit. The proposed circulation load for the closed circuit is 300%.

The grinding circuit will further grind the SAG mill product to 80% passing 70 µm. The major equipment in the secondary grinding circuit will include:

•	one 5.5 m diameter by 8.8 m effective grinding length (EGL) ball mill, driven by
	one 4,500 kW motor
•	one hydrocyclone cluster, consisting of ten 660 mm diameter hydrocyclones
	(8 operating, 2 in standby).

As required, steel balls will be added into the ball mill to maintain grinding efficiency. Zinc cyanide, a sulphide mineral suppressing reagent, will be added to the ball mill discharge pump box.

17.1.5 FLOTATION

The hydrocyclone overflow slurry will be processed by conventional flotation to recover the target minerals. Flotation will consist of talc pre-flotation, copper-lead bulk flotation to produce a bulk copper-lead concentrate, and subsequent copper and lead separation flotation to produce a copper concentrate and a lead concentrate. The copper-lead bulk flotation tailings will be conditioned and floated to produce a zinc concentrate. The zinc flotation tailings, which are the final tailings product, will be pumped to the TSF for storage.

TALC FLOTATION

The hydrocyclone overflow with 33% solids will be floated in two 100 m³ tank-type flotation cells to remove talc, which may impact copper and lead flotation and concentrate quality. Depending on the copper, lead, and zinc content of the talc product, the talc rougher flotation concentrate may be further cleaned to reject any entrained sulphide minerals. The pre-flotation tailings will be directed to the copper-lead bulk flotation circuit, while the talc concentrate will be sent to the final tailings pump box to be pumped to the TSF, together with the zinc rougher scavenger flotation tailings and the first zinc cleaner scavenger flotation tailings.

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Preliminary Economic Assessment Report on the Arctic
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COPPER-LEAD BULK FLOTATION AND REGRINDING

The talc pre-flotation tailings will flow into the copper-lead bulk flotation conditioning tank where copper-lead mineral collectors will be added. The conditioned slurry will be floated in six conventional 100-m³ tank-type flotation cells, five cells for rougher flotation and one for rougher/scavenger flotation. The concentrate from the rougher bulk flotation circuits will be reground in a tower mill, which is driven by one 750 kW motor. The mill will be in closed circuit with ten 250 mm diameter hydrocyclones to grind the bulk concentrate to a particle size of 80% passing 35 to 40 µm prior to being further upgraded by two stages of cleaner flotation.

The first stage of copper-lead bulk cleaner flotation will be conducted in one 4.0 m diameter by 11.0 m high flotation column. The first cleaner flotation tailings will be scavenged to further recover copper-lead minerals in two 20 m³ tank-type flotation cells. The concentrate from the first cleaner/scavenger flotation will flow to the regrinding pump box for further regrinding, while the tailings produced from the flotation stage will be sent to the copper-lead rougher/scavenger tailings pump box where the combined slurry will be pumped to the zinc flotation circuit.

The copper-lead concentrate from the first cleaner flotation stage will be pumped to the second cleaner flotation stage. Secondary cleaner flotation will be conducted in one 4.0 m diameter by 10.0 m high flotation column. The concentrate from the copper-lead second cleaner flotation will be pumped to the copper-lead bulk concentrate thickener prior to being pumped to the copper and lead separation flotation circuit. The overflow of the thickener will be sent to the process water tank for reuse. The tailings from the copper-lead secondary cleaner flotation will return to the copper-lead first cleaner flotation.

Copper-lead rougher flotation and cleaner flotation will be carried out at a pH of 8.0 to 8.5. Reagents used in the circuit include:

• zinc cyanide as depressant

• SIPX and 3418A as collectors

• MIBC as frother.

The major process equipment used in the copper-lead bulk flotation circuit includes:

• one 4.0 m diameter by 4.5 m high conditioning tank

• six 100 m³rougher and scavenger flotation cells

• one tower mill with an installed power of 750 kW

• ten 250 mm hydrocyclones (eight operating, two standby)

• one 4.0 m diameter by 11.0 m high flotation column for first cleaner flotation

• two 20 m³tank flotation cells for first cleaner/scavenger flotation

• one 4.0 m diameter by 10.0 m high flotation column for second cleaner flotation

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Preliminary Economic Assessment Report on the Arctic
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• one 10.0 m diameter high rate thickener for the cleaned bulk concentrate.

Copper-Lead Separation Flotation Circuit

The underflow of the copper-lead bulk cleaner concentrate thickener will be pumped to the copper-lead separation flotation conditioning tanks, where sodium cyanide and lime will be added to suppress copper minerals. The conditioned slurry will flow to lead rougher flotation cells. Lead collector will be added in the second conditioning tank to collect lead minerals. The lead concentrate produced from the lead rougher and rougher/scavenger flotation cells will be pumped to the first lead cleaner flotation circuit. The tailings of the rougher/scavenger flotation, which is the copper concentrate, will report to the copper concentrate thickener.

The lead concentrates from the lead rougher and scavenger flotation circuit will be further upgraded by three stages of cleaner flotation. The first stage of cleaner flotation will consist of first lead cleaner flotation and cleaner scavenger flotation. The lead concentrates from the first stage of cleaner flotation will be fed to the second stage of lead cleaner flotation. The lead cleaner/scavenger tailings, together with the lead rougher/scavenger concentrate, will report to lead rougher flotation. The second lead cleaner flotation concentrate will be further upgraded by the third stage of cleaner flotation to produce the final lead concentrate, which will report to the lead concentrate thickener. Tailings from the second and third stages of cleaner flotation will be forwarded to the proceeding flotation cell feed boxes.

Lead flotation will be conducted at a pH ranging from 9.0 to 9.5. Reagents used in the circuit include:

• lime for adjusting slurry pH

• sodium cyanide as a copper mineral depressant

• 3418A as collector

• MIBC as frother.

The copper- lead flotation circuit consists of:

• two 2.8 m diameter by 3.0 m high conditioning tanks

• five 20 m³lead rougher and rougher/scavenger flotation cells

• three 5 m³lead first cleaner and scavenger flotation cells

• one 5 m³lead second cleaner flotation cell

• one 5 m³lead third cleaner flotation cell.

Zinc Flotation and Regrinding Circuit

The copper-lead rougher scavenger tailings will be further floated to recover zinc in the zinc flotation circuit, along with the copper-lead first bulk cleaner scavenger tailings. The circuit will consist of feed slurry conditioning, rougher/scavenger flotation, zinc rougher concentrate regrinding, and two stages of cleaner flotation.

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Preliminary Economic Assessment Report on the Arctic
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Tailings from the copper-lead bulk flotation circuit will be sent to the head cell of a bank of zinc rougher and rougher scavenger flotation cells, after being conditioned with lime (to a pH above 10.5, to depress pyrite) and copper sulphate (to activate zinc minerals). The flotation circuit will produce a zinc rougher and scavenger concentrate and zinc rougher scavenger flotation tailings. The concentrates produced will report to the regrinding circuit. Tailings will flow to the final tailings pump box to be delivered to the TSF.

Concentrates from the rougher and scavenger flotation stage will flow to the regrinding pump box, from where the slurry will be pumped to the regrinding hydrocyclone cluster. With a particle size of 80% passing 35 to 40 µm, the overflow of the hydrocyclones will report to the zinc first cleaner and scavenger flotation circuit, while the underflow will gravity flow to a tower mill driven by a 750 kW motor. The slurry discharged from the mill will flow to the hydrocyclone feed pump box.

There will be two stages of zinc cleaner flotation. The zinc cleaner circuit will operate at a pH of 11 or above to reject pyrite. The first cleaner concentrate produced will be further upgraded in the second stage of cleaner flotation. The first cleaner flotation tailings will be further floated in cleaner scavenger flotation cells. Concentrate from the first stage of cleaner scavenger flotation will flow to the zinc regrinding pump box. Tailings from the first stage of cleaner scavenger flotation will flow to the final tailings pump box where tailings will be pumped to the TSF, along with the zinc rougher scavenger flotation tailings and the talc concentrate.

The final zinc concentrate produced from the second stage of cleaner flotation will report to the zinc concentrate thickener, while the second cleaner flotation tailings will be recycled back to the first cleaner flotation column feed pump box

Reagents used in the zinc flotation circuit include copper sulphate, lime, SIPX, and MIBC.

The main equipment used for the zinc flotation circuit consists of:

• two 4.0 m diameter by 4.5 m high conditioning tanks

• six 100 m³zinc rougher and rougher scavenger flotation cells

• one 750 kW tower mill

• eight 250 mm hydrocyclones (6 operating, 2 standby)

• one 4.0 m diameter by 10.0 m high flotation column for the first zinc cleaner flotation

• two 20 m³tank cells for the first zinc cleaner scavenger flotation

• one 3.0 m diameter by 8.0 m high flotation column for the second zinc cleaner flotation.

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Preliminary Economic Assessment Report on the Arctic
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PRODUCT DEWATERING

Each of the flotation concentrates will be thickened in its respective thickener, including copper, lead, and zinc concentrates. The concentrates will be further dewatered by pressure filters to a moisture content of approximately 9%. Pressure filtration with an air blowing function has been selected for the Arctic Project in order to achieve a low overall cake moisture.

Copper Concentrate Dewatering

Copper concentrate from the tailings of the lead rougher scavenger flotation will be pumped to a 10 m-diameter high-rate thickener. The copper concentrate will be mixed with diluted flocculant solution at the thickener feed well. The thickener will be operated together with variable frequency drive underflow slurry pumps to condense the underflow to approximately 65% solids. Prior to pressure filtration, the underflow will be pumped to an agitated concentrate stock tank, which is capable of storing the concentrate for more than eight hours.

Thickener overflow will be treated by a sulphur dioxide-air procedure to destroy residual cyanide and reduce dissolved heavy metals prior to being sent to the final flotation tailings pump box from where the slurry is delivered to the TSF.

The final filter cake moisture is expected to be 9%. The copper concentrate will be discharged into a stock bin, from which the concentrate will be automatically loaded onto haulage trucks for shipment. Filtrate from the filtration will return to the copper concentrate thickener.

The in-plant storage and dispatch facility will be capable of storing up to 10 days of copper concentrate production in case of potential truck haulage interruptions caused by inclement weather or other.

The equipment required for concentrate thickening and filtration includes:

• one 10 m diameter high rate thickener

• one 7.5 m diameter by 8.5 m high agitated concentrate stock tank

• two 100 m²pressure filters.

Lead Concentrate Dewatering

The lead concentrate will be blended with the flocculant solution and forwarded to a 3.5 m diameter high-rate thickener. Thickener underflow containing approximately 65% solids will be pumped to an agitated concentrate stock tank prior to pressure filtration. The concentrate stock tank will have the capacity to contain lead concentrate production for at least eight hours. The thickener overflow will be directed to the conditioning tank in the copper and lead separation circuit for reuse or sent to the water treatment circuit that is designed for treating the copper thickener overflow.

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Preliminary Economic Assessment Report on the Arctic
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The filtrated lead concentrate cake with 9% moisture will be bagged and then loaded into a 20 ft concentrate container, which will be hauled to the ocean port for shipment. The filtrate from the filter will return to the lead concentrate thickener feed well.

The in-plant lead concentrate storage and dispatch facility will be capable of storing at least 10 days of lead concentrate production.

The equipment required for concentrate thickening and filtration includes:

• 3.5 m diameter high rate thickener

• 4.0 m diameter by 4.0 m high agitated concentrate stock tank

• one 40 m²pressure filter.

Zinc Concentrate Dewatering

The final zinc concentrate will be directed to an 8.0 mm diameter high-rate thickener. Flocculant will be added to improve settling of the concentrate. Thickener underflow with a solid density of 65% will be stored in an agitated concentrate stock tank prior to pressure filtration. Thickener overflow will be delivered to the effluent treatment plant or to the final flotation tailings pump box. The concentrate stock tank is designed to be capable of containing zinc concentrate production for at least eight hours.

The final filter cake moisture is expected to be 9%. Filtrate from the filtration will return to the zinc concentrate thickener. The zinc concentrate will be conveyed and discharged to a concentrate stock bin, and then automatically loaded onto trucks for shipment off site to smelters.

The in-plant zinc concentrate storage and dispatch facility will be capable of storing at least 10 days of zinc concentrate production in case of potential interruptions to truck haulage due to inclement weather.

Equipment required for concentrate thickening and filtration includes:

• one 8.0 mm diameter high -rate thickener

• one 7.5 m diameter by 8.0 m high agitated concentrate stock tank

• two 70 m²pressure filters.

TAILINGS DISPOSAL

The final flotation tailings will flow by gravity to the TSF, and includes:

• the talc flotation concentrate

• the zinc rougher scavenger tailings

• the zinc first cleaner scavenger flotation tailings.

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Preliminary Economic Assessment Report on the Arctic
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The tailings pond will be installed with a reclaim water pump barge, which will reclaim water from the pond to either the process water tank or the water treatment plant.

Tailings management is discussed further in Section 18.0.

REAGENT HANDLING AND STORAGE

Various chemical reagents will be added to the grinding and flotation circuits to modify the mineral particle surfaces and enhance the floatability of the valuable mineral particles into the concentrate products.

The reagents will be prepared and stored in a separate, self-contained area within the process plant and delivered by individual metering pumps or centrifugal pumps to the required addition points. All reagents will be prepared using fresh water.

Preparation of the various reagents will require:

• a bulk handling system

• mixing and holding tanks

• metering pumps.

Collectors

The collector SIPX in a solid form will be shipped to the mine site in bags. The SIPX will be diluted to 20% solution strength in a mixing tank and stored in a holding tank, before being added to the copper-lead bulk flotation circuit and the zinc flotation circuit via metering pumps.

The collector 3418A will be received as a liquid in drum. This collector will be delivered to the lead flotation and the copper-lead rougher flotation circuit via metering pumps without dilution.

Frother

MIBC frother will be received as a liquid in 500 kg drums. The reagent will be used at the supplied solution strength. Metering pumps will deliver the frother to the talc, copper, lead, and zinc flotation circuits.

Lime

Lime will be trucked to the site as unslaked pebbles, and stored in a 100 t-capacity bulk lime silo. Lime will be conveyed to the lime slaker and slaked with water. The slaked lime will be stored in an agitated mixing tank and distributed to the addition points via a lime slurry loop, mostly to the zinc flotation circuits. The slaker will be controlled automatically based on the lime slurry levels in the holding tank.

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Preliminary Economic Assessment Report on the Arctic
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Flocculant

Solid flocculant will be prepared in a packaged preparation system, including a screw feeder, a flocculant educator, and mixing devices. Flocculant will be diluted to a 0.2% solution strength and added via metering pumps to the copper, lead, and zinc thickeners’ feed wells where the reagent will be further diluted.

Other Reagents

Sodium cyanide, zinc sulphate, and copper sulphate will be supplied in powder/solid form, and will be dissolved and diluted by fresh water. The strength of the reagent solutions will be approximately 20%. Cyanide monitoring/alarm systems will be installed at the cyanide preparation areas. Emergency medical stations and emergency cyanide detoxification chemicals will be provided at the areas as well.

Zinc cyanide will be prepared by mixing the sodium cyanide solution, the zinc sulphate solution, and the lime slurry.

Anti-scale chemicals may be required to minimize scale build-up in the reclaim or recycle water lines. These chemicals will be delivered in liquid form and metered directly into the intake of the reclaim water pumps or process water tank.

A separate reagent system will be provided to occasionally determine the effect of any new reagent on metal recovery and concentrate grading.

Storage tanks will be equipped with level indicators and instrumentation to ensure that spills do not occur during normal operation. Appropriate ventilation, fire and safety protection equipment and devices will be provided at reagent preparation areas.

WATER SUPPLY

There will be two separate water supply systems: a fresh water supply system and a process water supply system. There will also be an effluent treatment plant to treat effluents generated from process and mining operations.

Fresh Water Supply System

Process fresh water will be supplied from the water treatment plant or local creeks. A 10,000 mm by 8,500 mm fresh water/fire water storage tank will hold operating fresh water prior to distribution within the plant. Fresh water will be mainly used for the following purposes:

• fire water

• gland and seal water

• mill lubrication cooling water.

Potable water will be supplied from local creeks and chlorinated prior to delivery to the potable water distribution loop.

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Preliminary Economic Assessment Report on the Arctic
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Process Water

Process water will be made up of fresh water, and water reclaimed from the TSF. Reclaimed water and fresh water will be directed to an 8.5 m diameter by 10.0 m high process water tank, from where the water will be distributed to the process plant and other service locations.

AIR SUPPLY

Air service systems will supply air to the grinding and flotation plant, as follows:

•	Flotation:
•	-	Low-pressure air will be provided for flotation tank cells by air blowers. High- pressure air will be provided for the flotation column by dedicated air compressors.
	Filtration Circuit:
	-	High-pressure air will be provided for filtration and drying by dedicated air compressors.
•	Plant Air Service:
	-	High-pressure air will be provided for various services by dedicated air compressors.
•	Instrumentation Air:
	-	Air will come from the plant air compressors; it will be dried and stored in a dedicated air receiver.

A separate high-pressure air service system will supply air to the crushing plant by a dedicated air compressor. The air will be provided for dust suppression and equipment services.

ASSAY/METALLURGICAL LABORATORY AND QUALITY CONTROL

The final concentrates and intermediate streams will be monitored by two online x-ray diffraction analyzers. The analyzers will be equipped with pH monitoring systems. The assay data will be fed back to the central control room and used to optimize the process conditions. Routine samples of head, intermediate products, tailings, and final products will be collected and assayed in the assay laboratory, where standard assays will be performed. The data obtained will be used for product quality control and routine process optimization.

The assay laboratory will consist of a full set of assay instruments for base metal analysis as well as gold and silver assays, including:

• fire assay equipment

• an AAS

• an ICP for the routine assay laboratory

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Preliminary Economic Assessment Report on the Arctic
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• an inductively coupled plasma-mass spectrometry (ICP-MS) for the environmental laboratory

• a Leco furnace

• other determination instruments such as pH and redox potential metres and experimental balances.

The metallurgical laboratory will perform metallurgical tests for quality control and process flowsheet optimization. The laboratory will be equipped with laboratory crushers, ball mills, particle size analysis devices, laboratory flotation cells, balances, and pH metres.

17.2	PLANT PROCESS CONTROL
17.2.1	OVERVIEW

PLANT CONTROL

The plant control system will consist of a distributed control system (DCS) with personal computer-based operator interface stations (OIS) located in the control rooms of the following areas:

• Primary Crushing: A control room will be provided in the primary crushing area with a single OIS. Control and monitoring of all primary crushing and conveying operations will be conducted from this location.

• Mill Building: A central control room will be provided in the mill building with required an OIS.

In conjunction with the OIS, the DCS will perform all equipment and process interlocking, control, alarming, trending, event logging, and report generation.

The plant control rooms will be staffed by trained personnel 24 h/d.

Programmable logic controllers (PLCs) (or other third party control systems supplied as part of mechanical packages) will interface with the plant control system via ethernet network communication systems when possible.

Operator workstations will be capable of monitoring the entire plant site process operations, and will be capable of viewing alarms and controlling equipment within the plant. Field instruments will be microprocessor-based “smart” type devices. Instruments will be grouped by process area, and wired to each respective area’s local field instrument junction boxes. Signal trunk cables will connect the field instrument junction boxes to DCS input/output (I/O) cabinets.

Intelligent-type motor control centres (MCCs) will be located in electrical rooms throughout the plant. Utilizing an industrial communication protocol interfaced to the DCS, a serial bus network will facilitate remote operation and monitoring of the MCC.

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An automatic sampling system will collect samples from various product streams for online analysis and assessment of daily metallurgical balance.

17.2.2 PRODUCTION PROJECTION

Based on the preliminary mine plan and preliminary metallurgical test work, the projected annual concentrate production from the proposed plant is provided in Table 17.2.

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Table 17.2 Annual Concentrate Production Projection

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Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining
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18.0 PROJECTINFRASTRUCTURE

18.1 OVERVIEW

The proposed Arctic mine site is spread over a distance of approximately 6 km within the upper reaches of the Sub-Arctic Creek Valley. The proposed development for the Arctic Project consists of the following major infrastructure:

• roads and an airstrip

• mill buildings and related services facilities including maintenance and truck shops, and assay lab

• water supply and distribution

• waste management

• fuel storage

• on site explosive storage

• power supply

• TSF and water management

• water treatment plant

• construction and permanent camp accommodation

• waste rock storage facilities

• communication.

All buildings and facilities will be constructed with appropriate HVAC and fire protection systems, water distribution and plumbing systems, and dust control systems.

The location of major project facilities is shown in Figure 18.1.

Mine-site infrastructure has been located to:

• take advantage of local topography

• minimize pumping requirements from the mill building to the TSF

• minimize environmental impacts to Sub-Arctic Creek

• minimize snow avalanche mitigation requirements

• reduce the haul distance from the pit to the primary crusher and TSF.

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Preliminary Economic Assessment Report on the Arctic
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Crushed material will be conveyed, via an overland conveyor, from the primary crusher to the crushed material stockpile, from which the material will be conveyed to the SAG mill for processing.

At this PEA level review, little survey, geotechnical, hydrological, avalanche, environmental or other constraint information was provided. There was no site specific geotechnical information although permafrost is assumed to exist at the proposed mine site. Tetra Tech did not review any hydrological, avalanche, environmental or other constraint information for the proposed mine site.

No specific civil engineering design criteria were developed at this stage of the Arctic Project review. Instead, typical design criteria developed for similar climate/arctic and remote conditions were applied to the PEA design. This included internal roads and grading criteria. Assumed quantities for drainage systems were incorporated into the cost estimate as drainage channels and ditches were not designed. While an approximate retaining wall quantity was included in the cost estimate, no retaining wall designs were completed for the PEA.

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Preliminary Economic Assessment Report on the Arctic
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Figure 18.1 Final General Arrangement Layout after 12 years of Operation

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Preliminary Economic Assessment Report on the Arctic
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18.2	ROADS AND AIRSTRIP
18.2.1	AMBLER MINING DISTRICT INDUSTRIAL ACCESS ROAD

There is currently no developed surface access to the Arctic Project area. Access to the Project is proposed to be via a road approximately 340 km (211 miles) long, extending west from the Dalton Highway where it would connect with the proposed Arctic Project area. Although the capital costs of the road are not yet known, NovaCopper has been in discussions with AIDEA regarding the viability of permitting and constructing an access road. Further information regarding the AMDIAR can be found in see Section 21.5.6.

18.2.2 ACCESS ROAD

The Arctic Deposit access road branches off from the proposed AMDIAR and ends at the Arctic Deposit. Figure 18.2 shows the overall road location plan showing the relationship of the access road to the proposed Brooks East corridor which eventually connects to the existing Dalton Highway 320 km from the access road intersection.

Figure 18.2 Access to the Arctic Project

The access road is 17 km long from the AMDIAR corridor to the on site mine roads. The grades are relatively gentle in the order of 3 to 5%, with the maximum grade at less than 9%.

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The road is located in a permafrost environment. The initial 10 to 12 km is in a lowlands area that shows strong indications of being ice-rich. The recommended approach to construct a road in a permafrost environment is to protect the permafrost from degradation by avoiding cuts and providing sufficient embankment. The embankment height will be greater in the lowlands as it is ice-rich and more susceptible to degradation.

The last 5 km of the road transitions from the valley floor and rises 300 m in elevation, with the last 3 km traversing steep side hills. Where the alignment traverses steep side hill, roadway cuts may be necessary. Cuts would be recommended only if bedrock was very close to the surface.

More information on the access road is provided in Appendix B.

18.2.3 HAUL ROADS, SITE ROADS, AND PAD AREAS

The mine site is spread over a distance of approximately 6 km. A network of site roads connecting the pit and the various on site facilities will be constructed throughout the site.

Major site roads will include mine haul roads from the open pit to the primary crusher receiving area, waste rock storage areas and tailings storage areas, and roads to and from the truck shop. These major roads will generally be gravel-surfaced and 30 m wide. The road sub-base and base requirements will be governed by the quality of the subgrade; overall road thickness is expected to be approximately 1 m. The maximum road grade will be 10%. Safety berms will be provided wherever needed.

Other site service roads will interconnect the following facilities and areas:

• open pit

• process plant

• coarse crushed stockpile

• main substation/power generation

• primary crushing facility

• maintenance facilities

• camp

• explosives storage

• overland conveyor right of way.

These roads will generally be 6 to 8 m wide and gravel-surfaced. The roads will be constructed with a maximum grade of 10%.

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18.2.4 AIRSTRIP

The proposed location for an airstrip sufficient to support project activities is located in the valley approximately 21 km from the Arctic Project Site (Figure 18.2). Geotechnical data is not available for the site but it is assumed that the facility will require permafrost protection to ensure year round operations.

The airstrip will operate as a private aerodrome and prior permission will be required for all aircraft utilizing the site. The design criteria for the airstrip will include:

• Design Aircraft – Dash 8 – Q400/Hercules C130

• Gravel Airstrip, surfaces will be treated for dust control.

• general dimensions:

- runway length: 1,524 m

- runway width: 45 m

- graded strip either side of runway: 7.5 m

- side slope, beyond graded strip: 25% (4:1)

- runway end safety area: 60 m both ends

- apron capable of parking two Dash 8 – Q400 aircraft

- 2 m permafrost protection.

The site will be provided with:

• medium intensity light-emitting diode (LED) airfield edge, end, taxiway lighting

• apron flood lighting

• backlit mandatory signage

• lighted wind direction indicators

• Runway End Identification Lighting System (REILS) at both approaches to the runway

• Precision Approach Path Indicators (PAPIs) at both runway touchdown points

• Localizer Performance with Vertical Guidance (LPV) Wide Area Augmentation System (WAAS)

• GPS approach and departure procedures (with vertical guidance) would be developed for both runway.

More information on the airstrip is provided in Appendix C.

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18.3	BUILDINGS AND RELEVANT BUILDING SERVICES
18.3.1	PRIMARY CRUSHING BUILDING

The primary crusher will be housed in a pre-engineered building that will also house a control room, an apron feeder, and a crusher discharge conveyor.

ROM material trucked from the pit will be discharged into a dump bin equipped with a stationary grizzly screen. The receiving bin is located outside of the crushing building. The ROM material will be transported to the crusher by an apron feeder. The crushed material will discharge onto the crusher discharge conveyor and then to the overland conveyor to the coarse material stockpile. Sufficient space will be provided inside the building for ongoing operation and maintenance.

18.3.2 CRUSHED MATERIAL STOCKPILE

The crushed material stockpile will be fed by the overland conveyor connecting with the primary crushing building. The stockpile with a live capacity of 10,000 t will be covered and equipped with a dust control system. Three apron feeders inside the concrete reclaim tunnel will reclaim the stocked material onto a SAG mill feed conveyor.

18.3.3 CONVEYING

The conveyors connecting the stockpile and the SAG mill feed chute and the primary crusher discharge conveyor will be vendor-supplied systems that will include all structural support frames, trusses, bents and take-up structures. The overland conveyors will be supported on concrete pre-cast strip panels/sleepers spaced at regular intervals. Elevated sections of conveyor systems will be supported on vendor-supplied steel trusses spanning between steel bents on concrete spread footings.

18.3.4 MILL BUILDING, MAINTENANCE, TRUCK SHOP, ASSAY AND METALLURGY LAB

The mill building will be an engineered post-and-beam steel structure with insulated steel roof and walls. The building foundation will comprise concrete spread footings, grade walls along the building perimeters, and slab-on-grade floor. The floor surfaces will have localized areas sloped toward sumps for cleanup activities.

Interior steel platforms on multiple levels will be provided for ongoing operation and mill facility equipment maintenance. The building will house milling, flotation, reagent preparation/delivering, concentrate dewatering, concentrate storage, concentrate load out, and some offices for mill staff. The building will also house the truck shop and maintenance facilities – (including a wash bay and three repair bays), warehouse, first aid room and emergency vehicle parking, and the assay and metallurgical laboratories. The truck shop will have sumps and trenches constructed to collect waste water in the maintenance bays. Floor hardener will be applied to concrete surface for high traffic areas. The warehouse and repair bays will be serviced by two 10 t overhead cranes.

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All water supply tanks for fresh water and process water supplies will be placed inside the mill building, except diesel fuel storage tanks. The process areas will be serviced by three cranes of various capacities that match specific functions.

The assay and metallurgical laboratories will house all necessary laboratory equipment for sample assay, product quality control and process optimization. The laboratories will be equipped with all appropriate HVAC and chemical disposal equipment as needed. The facility floor will be reinforced as needed to accommodate specialized equipment.

18.3.5 ADMINISTRATION AND MINE DRY

The administration and mine dry complex will be a part of the camp complex. This facility will house a mine dry and lockers to service on site mining, process, surface services, and administration personnel and office areas for staff and supervisors. A mud room will also be attached.

18.3.6 ARCTIC CORRIDORS

Heated arctic corridors will be provided for worker access to main buildings, and to house power cables, heating system piping, and other pipe lines. The corridors will connect the camp, main process plant, and diesel power plant.

Two boilers will provide emergency heat supply for the camp and the main plant building when power generators are not able to produce sufficient heat.

18.3.7 COLD STORAGE WAREHOUSE

The cold storage warehouse will be a pre-engineered galvanized steel structure with an un-insulated fabric cover. The building will be supported on pre-cast concrete lock-blocks on a prepared gravel surface.

18.3.8 HVAC AND FIRE PROTECTION

All process areas will be heated to a minimum temperature of 5°C on a typical winter day, by recovered waste heat from diesel power generators.

Large process buildings will be ventilated year-round to prevent a build-up of contaminants and humidity. The ventilation rate will rise during summer months. During winter months, heat will be supplied by waste heat from the diesel power plant via glycol pumping.

All staff-occupied areas, such as offices, washrooms, and change rooms, will be heated to a minimum temperature of 20°C on a typical winter day, by supplying filtered and tempered outdoor air mixed with return air. The air will be distributed through ductwork into the individual rooms.

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Air conditioning will be limited to control rooms, laboratories, and electrical rooms in which heat generated from electrical equipment is expected to be excessive. Electrical rooms in which heat is not expected to be significant will be cooled with filtered outdoor air. Small rooms, electrical rooms, and remote buildings will be heated using electric heat.

Washrooms, change rooms, print rooms, janitor’s rooms, showers and lunch rooms will be mechanically exhausted to atmosphere. Make-up air will be transferred from adjacent areas, or be supplied with filtered tempered outdoor air.

18.3.9 FIRE PROTECTION

Each building will be equipped with fire water tanks capable of sustaining enough water for two hours of firefighting at a flow rate equal to the largest sprinkler flow, plus flow from an inside hose and an outside hydrant. The fire water system will be pressurized by a fire water pump package comprising a jockey pump, a main electric pump, and a standby diesel-fired pump.

18.3.10 DUST CONTROL

Dust control systems, comprising dry baghouses, will be installed at the primary crusher building and the coarse crushed material reclaim tunnel.

The dust will be pneumatically conveyed from each exhaust hood to the dust collector through steel ducting, which will include test ports, dampers and cleanouts. Collected dust fines will be returned to the process stream.

Dust will also be controlled at the truck dump, using a series of water spray nozzles, whenever the outdoor temperature is above freezing.

18.4 WATER SUPPLY AND DISTRIBUTION

The construction camp will require approximately 320 L/person/d (or 80 m³/d) of potable water to supply a peak of approximately 700 direct workers, staff, and indirect operations personnel during the construction phase and 350 direct workers, staff, and indirect operations personnel during operation of the Arctic Project. A potable water package treatment unit, including a bottling facility for drinking water, will be located close to the administration building. The potable water will undergo chlorination and ultraviolet light treatment.

Fresh water will be pumped from local groundwater wells to fresh and firewater storage tanks, located near each major facility. To ensure adequate firewater levels, two individual reservoirs inside the tank will be separated by an internal standpipe. The tank will be equipped with a level detection control system.

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The fire/fresh water tank will be located inside the process building to prevent freezing during winter months. The water pressure will be maintained by booster pumps, jockey pumps and diesel back-up pumps. A plant site alarm will signal a low system pressure condition.

Emergency showers and eyewash stations will be installed throughout the process building.

18.5	WASTE MANAGEMENT
18.5.1	SEWAGE DISPOSAL

Sewage is calculated based on a provision for 120 L/person/day of sewage to be collected, approximately 40 L/person/d of which will be treated. The proposed treatment plant will be a membrane bioreactor system and chemical phosphorus removal.

18.5.2 DOMESTIC WASTE DISPOSAL

Construction and industrial waste will be landfilled within the waste storage bin (WSB) laydown area. Non-hazardous solid construction industrial waste will be buried in the WSB laydown area. Domestic waste from the construction camp and operating areas will be incinerated with the ash buried in the WSB laydown area.

Scrap metal will be collected in bins and recycled by a qualified local contractor.

Several forms of domestic and industrial solid waste will be generated over the life cycle of the mine. All avenues of re-use, reduction and recycling of materials will be examined and implemented prior to disposal of any waste.

18.6 FUEL STORAGE

The diesel fuel storage comprising three 500,000 L tanks will be housed in a high-density polyethylene (HDPE)-lined, dyked tank farm, located near the truckshop. Fuel dispensing facilities, including fill system for light vehicles as well as fast-fill facilities for mining equipment, will be included.

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18.7	ON SITE EXPLOSIVES STORAGE
18.7.1	EMULSION PLANT

The emulsion plant will be a 50 m by 50 m facility located near the open pit, approximately 1,000 m from the nearest road or mine building. The explosive facility will include an emulsion plant and emulsion storage.

18.7.2 DETONATOR AND EXPLOSIVE STORAGE MAGAZINE

The magazine will be fabricated from a 6 mm metal plate the walls will comprise at least 7.6 cm of bullet-resistant material and the roof will be a 4.7 mm (or thicker) metal plate. The magazine will be mounted on large metal I-beam skids, providing a minimum ground clearance of 10 cm or more for structural rigidity and portability. The magazine will be fabricated in accordance with current storage standards for industrial explosives.

18.8 POWER SUPPLY TO PLANT SITE

The Arctic Project will require 15 MW of peak load for 10,000 t/d operation demand. Power will be generated by five self-contained 3.6 MW prime diesel generators. Four units will be in service with the fifth unit reserved for maintenance. Heat will be recovered from the generators and used to heat the mill, camp and related facilities.

18.9 TAILINGS STORAGE FACILITY

The co-disposal TSF will be a fully lined facility consisting of rockfill embankment constructed across the Arctic Creek drainage, creating an impoundment that will extend up the drainage. The rockfill embankment will be constructed to an ultimate crest elevation of 655 mamsl with the embankment being raised in stages to minimize the initial capital construction cost. During operations, PAG waste rock will be dumped at the bottom and sides of the basin forming layers with consecutive disposal on tailings that will be filling the voids. The tailings has the potential to generate acid and, therefore, the tailings and the PAG waste rock will be placed under water and remain permanently submerged in order to reduce the potential for acid generation. Additional studies will be required to determine the most suitable method of co-disposal and potential requirements for ARD management and mitigation programs will need to be part of the design of the TSF.

Further details regarding the TSF can be found in Appendix D.

The TSF is designed to contain a total of 110.5 Mm³ over the 12 year life of mine, with 23.8 Mm³to accommodate the tailings at an assumed stored dry density of 1.5 t/m³, and 86.7 Mm³ of PAG waste rock at an assumed stored dry density of 1.9 t/m³. The TSF will be sited as a staged rockfill embankment with an upstream geomembrane liner. The starter embankment will have a crest elevation of 560 m and impound 1 year of mining/processing production, which is approximately 670,000 m³ of tailings and 12.3 Mm³ of waste rock (Figure 18.3).

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Figure 18.3 Tailings Storage Facility

18.9.1 TSF EMBANKMENT

The TSF embankment has 2.5:1 (2.5 horizontal to 1 vertical) upstream slope and a 2:1 downstream slope. The ultimate height of the embankment will be approximately 195 m, measured from the downstream toe to the crest of the embankment. The crest of the embankment is designed at 12 m wide to accommodate vehicles and equipment. The embankment is a homogeneous rockfill structure founded on bedrock. NAG waste rock will be used to construct the embankment. A geotextile material will be placed over the upstream slope of the rockfill embankment and below the geomembrane liner, to prevent damage to the lining system from ice. The TSF embankment sections are illustrated in Figure 18.4.

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18.9.2 TAILINGS IMPOUNDMENT

It has been assumed that low permeability soils do not naturally exist within the impoundment area of the TSF and a synthetic geomembrane lining system will be required. This system will tie into the geomembrane installed on the embankment, creating a fully lined facility. Construction of the impoundment area will consist of removing the topsoil and vegetation, compacting exposed fine-grained soils and installing the geomembrane. A protective layer will be placed over the geomembrane to prevent damage to the liner.

Figure 18.4 Tailings Embankment Cross Section

18.9.3 RECLAIM WATER SYSTEM

Water liberated from the tailings will be collected in a free water pool or supernatant pond formed at the tailings beach/ground interface. This water will be pumped back to the mill for re-use. During operations, the tailings deposition from the embankment crest and southern limit of the impoundment will result in a supernatant pond located at the northern limits of the TSF.

Water from the supernatant pond will be returned to the mill for re-use in the process via a floating barge pump system. The solution will be transferred to the mill via an overland HDPE pipeline.

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18.9.4 SEEPAGE COLLECTION SUMP

The seepage collection sump will collect seepage from below the TSF embankment and impoundment for routing back to the supernatant pond. The conceptual design for the seepage collection sump is a vertical sump that is piped to the embankment toe drains, consisting of perforated pipes encapsulated in gravel, and conveyed to the sump where it will be pumped back into the TSF impoundment.

18.10 WASTE ROCK FACILITIES

The waste rock produced over the projected 12-year mine life (including one year of preproduction) is summarized in Sections 16.3 and 16.4. The waste schedule is plotted in terms of annual tonnages by waste type: PAG and PAG (Table 16.6). Of the 299.4 Mt of waste rock, 164.8 Mt is currently classified as PAG and 134.6 Mt as NAG.

18.10.1 NON-ACID GENERATING WASTE ROCK

NAG waste rock will first be used for the construction of the TSF embankment. Excess NAG waste rock will be stockpiled between the plant site and the TSF (Figure 18.3). Through the LOM, approximately 68.6 Mm³of NAG waste rock will be generated from the mine, while 32.2 Mm³will be used in the construction of the TSF embankment. The remaining 36.4 Mm³will be stored in the stockpile with an ultimate elevation of 1,000 mamsl. The stockpile was designed with a 2:1 slope and will have an ultimate height of 320 m measured from the toe to the top. The waste rock that is NAG will not be submerged upon closure, and will be utilized to construct facility infrastructures (i.e. TSF) throughout the duration of the mine life or will be stored in the NAG stockpile (Figure 18.3).

18.10.2 POTENTIALLY ACID GENERATING WASTE ROCK

Approximately 86.7 Mm³of the waste rock is estimated to be PAG waste rock. In order to minimize future oxidation of this PAG waste rock and the associated potential release of contaminants, it is planned to place the PAG waste rock together with the tailings (co-disposal) in the lined TSF keeping the PAG rock permanently under water to limit oxidation and thus prevent generation of ARD (Figure 18.3).

18.11 WATER MANAGEMENT

This section describes the proposed measures required for effective management of surface water and sediment during construction and operation of the Arctic Project. Further details can be found in Appendix D.

The overall goal of the water management plan is to:

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•	minimize the amount of contact water (water resulting from interaction with mine waste and other elements from mining activities) produced during project construction, operation and following closure
•	collect contact water to enable reuse, or discharge after treatment (unless
	compliance with water quality standards can be demonstrated)
•	provide and retain water for mine operations
•	provide a basis for management of the freshwater on the site
•	minimize the impact of mine construction and operations activities on the natural surface water drainages
•	optimize the use of surface water generated within the footprint area of each facility.

To meet these objectives, the applied design methodology focused on identifying and managing separately all contact water and non-contact water types and also sediment.

The definition and design of separate conveyance systems for each water type was based on the following criteria:

• Non-contact water is defined as all surface and subsurface waters from areas not disturbed by the mining project, regardless of its quality. This is water from natural drainage channels that may be intercepted, diverted and discharged into natural drainages located downstream of the project facilities. Non-contact water will be managed mainly through diversion works around specific mine infrastructure and through temporary storage ponds.

• Contact water is defined as all water which will be affected by the proposed mining activities and will require special handling or treatment prior to use or being returned to the natural drainage, to ensure compliance with corresponding Alaskan water quality standards. Contact water is generated when naturally occurring waters come into contact with mining infrastructure, mining generated waste rock or processed materials that alter its physical-chemical properties. This classification also includes water from areas disturbed by mining activities with high potential for generation of increased sediment loads. Contact waters with increased sediment loads will be transported to sediment retention ponds to facilitate the reduction of suspended solids loads through settlement. After reduction of sediment loads, all contact water will be transported to the acid water treatment plant prior to its use in the mineral process or discharged to the natural drainage.

In summary, contact water is expected in facilities such as the pit, TSF, and NAG stockpile. This water will be collected in designated ponds for treatment, as necessary, prior to its release and will be used in the process to the maximum extent possible. Contact water from direct run-off from the NAG stockpile (and from un-diverted runoff from contributing basins) and other facilities will be diverted to a collection pond, for treatment (if required) and subsequent downstream discharge to natural stream courses.

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Contact water from direct run-off from the open pit and any possible groundwater (if any) will be accumulated in the bottom of the operating pit and subsequently pumped via gravity to the tailings pond.

On the other hand, non-contact water will be directed around the facilities via diversion channels or culverts and discharged into the natural drainages upon achievement of sediment removal to acceptable levels.

18.11.1 SITE FACILITIES

The following section describes the water management structures expected at each of the following facilities:

• open pit

• plant site area

• waste rock facilities

• TSF.

Separate strategies are considered for the management of contact and non-contact waters and sediments. A summary of the required infrastructure for each plan is shown in Table 18.1 and Table 18.2.

Table 18.1 Components for Management of Non-contact Surface Waters

Infrastructure	Non- contact Water System	Final Disposal
NAG Stockpile	Diversion Channel	Discharge to Natural Drainage
Process Plant	Diversion Channel	Discharge to Natural Drainage
Tailings Dam	Diversion Channel	Discharge to Natural Drainage

Table 18.2 Components for Management of Contact Surface Waters and Sediments

Infrastructure	Contact Water System	Final Disposal
NAG Stockpile	Drains, Perimeter Channel and Sediment Ponds	Tailings Pond or Acid Water Treatment Plant
Tailings Dam	Drains And Collection Sump	Acid Water Treatment Plant (if needed)
Open Pit	In-pit Sump	Tailings Pond

OPEN PIT

The open pit footprint will be dynamic as mining progresses and will be treated as a closed system with all direct rainfall and local runoff being treated as contact water and collected in a sump located in the pit bottom. The captured water will be pumped and conveyed via gravity to the tailings pond.

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PLANT SITE

The plant site facility will be located north of the NAG stockpile. This facility will have a diversion channel around the entire facility and will route non-contact surface runoff from west, north and east catchments areas. All the runoff collected by the diversion channel will discharged to the natural drainage located downstream of the TSF collection sump.

WASTE ROCK DUMP

Given that all contact waters are linked to the waste rock facilities, proper water management will require that all direct surface runoff and all seepage water from the waste rock facilities is collected and conveyed via gravity to the tailings pond.

All direct surface runoff will be collected and controlled via a series of surface channels and drains around the immediate perimeter of the slopes of the waste rock facilities. Seepage collection structures will comprise of a set of perforated pipes and drains that will be located beneath and at the base of the waste rock facilities in order to capture and convey water to the toe of the waste rock facilities. Both surface water runoff and seepage collection systems will discharge via gravity to the tailings pond. From the tailings pond, water will be pumped to the water treatment plant when needed.

In addition to the contact water management structures, the waste rock facilities will have diversion channels located on the east and west side of the structure. This channels will conveyed non-contact surface runoff from existing catchments to the natural drainage located downstream of the TSF collection sump.

TSF

Water management structures expected for the TSF include (1) reclaim water system, (2) seepage collection system and (3) diversion channels.

Reclaim Water System

All water from the tailings will be collected in a supernatant pond formed at the tailings and will be returned to the mill for re-use in the process via a floating barge pump-back system. This solution will be transferred to the mill by a pipeline.

Seepage Collection System

The seepage collection sump will collect seepage from below the TSF embankment (baseflow expected to come from the existing upstream watercourse) via an under-drain system. The under-drain system will be composed of perforated pipes encapsulated in gravel. The seepage collection system will be used for monitoring and will be directed to the AWTP if needed.

Diversion Channels

Two diversion channels (west and east) will be constructed to direct surface water runoff around the TSF during the life of the Arctic Project. The water diversion system associated to the TSF will collect and divert non-contact surface runoff from upstream to the natural drainage located downstream of the proposed collection sump.

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It is estimated that temporary diversion channels will be constructed during operations and will be removed as operations progress.

18.12 WATER TREATMENT PLANT

An allowance of $9 million has been included in the capital cost estimate for a water treatment facility.

18.13 CONSTRUCTION AND PERMANENT CAMP ACCOMMODATION

On site camp accommodations for construction personnel will be provided during the construction phase. The current design will include:

• 700 single-occupancy rooms

• a dormitory building

• a kitchen

• a recreation complex

• an administration complex

• a sewage treatment facility

• potable water supply system.

After construction is completed, a portion of the temporary camp will be dismantled; the remaining camp will be refurbished and will comprise 350 single-occupancy rooms as well as the above facilities.

18.14 COMMUNICATIONS

The telecommunications design for the Arctic Project will incorporate proven, reliable, and state-of-the-art systems to ensure that personnel at the mine site will have adequate data, voice, and other communications channels available.

The telecommunications system will be supplied as a design-build package. The base system will be installed during the construction period, and then expanded to encompass the operating plant. The design will include:

• a Voice over Internet Protocol (VoIP) telephone system

• satellite communications for voice and data

• Ethernet cabling for site infrastructure

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• wireless Internet access

• two-way radio communications at site

• satellite TV.

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19.0 MARKETSTUDIESANDCONTRACTS

The Arctic Project will produce copper, lead and zinc concentrates. The copper and zinc concentrates will be shipped out in bulk while the lead concentrate will be shipped off site in containers. These concentrates will be trucked to Port MacKenzie located north of Anchorage, then the concentrates will be transported by ocean to smelters. This port operates 12 months per year and provides good access to Asia.

There are no market studies or contracts material to the Arctic Project, other than the NANA Agreement described in Section 4.3.2, that are required for property development.

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20.0	ENVIRONMENTAL STUDIES, PERMITTING,
	AND SOCIAL OR COMMUNITY IMPACT

This section characterizes the existing environmental baseline data for the Arctic Project area, makes suggestions for additional studies that would provide a basis for the mine permitting efforts, describes the major environmental permits that will likely be required for the Arctic Project, and identifies potential significant social or community impacts.

20.1 ENVIRONMENTAL STUDIES

The Arctic Project area includes the Ambler Lowlands and the Subarctic Creek drainage and would be sited within the Shungnak River drainage. To date, a limited amount of baseline environmental data collection has occurred in the area including surface water quality sampling, wetlands mapping, stream flow monitoring, aquatic life surveys, subsistence, meteorological monitoring, and acid base accounting data. The existing data are summarized in Sections 20.1.1 to 20.1.5.

20.1.1 HYDROLOGY AND WATER QUALITY DATA

A number of sampling efforts have been used to characterize the hydrology in the Arctic Project area. In 2007, 2008 and 2009, Shaw Environmental collected water quality samples and measured stream flow at 13 stations on the Shungnak River, Subarctic Creek, Arctic Creek, and the Kogoluktuk River (Shaw 2007 to 2009).

In July 2010, Tetra Tech performed baseline studies to characterize flow and water quality in streams and other actively flowing drainages that could be potentially impacted by construction and operation of a proposed access road between the Bornite airstrip and the Arctic airstrip, and the existing road between the Arctic airstrip and the Arctic Deposit. Tetra Tech collected water quality and flow data at 14 sites. The results of the Tetra Tech sampling program indicate that, in general, the water quality for all meets applicable Alaska State water quality standards (WQS) for the parameters analyzed. It is noteworthy that the samples were collected during base flow conditions in the absence of any rainfall events. Since these data were conducted during low flow, they can be used to characterize baseline water quality conditions prior to any site disturbance. The Tetra Tech studies are described in detail in Arctic Deposit Access Environmental Baseline Data Collection – Hydrology, Ambler Mining District, Alaska (Tetra Tech 2010a).

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20.1.2 WETLANDS DATA

Tetra Tech also performed a program of jurisdictional wetlands identification in a portion of the Arctic Project area in 2010 as part of a study to identify potential road alignment alternatives between the Bornite and Arctic airstrips. The work included data review, vegetation mapping, aerial photographic interpretation (segmentation), and field soil surveys. The work is described in detail in Arctic Deposit Access Environmental Baseline Data Collection - Wetlands & Vegetation, Ambler Mining District, Alaska (Tetra Tech 2010b) and is summarized in this section.

The area between the Bornite airstrip and the Arctic airstrip consists of a wide valley containing the Ambler Lowlands and the Shungnak River. Wetlands are prevalent throughout much of the Ambler Lowlands. The majority of the wetlands within the area occur within tundra communities composed primarily of ericaceous shrubs, such as bog blueberry shrubs (Vaccinium uliginosum and V. vitis-ideae) and graminoids, such as cottongrass (Eriophorum vaginatum) and sedges (Carex bigelowii and C. aquatilis). Spruce forests (Picea glauca and P. mariana) and shrub birch communities (Betula nana and B. glandulosa) make up most of the upland communities.

Owing largely to the low resolution of the aerial imagery, Tetra Tech ultimately used secondary indicators including colour, texture (shadows), local topography, orientation, and slope in the vegetation mapping process. Using the combination of the aerial imagery, ground field reconnaissance surveys, contour mapping and these secondary indicators, Tetra Tech was able to make a distinction between wetland and upland vegetation communities.

20.1.3 AQUATIC LIFE DATA

Tetra Tech’s sampling efforts in 2010 included baseline aquatic life surveys in the area along the proposed road alternatives between the Bornite airstrip and Arctic airstrip, and along the Arctic airstrip to Arctic Deposit road in Subarctic Creek. The purpose of this study was to characterize the aquatic life within the Shungnak River and potentially impacted tributaries. Opportunistic observations were also collected in the Kogoluktuk River. Fish and macroinvertebrate data were collected from July 8 to 14, 2010. The study is described in detail in Arctic Deposit Access Environmental Baseline Data Collection – Aquatics, Ambler Mining District, Alaska (Tetra Tech 2011) and is summarized in this section.

Tetra Tech employed active fish capture methods to assess the local fishery population, and backpack electrofishing gear to sample reaches of smaller streams. Tetra Tech also employed passive fish capture techniques, including gill netting, minnow traps, hoop nets and visual observation.

The aquatics sampling plan closely followed the streams included in Tetra Tech’s water quality sampling campaign. Six different fish species were captured or observed in the study area as summarized in Table 20.1 and excerpted from Tetra Tech (2011). The lack of large or anadromous fish in sampled sections of the Shungnak River is likely the result of the presence of a waterfall, estimated to be 4.5 to 9 m tall, situated downstream of the Arctic Project area.

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Table 20.1 Captured or Observed Fish Species

Common Name	Scientific Name	Number Captured/Observed	Capture/Observation Location*
Slimy sculpin	Cottus cognatus	122	SUB-01, RUB-01, STR-01, STR-02, STR-03, STR-05, SHU-03
Alaska blackfish	Dallia pectoralis	1	STR-05
Northern pike	Esox lucius Linnaeus	2	SHU-04, KOG
Round whitefish	Prosopium cylindraceum	140	RUB-01, SHU-01, SHU-02, SHU-03, KOT-01
Dolly Varden	Salvelinus malma	39	SUB-01, SUB-02, RUB-01, STR-03, STR-04, STR- 05, KOT-01
Arctic grayling	Thymallus arcticus	159	SUB-01, STR-03, STR-05, SHU-01, SHU-02, SHU-03, SHU-04, KOG, KOT-01

Notes:^*SUB = Sub-Arctic Creek; RUB = Ruby Creek; STR = Unnamed Stream Designation;
SHU = Shungnak River; KOG = Kogoluktuk River; KOT = Unnamed Kogoluktuk River Tributary

Benthic macroinvertebrates were also collected during the July 2010 survey period from nine monitoring sites. The overriding goal of this assessment was to describe conditions as they existed at that time as there were no known previous data that could be used for comparison. Applying indices of water quality and overall stream health, to the results of the macroinvertebrate sampling, showed that most of the sampled streams in the Arctic Project area are in good ecological condition. They generally demonstrate relatively high levels of species diversity and species richness although results varied from stream to stream.

Total species abundance and richness, Ephemeroptera, Plecoptera, Trichoptera (EPT) index, and Fine Sediment Biotic Index (FSBI) metrics were plotted as a measure of stream health in the survey area. That data is illustrated in Figure 20.1 and excerpted from Tetra Tech (2011).

Based on these metrics, six monitoring sites in Subarctic Creek, Ruby Creek and an unnamed creek (SUB-01, SUB-02, STR-03, STR-04, RUB-01 and SHU-01) exhibit higher levels of stream health and therefore have greater sensitivity to impairment from sediment or changes in water quality. Analysis of monitoring sites in some of the unnamed creeks (STR-01, STR-02 and STR-05) suggest these streams exhibit lower levels of stream health and would likely show less sensitivity to effects from development.

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Figure 20.1 Comparison of Benthic Macroinvertebrate Community Measures

20.1.4 SUBSISTENCE DATA

Access to the Arctic Project requires access through private lands owned by NANA. NovaCopper acknowledges the importance of subsistence, as such, a Subsistence Subcommittee of the Oversight Committee comprised of locally appointed residents from six potentially affected communities in the region has been formed to review and discuss subsistence issues related to the project and to develop future compliance plans. A representative from NANA and NovaCopper facilitate the meetings and report a summary of the discussions and recommendations provided by the Subsistence Subcommittee to the Oversight Committee.

A formal subsistence survey has not been performed in the immediate vicinity; however, NovaCopper has established a workforce “Wildlife Log” to document potential subsistence resources, species diversity and human/wildlife encounters. In 2012, Stephen R. Braund & Associates completed a subsistence data gap memo under contract to the Alaska Department of Transportation and Public Facilities as part of the baseline studies associated with a proposed road to the Ambler mining district. The purpose of this analysis was to identify what subsistence research had been conducted for the potentially affected communities, determine if subsistence uses and use areas overlap with or may be affected by the project, and identify what, if any, additional information (i.e., data gaps) needed to be collected in order to accurately assess potential effects to subsistence (Braund 2012). Among other topics the report outlined historic subsistence uses including maps and a literature review, and also provided a synopsis by village including those villages closest to the Arctic Project, and suggested further study.

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Previous sampling efforts have established the presence of various salmon species, northern pike and sheefish in the Kogoluktuk River. Sampling efforts in the Shungnak River have established the presence of northern pike. The presences of fish are good indicators of the possibility of subsistence use of these rivers, but boat access is limited due to waterfalls and rapids. In comparison, the Kobuk River, a wide and easily navigable river on which the communities of the region exist, supports the bulk of subsistence fishing.

Determining the presence and distribution of caribou is complex because of seasonal and annual variability in migration patterns. In 2012, Tetra Tech provided a report to NovaCopper outlining options for documenting the caribou migration patterns in the area and the subsistence use of caribou and other resources by the three communities (Ambler, Shungnak, and Kobuk) closest to the Arctic Project area. The ADF&G also has information about caribou population density and migration patterns in the area (ADF&G 2011).

The NWAB, through its Title 9 Conditional Use permit, regulates the project with respect to caribou interactions to assure the migration is minimally affected by mining and exploration activities. To this end, NovaCopper has been in contact with the ADF&G wildlife biologists to monitor movements of caribou in the spring and fall in proximity to the Arctic Project. Radio collar data provided by the State of Alaska is used to track herd movements. Summary maps of those movements constructed from years of radio collar information indicate three main migration corridors to the west of the Arctic Project area, the nearest one is approximately 48 km west of the Arctic Project area.

20.1.5 ACID BASE ACCOUNTING DATA

Two sampling efforts have been used to characterize the acid generation potential of the mine waste for the Arctic Project, and the constraints that this may place on either open pit or underground mining development. In 1998, Robertson collected 60 representative core samples from the deposit for their acid/base characteristics; these samples provided a broad assessment of ARD at the Arctic Deposit with a focus on characterization for surface development. In 2010, SRK collected 148 samples and prepared a preliminary ML and ARD analysis of the ML/ARD potential of waste rock at the Arctic Deposit (SRK 2011). The SRK report focuses on characterization for underground development rather than an open pit scenario; however, it does provide a more refined analysis of ARD potential based on advances that have been made in understanding the importance of sulphide mineralogy in assessing ARD. The criteria used for classifying ARD potential also differs slightly from the Robertson era work.

Based on recommendations from SRK, Tetra Tech used total sulphur percent as a proxy for AP and calcium percent as a proxy for NP to better characterize ABA in an open pit mine plan scenario. Overall, 55% of the mine waste rock is currently classified as PAG.

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20.1.6 ADDITIONAL BASELINE DATA REQUIREMENTS

Additional baseline environmental data in the Ambler Lowlands, the Subarctic Creek drainage, the Shungnak River drainage and downstream receiving environments will be required to support future mine design, development of an EIS, permitting, construction and operations. NovaCopper is advised to consult with state, local and federal regulatory agencies and their consultants to further develop a comprehensive environmental baseline program. Owing to the long lead-time to collect data (years), it is important that the comprehensive environmental baseline program generates adequate data in terms of type, quality and quantity for each of the disciplines of interest. Recommendations for additional baseline studies are included Table 20.2.

Table 20.2 Additional Recommended Environmental Baseline Studies

Discipline	Recommended Studies
Acid-Base Accounting	Additional static test work of waste lithogeochemical domains within and adjacent to the proposed open pit and borrow sources and tailings followed by kinetic test work.
Archaeology	Assessment of cultural resources, cultural site clearance
Aquatic Life	Expanded aquatic surveys (invertebrates)
Ecosystem and Soils	Permafrost and wetlands delineation mapping; vegetation surveys
Fisheries	Expanded fisheries survey
Hydrogeology	Installation and monitoring of groundwater wells in the Subarctic Creek drainage areas near the site of, and down gradient of, the proposed tailings and waste rock storage facilities and in alternative sites for tailings and waste rock disposal
Hydrology	Snow survey
Meteorology, Air	Expansion of the meteorological program to additional locations to be determined;
Quality, and Noise	air quality monitoring
Wildlife	Avian survey, large mammal survey.

All of the data are important to the development of an accurate environmental baseline and water balance model for the Arctic Project area. These studies would need to be completed in sufficient depth to cover all reasonably foreseeable baseline work that may be requested by the regulatory agencies. The risks that come with insufficient baseline data include delays in the permitting process, poorly constrained pre-mining characterizations, inappropriate trigger levels in permits and inaccurate water balance models that can negatively affect operations and otherwise result in unforeseen and potentially costly circumstances during permitting or mine operations and closure.

20.2	PERMITTING
20.2.1	EXPLORATION PERMITS
	NovaCopper performs mineral exploration at the Arctic Deposit under State of Alaska and NWAB permits.

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NovaCopper is presently renewing their multi-year hardrock exploration permit issued by the ADNR. The renewal is anticipated by July 2013. NovaCopper is not currently performing mineral exploration activities on the Arctic Project during 2013. Cumulative surface disturbance for exploration activities on the Arctic Project remains less than 5 acres and therefore there are currently no State of Alaska requirements for reclamation bonding for the Arctic Project.

NovaCopper reports that the NWAB Title 9 Conditional Use Permit authorizing exploration and bulk fuel storage expires on December 31, 2014, and the permit for landfill, gravel extraction, and bridge construction permit expires after December 31, 2015. No bonding is required for the borough permits.

NovaCopper has obtained several other permits for camp support operations. These permits include a drinking water permit, a wastewater discharge permit (pending final approval), camp establishment permits, and construction and operation of a Class III Camp Municipal Landfill, all of which are issued by the ADEC. Temporary water use authorization issued by the ADNR, a Title 16 Fish Habitat permit and a wildlife hazing permit issued by the ADF&G.

20.2.2 MAJOR MINE PERMITS

The following discussion identifies the major permits and approvals that are likely required for the Arctic Project to be developed into an operating mine.

The types of major mine permits required by this project are largely driven by the underlying landownership; regulatory authorities vary depending on land ownership. The Arctic Project area includes patented mining claims (private land under separate ownership by NovaCopper and NANA), State of Alaska land, and NANA land (private land). In addition, federal land would likely be part of any access road between the Dalton Highway and the Arctic Project area. However permits associated with such an access road are being investigated in a separate action by AIDEA as part of their permitting of the AMDIAR and not addressed in this report. A list of likely major mine permits is included in Table 20.3.

Because the Arctic Project is situated at least partially on State of Alaska lands, it will likely be necessary to obtain a Plan of Operation Approval (which includes the Reclamation and Closure Plan) from the ADNR. The ADNR will also require certificates to construct and then operate a dam(s) (tailings and water storage), water use authorizations, an upland mining lease and a mill site lease, as well as several minor permits including those that authorize access to material sites.

The ADEC would require an integrated waste management permit, air permits for construction and then operations, and an Alaska Pollution Discharge Elimination System (APDES) permit for any wastewater discharges to surface waters, and a multi-sector general permit for stormwater discharges. The ADEC would also be required to provide a 401 Certification for the Federal Section 404 permit (Table 20.2).

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The ADF&G would require permits for any culverts or bridges that are required to cross fish-bearing streams or for other impacts to fish-bearing streams.

The USACE would require aClean Water Act (CWA) Section 404 permit for dredging and filling activities in “waters of the US,” including jurisdictional wetlands.

The USACE Section 404 permitting action would require the USACE to comply with the NEPA and, for a project of this magnitude, the development of an EIS. The USACE would likely be the lead federal agency for the NEPA process. The NEPA process will require an assessment of direct, indirect and cumulative impacts of the Arctic Project and the identification of project alternatives, and include consultation and coordination with additional federal agencies, such as the US Fish and Wildlife Service (if endangered or threatened species are present) and National Marine Fisheries Service (if essential fish habitat is present), and with the State Historic Preservation Office and Tribal Governments under Section 106 of theHistorical and Cultural Resources Protection Act.

The Arctic Project will have to meet USACE wetlands guidelines to avoid, minimize and mitigate impacts to wetlands and, as part of the Section 404 permitting process, the USACE will require NovaCopper to develop a compensatory wetlands mitigation plan. This will be an in-depth plan for mitigating unavoidable wetlands impacts and should include the input from reclamation and mitigation experts as well as the purchase of wetland “credits” through an in-lieu fee program.

The Arctic Project will also have to obtain approval for a Master Plan from the NWAB. In addition actions will have to be taken to change the borough zoning for the Arctic Project area from Subsistence Conservation to Resource Development.

The overall timeline required for permitting would be largely driven by the time required for the NEPA process, which is triggered by the submission of the 404 permit application. The timeline includes the development and publication of a draft and final EIS and ends with a Record of Decision and an appeal period. In Alaska, the EIS and other State and Federal permitting processes are generally coordinated so that permitting and environmental review occurs in parallel. The NEPA process could last between 1.5 to 3 years, but could potentially be longer.

Table 20.3 Major Mine Permits Required for the Arctic Project

Agency	Authorization
State of Alaska
ADNR	Plan of Operations Approval (including Reclamation and Closure Plan)
	Upland Mining Lease
	Mill Site Lease
	Reclamation Bond
	Certificate of Approval to Construct a Dam
	Certificate of Approval to Operate a Dam
	Water Rights Permit to Appropriate Water

table continues…

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Agency	Authorization
ADF&G	Title 16 Permits for Fish Habitat (authorize stream crossings)
ADEC	APDES Water Discharge Permit
	Alaska Multi-Sector General Permit (MSGP) for Stormwater
	Stormwater Discharge Pollution Prevention Plan (part of MSGP)
	Section 401 Water Quality Certification of the CWA Section 404 Permit
	Integrated Waste Management Permit
	Air Quality Control – Construction Permit
	Air Quality Control – Title V Operating Permit
	Reclamation Bond
Federal Government
EPA	Spill Prevention, Control, and Countermeasure (SPCC) Plan (fuel transport and storage)
USACE	CWA Section 404 Dredge and Fill Permit
NWAB
NWAB	Master Plan Approval and rezoning lands from Subsistence Conservation to Resource Extraction

Note: ”Major” permits define critical permitting path. Additional “minor” permits are also required.

20.3 SOCIAL OR COMMUNITY CONSIDERATIONS

The Arctic Project is located approximately 25 miles northeast of the native villages of Shungnak and Kobuk, and 40 miles east-northeast of the native village of Ambler. The population in these villages are 151 in Kobuk (2010 Census), 262 in Shungnak (2010 Census) and 258 in Ambler (2010 US Census). Residents live a largely subsistence lifestyle with incomes supplemented by trapping, guiding, local development projects, government aid and other work in, and outside of, the villages.

The Arctic Project has the potential to significantly improve work opportunities for village residents. NovaCopper is working directly with the villages to employ residents in the ongoing exploration program as geotechnicians, drill helpers, and environmental technicians. NovaCopper and NANA have established a Workforce Development Subcommittee of the Oversight Committee, described below, to assist with developing a local workforce. In addition, NovaCopper has existing contracts with native-affiliated companies (such as NANA Camp Services (NMS) and WHPacific Inc.) that are providing camp catering and environmental services for the project, respectively.

In October 2011, NovaCopper signed an agreement with NANA. In addition to consolidating landholdings in the Ambler District, the agreement has language establishing native hiring preferences and preferential use of NANA subsidiaries for contract work. Furthermore, the agreement formalized an Oversight Committee, with equal representation from NovaCopper and NANA, to regularly review project plans and activities. In addition, a Subsistence Subcommittee has been formed to protect subsistence and the Iñupiaq way of life and a Workforce Development Subcommittee is also in place to address current and future employment needs on the project through the development of training and educational programs that build skill sets for local residents interested in exploration and mining careers. The agreement also includes a scholarship funded annually by NovaCopper that promotes education for youth in the region. NovaCopper meets monthly, during summer months, with the residents of Kobuk, Shungnak and Ambler, the three villages closest to the project area. NovaCopper also meets annually with eight other NANA region villages including Noatak, Kivalina, Kotzebue, Kiana, Deering, Buckland, Selawik and Noorvik, for the purpose of updating residents on project plans and fielding their questions and concerns. NovaCopper has also developed a good working relationship with the NWAB government.

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In general terms, rural Alaska residents are often concerned about potential mining impacts to wildlife and fish for those projects within their traditional use areas NovaCopper appears to have acknowledged these views and concerns and is taking substantive steps to address them during the current exploration stage of the Arctic Project.

Local community concerns will also be formally recognized during the scoping stage at the beginning of the NEPA process. At that time, the lead federal agency (likely the USACE) will hold scoping meetings in rural villages to hear and record the concerns of the local communities so that the more significant of these can be addressed during the development of the EIS. In addition, the USACE would have government-to-government consultations with the Tribal Councils in each of the villages, as part of the NEPA process, to discuss the project and discuss Council concerns.

Characterizing the level of support or opposition to the Arctic Project would be speculative at this time. A poll conducted by Dittman Research for the 2011 NANA Shareholder opinion survey asked if Shareholders supported or opposed road projects on NANA land to assist in economic and potential mineral development. Eighty three percent supported the concept while 15% opposed. Surveys of this sort show a broad support for infrastructure and of mineral development indirectly in the region as long as regional interests are met. Regional engagement by NovaCopper has also encountered a strong desire for the economic benefits that come with mining projects. However, like most mining projects there will likely be some opposition to this project.

20.4 MINE RECLAMATION AND CLOSURE

Mine reclamation and closure are largely driven by State of Alaska regulations (11 AAC 86.150, 11 AAC 97.100 -910, and 18 AAC 70) and statutes (Chapter 27.19) that specify that a mine must be reclaimed concurrent with mining operations to the greatest extent possible and then closed in a way that leaves the site stable in terms of erosion and avoids degradation of water quality as a result of ARD or ML on the site. A detailed reclamation and closure plan will be submitted to the agencies for review and approval in the future, during the formal mine permitting process. The approval process for the plan varies somewhat depending on the land status for any particular mine. Owing to the fact that the Arctic Project is likely to have facilities on a combination of private (patented mining claims and native land) and State lands, it is likely that the reclamation plan will be submitted and approved as part of the plan of operations, which is approved by the ADNR. However since the reclamation plan must meet regulations of both ADNR and the ADEC, both agencies will review and approve the Reclamation and Closure Plan. Generalized reclamation and closure strategies are presented in Section 20.5.1.

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20.4.1 RECLAMATION AND CLOSURE PLAN

A formal reclamation and closure plan for the Arctic Project will be developed as part of the formal mine permitting process in the future. The following is a general discussion of the reclamation and closure as it applies to the Arctic Project.

If the Arctic Project is developed as an open-pit mine, it will generate approximately 300 Mt of waste rock that will be managed during mining operations and permanently disposed of on the site as part of concurrent reclamation, mine closure and final reclamation. Waste rock will be characterized and a waste rock management plan will be developed and incorporated into the mining operations and final closure strategy. Preliminary acid-base-accounting data obtained from drill core samples indicate that the waste rock and tailings could potentially exhibit ARD/ML. Depending on the geochemical character, all or a portion of, the waste rock may be covered with an engineered soil or geomembrane cover at closure to minimize water and oxygen infiltration and oxidation of the waste rock and/or tailings. Under the current mine plan the waste rock with the potential to generate ARD/ML will be co-disposed with the tailings and managed under a wet cover to minimize oxidation

The reclamation and closure plan will also address water management, including the monitoring of surface and groundwater at the site that will be required after mine closure.

The mineralized material will be processed in a mill at site and the tailings remaining after mineral extraction will report to the tailings waste stream. Under the present mine plan, the tailings will be conveyed by pipe as a slurry to a lined tailings facility. Closure for the tailings facility will be a “wet” closure, where water is managed in order to keep the tailings and waste rock submerged in perpetuity to minimize the potential of ARD/ML. Long term water management may be required for the wet closure scenario. The reclamation and closure plan will address the closure of the tailings facility in more detail.

Typically, mine buildings and other mine infrastructure are removed at mine closure, and disturbed or otherwise-impacted land is re-graded, covered with growth media and re-vegetated. This would be the general approach to closure for the Arctic Project, except that State of Alaska regulations allow some flexibility to accommodate the preferences of the surface land owner and the intended post-mining land use. In this case, the mine facilities may be located on a combination of private and State lands. It would be permissible, under current regulations, to leave certain buildings or other facilities in place at closure if the private land owners (patented claim owners and/or NANA) prefer. Those details will be included in the reclamation and closure plan that will be submitted during the formal mine permitting process. State of Alaska agencies require a letter of consent from the surface land owner that acknowledges that the intended mining use and closure plans are acceptable to them. The letter is required in advance of any mine construction activities.

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The reclamation and closure plan will also include many other important details including growth media inventories, water management plans and post closure monitoring plans that will measure the efficacy of the mine reclamation activities with regard to protecting the environment for a significant period following mine reclamation and closure, likely 30 years.

A number of studies will be required to develop a more detailed reclamation and closure plan, and support mine permitting and mine design, prior to mine development.

A site-wide surface water and constituent load balance will be required to predict water shortfalls or surpluses during operations and at closure. The results of the study will drive the need for any water use authorizations, surface water discharge permits and establish any need for water treatment capabilities for the Arctic Project.

Groundwater studies will be required to provide a range of important information including the flood rate of the open pit following closure. In addition, it will be necessary develop a model to predict quality of the pit lake water after closure in order to assess the possible need to treat that water if modeling predicts that the lake will eventually fill and discharge water to the environment.

Acid-base-accounting study work will need to be performed in order to characterize the geochemistry of the tailings and the waste rock. The results of these studies will be used to develop waste rock and tailings management plans for operations and also to support the closure strategy as it applies to the tailings and waste rock storage facilities.

During operations additional studies using test plots on the site may be performed that will help develop the final details for tailings and waste rock cover designs, re-vegetation, and water management of the site during closure and reclamation.

Reclamation and closure costs will be developed as more detailed engineering and study work is completed that is sufficient to support a detailed reclamation and closure plan. Current financial assurance amounts, which are reflective of closure and reclamation costs, for Alaska’s six operating metal mines range from $6 million to $305 million. Of those, only two mines are open pit mines (Red Dog and Fort Knox mines) and the financial assurance amounts for those are $68 million and $305 million, respectively. The Red Dog financial assurance amount is largely driven by the need for perpetual water treatment which accounts for approximately $255 million of the total financial assurance amount. For this PEA, Tetra Tech estimates that the closure costs for the Arctic Project will be between $30 million and $60 million, exclusive of any potential need for long-term water treatment following mine closure. This estimate is based on: 1) comparisons with other mines that share similarities with the Arctic Project, and 2) other industry experience.

20.4.2 RECLAMATION AND CLOSURE FINANCIAL ASSURANCE

There are two State of Alaska agencies that require financial assurance in conjunction with approval and issuance of large mine permits. The ADNR, under authority of Alaska Statute 27.19, requires a reclamation and closure plan be submitted prior to mine development and requires financial assurance to assure reclamation of the site, typically prior to construction. The ADEC requires financial assurance both during and after operations, and to cover short and long-term water treatment if necessary, as well as reclamation and closure costs, monitoring, and maintenance needs. The financial assurance must also include the property holding costs for a one-year period. The financial assurance amount is calculated through the process of reviewing and approving the Arctic Project reclamation and closure plan during the permitting process. Both State agencies also have requirements that any private land owners (patented mining claim owners and NANA) provide their concurrence with the use of their lands for mining purposes and that the reclamation and closure plan for the mine is consistent with their intended post-mining land use.

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NovaCopper may satisfy the State of Alaska financial assurance requirement by providing any of the following: (1) a surety bond, (2) a letter of credit, (3) a certificate of deposit, (4) a corporate guarantee that meets the financial tests set in regulation by the ADNR commissioner, (5) payments and deposits into the trust fund established in AS 37.14.800, or (6) any other form of financial assurance that meets the financial test or other conditions set in regulation by the ADNR commissioner. The adequacy of the reclamation and closure plan, and the amount of the financial assurance, are reviewed by the State agencies at a minimum of every five years and may be reviewed whenever there is a significant change to the mine operations, or other costs that could affect the reclamation and closure plan costs.

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21.0 CAPITALANDOPERATINGCOSTS

21.1 INTRODUCTION

Tetra Tech has developed a capital cost estimate (CAPEX) and operating cost estimate (OPEX) for the Arctic Project. A summary of both the CAPEX and OPEX is provided in Table 21.1, and is discussed in greater detail in the subsections that follow. The CAPEX and OPEX provide the basis for the economic analysis in Section 22.0.

Table 21.1 Summary of Capital and Operating Costs

Cost Type	Total ($ million)	Unit Cost ($/t milled)	Estimate Accuracy Range (%)
Total Capital Costs	717.7	-	±35
Total Sustaining Capital for LOM	164.4	-	±35
Total Operating Costs*	-	63.93	±35

Note:^*Excluding pre-production cost and including the road toll cost for the proposed AMDIAR.

21.2	CAPITAL COST ESTIMATE
21.2.1	INTRODUCTION
	The total estimated CAPEX for the design, construction, installation and commissioning of the Arctic Project is $717.7 million. Closure costs are estimated to be $81.6 million and are included in the financial analysis.
21.2.2	CAPITAL COST SUMMARY
	The total initial and sustaining capital cost summary is shown in Table 21.2 and  illustrates the costs separated out by the work breakdown structure (WBS) area.
21.2.3	TOTAL SUSTAINING CAPITAL COSTS
	The total sustaining capital costs of $164.4 million for the 12 year LOM including equipment, tailings and other items are summarized in Table 21.3.

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Table 21.2 CAPEX Summary

Item		Total Cost ($ million)
Initial Capital Estimate
10	Overall Site	82.5
20	Open Pit Mining	119.7
35	Mineralized Material Handling	17.4
40	Process Plant	122.2
50	Tailings and Water Management	21.0
70	On Site Infrastructure	49.1
75	Airstrip	14.2
86	External Access Roads	27.2
87	Temporary Services	23.1
Subtotal		476.4
90	Indirect Costs	130.9
98	Owner’s Costs	18.6
99	Contingency	91.9
Initial Development Capital		717.7

Note: Total made not add up due to rounding.

Table 21.3 Summary of Total Sustainable Capital Costs for LOM

Item	Total Cost ($ million)
Mining Equipment	45.6
Tailings	112.8
Other Equipment	6.0
Total Sustaining Capital	164.4

21.2.4 PURPOSE AND CLASS OF ESTIMATE

The purpose of this CAPEX is to estimate the capital cost of the Project contemplated in the scope of work included in the NI 43-101 technical report and PEA.

This estimate has been prepared in accordance with the recommended practices of the AACE International. In accordance with the AACE’s International Estimate Classification System, this cost estimate meets or exceeds the specifications for a Class 5 Estimate and has a deemed accuracy of ±35%.

21.2.5 ESTIMATE BASE DATE, EXCHANGE RATE, AND VALIDITY PERIOD

Tetra Tech prepared this estimate with a base date of Q2 2013. Quotations for major pieces of equipment used in this estimate were obtained in Q2 2013 and have a 90-day period of validity. Costs for all other equipment and materials were obtained from similar projects and are not based on definitive quotations.

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The CAPEX uses US dollars as the base currency. Foreign exchange rates are noted and applied as required (Table 21.4). All costs presented in this section are in US dollars unless otherwise stated.

Table 21.4 Foreign Exchange Rate Summary

Base Currency	Foreign Currency
US$1.00	Cdn$1.00
US$1.00	AUD$1.00
US$1.00	€0.81

Duties and taxes have not been included in this CAPEX.

The International System of Units (SI) has been used in the preparation of this estimate.

21.2.6 CONTRIBUTORS TO THE ESTIMATE

Tetra Tech developed the estimate in conjunction with:

• EBA (a Tetra Tech company): access road and airport

• NovaCopper: Owner’s costs.

21.2.6 ELEMENTS OF COST

The estimate consists of four main parts:

• direct costs

• indirect costs

• contingency

• Owner’s costs.

21.3.1 DIRECT COSTS

AACE defines direct costs as:

“…costs of completing work that are directly attributable to its performance and are necessary for its completion. In construction, (it is considered to be) the cost of installed equipment, material, labor and supervision directly or immediately involved in the physical construction of the permanent facility.”

Examples of direct costs are mining equipment, process equipment, mills and permanent buildings.

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Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

The largest components included in the direct costs are the mining pre-production costs ($62.5 million), the mining equipment ($54.6 million), the mineralized material handling ($17.4 million) and the processing costs (mill building: $59.9 million; grinding/flotation/refining: $45.7 million; concentrate handling: $12.3 million). Tailings disposal and reclaim cost total $21.0 million. The total estimated cost for on-site infrastructure is $49.1 million and consists of $22.1 million for ancillary buildings, $18.8 million for site services and utilities and $8.1 million for the plant mobile fleet. The largest components of the off-site infrastructure include the airstrip ($14.2 million) and external access road ($27.2 million).

The construction camp ($12.6 million) has been included in direct costs and it will be refurbished to become part of the permanent operations camp. The full WBS for the Arctic Project and the labour summary can be found in Appendix E.

The total direct costs for the Arctic Project are $476.3 million.

21.3.2 INDIRECT COSTS

AACE defines indirect costs as:

“…costs not directly attributable to the completion of an activity, which are typically allocated or spread across all activities on a predetermined basis. In construction, (field) indirects are costs which do not become a final part of the installation, but which are required for the orderly completion of the installation and may include, but are not limited to, field administration, direct supervision, capital tools, startup costs, contractor's fees, insurance, taxes, etc.”

Examples of indirect costs in this estimate are construction support equipment including all contractor indirects ($52.0 million), all temporary facilities or equipment to support construction, travel costs (travel and labour costs), spares ($3.9 million), freight and logistics ($21.8 million), initial fills ($2.2 million), commissioning ($1.1 million) and engineering, procurement and construction management (EPCM) ($46.8 million).

The total indirect costs for the Arctic Project are $130.9 million.

21.3.3 OWNER’S COSTS

Owner’s costs are costs borne by the owner in support and execution of the Arctic Project.

The project execution strategy involves an EPCM organization supervising one or more general contractors. The allowance for Owner’s costs has been provided by Tetra Tech and confirmed by NovaCopper. Some of the items include home office staffing, home office travel, home office general expenses, field staffing, field travel, general field expenses, and Owner’s contingency.

The Owner’s costs for the Arctic Project are $18.6 million.

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Project, Ambler Mining District, Northwest Alaska

21.3.4 CONTINGENCIES

When estimating the cost for a project, there is always uncertainty as to the precise content of all items in the estimate, how work will be performed, what work conditions will be like when the project is executed, and so on. These uncertainties are risks to a project. The estimator is often aware of the costs to these risks, and based on experience, can even estimate their probable costs. The estimated costs are referred to by cost estimators as “cost contingency”.

The level of engineering effort and risk has been assessed for each area and a contingency allowance has been assigned for each activity. The contingency ranges from a high of 25% for the airstrip, mine access road and tailings and waste dump areas to a low of 5 % for construction indirects and EPCM. Concrete and steel have been assigned a cost contingency of 15%. Owners cost have been assigned a contingency of 10%.

The total contingency for the Project is $91.9 million. This amount represents 19.3% of the direct costs.

21.3.5 EXCLUSIONS

The following are not included in the CAPEX:

• force majeure

• schedule delays such as those caused by:

- major scope changes

- unidentified ground conditions

- labour disputes

- environmental permitting activities

- abnormally adverse weather conditions

• receipt of information beyond the control of the EPCM contractors

• cost of financing (including interests incurred during construction)

• duties

• taxes

• schedule acceleration costs.

21.3.6 COSTS INCURRED PRIOR TO RELEASE OF DETAIL ENGINEERING AND CONSTRUCTIONASSUMPTIONS

The following assumptions have been made in the preparation of this estimate:

• All material and installation subcontracts were competitively tendered on an open shop, lump sum basis.

• Site work is continuous and is not constrained by the Owner or others.

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• Skilled tradespersons, supervisors, and contractors are readily available.

• The geotechnical nature of the site is assumed to be sound, uniform, and able to support the intended structures and activities. Adverse or unusual geotechnical conditions requiring piles or soil densification have not been allowed for in this estimate.

21.4 MINING CAPITAL COST ESTIMATE

The mining CAPEX includes investment and operating costs associated with the construction period and considers the following items:

• pre-production costs

• initial investment, replacement spare components and contingencies associated with the mine equipment fleet

• other investment such as dispatch, training sets, etc.

A summary of pre-production mining costs is presented in Table 21.5.

Table 21.5 Summary of Mining Costs

Pre-production	Costs ($ million)
Mining Operations	62.5
Mining Equipment	54.6
Facilities	2.0
Dewatering	0.6
Total Pre-production	119.7

PRE-PRODUCTION COST

Pre-production cost includes the mining cost associated with the pre-stripping before the start-up of the processing plants. The estimated pre-production tonnage is 16 Mt and it is expected to take one year to remove this material and expose sufficient mineral to initiate production. Details of the pre-production costs are shown in Table 21.6.

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Preliminary Economic Assessment Report on the Arctic
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Table 21.6 Mining Pre-production Unit Costs

Pre-production	Costs ($ million)	Costs ($/t mined)
Loading	2.54	0.16
Hauling	15.85	0.99
Drilling	2.64	0.16
Blasting	10.27	0.64
Ancillary	1.19	0.07
Support	5.92	0.37
Labour	22.28	1.39
Dewatering	0.80	0.05
Other	0.97	0.06
Total	62.45	3.90

MINING FLEET INVESTMENT

The mining fleet investment has been developed in accordance with the mine development schedule and considered the useful life of each type of equipment. The full purchase prices of the mining equipment were obtained from estimations prepared by Tetra Tech, based on costs from quotations from equipment vendors and “Mine and Mill Equipment Costs an Estimator’s Guide 2010 and 2011” escalated to 2013 figures. The prices included freight to the site and commissioning. Table 21.7 and Table 21.8 outline the details of initial and sustaining capital mine equipment purchases.

Table 21.7 Initial Mining Equipment Fleet

Equipment	Number of Units	Total Capital Cost ($)
Hydraulic Shovels 14.4 yd3	2	7,249,032
Haul Truck (100 ton)	16	24,045,136
Articulated Truck (43.5 ton)	2	1,510,587
Diesel Drill (4.5 to 8.5' in)	4	5,060,000
Secondary Drill (4.5 to 5.5 in)	1	806,910
Wheel Loader 15 yd3	1	2,171,736
Wheel Loader 9 yd3	1	765,827
Track Dozer (17.3 ft)	4	5,919,850
Wheel Dozer (15.2 ft)	1	1,175,949
Grader (16 ft)	2	1,708,373
Water Truck (10,000 gal)	1	676,288
Service Wheel Loader	1	381,430
Vibratory Compactor	1	216,637
Integrated Tool Carrier	1	178,499
Excavator	3	1,010,312

table continues…

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Equipment	Number of Units	Total Capital Cost ($)
Flatbed Truck	1	132,413
Fuel/Lube Truck	2	1,183,191
Mechanics Service Truck	1	108,119
Welder Truck	1	81,599
Tire Service Truck	1	403,917
Snow Plow/Sanding Truck	1	432,477
Pick-up Truck	8	399,837
Truck Mounted Hydraulic Crane	1	363,426
Rough Terrain Forklift	1	134,639
Shop Forklift	1	73,439
Light Plant/Towers	8	114,239
Mobile Radios	50	51,000
Safety Equipment	1	81,599
Engineering/Geology Equipment (computers, software, licenses)	1	165,239
Maintenance Management System	1	5,610
Explosives Plant (facilities and buildings)	1	1,200,000
Surveying	1	50,000
Total		57,857,309

Note: Rounding as required by reporting guidelines may result in apparent summation differences.

Table 21.8 Sustaining Capital Mining Equipment Fleet

Equipment	Number of Units	Total Capital Cost ($)
Hydraulic Shovels 14.4 yd3	2	7,249,032
Haul Truck (100 ton)	8	12,022,568
Articulated Truck (43.5 ton)	2	1,510,587
Diesel Drill (4.5 to 8.5 in)	3	3,795,000
Secondary Drill (4.5 to 5.5 in)	1	806,910
Wheel Loader 15 yd3	1	2,171,736
Wheel Loader 9 yd3	1	765,827
Track Dozer (17.3 ft)	4	5,919,850
Wheel Dozer (15.2 ft)	3	3,527,846
Grader (16 ft)	4	3,416,746
Water Truck (10,000 gal)	0	0
Service Wheel Loader	1	381,430
Vibratory Compactor	1	216,637
Integrated Tool Carrier	0	0
Excavator	3	1,010,312
Flatbed Truck	0	0
Fuel/Lube Truck	0	0

table continues…

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Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Equipment	Number of Units	Total Capital Cost ($)
Mechanics Service Truck	1	108,119
Welder Truck	1	81,599
Tire Service Truck	0	0
Snow Plow/Sanding Truck	0	0
Pick-up Truck	22	1,099,551
Truck Mounted Hydraulic Crane	1	363,426
Rough Terrain Forklift	1	134,639
Shop Forklift	1	73,439
Light Plant/Towers	16	228,478
Mobile Radios	131	133,619
Safety Equipment	2	163,199
Engineering/Geology Equipment (computers, software, licenses)	2	330,477
Maintenance Management System	2	11,220
Explosives Plant (facilities and buildings)	0	0
Surveying	1	50,000
Total		45,572,248

Note: Rounding as required by reporting guidelines may result in apparent summation differences.

21.5	OPERATING COST ESTIMATE
21.5.1	SUMMARY

The total LOM operating cost for the proposed mine is estimated at $63.93/t milled on average. The estimate includes mining, processing, tailing management, G&A and surface service operating costs.

A total of 35.68 Mt of mineralization material from the open pit mine is expected to be processed during LOM based on the proposed mining schedule. The nominal annual process rate is approximately 3,650,000 t/a (LOM average annual process rate is approximately 2,973,435 t/a) or 10,000 t/d (LOM daily average rate is 8,146 t/d) at 365 d/a. The unit cost is estimated based on the LOM average mill feed rate. The accuracy for the estimate is expected to be within a range of ±35%. The breakdown operating cost estimates are presented in Table 21.9. Figure 21.1 shows the cost breakdown at the LOM average throughput.

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Project, Ambler Mining District, Northwest Alaska

Table 21.9 Overall Operating Cost Estimate

Area	LOM Average Unit Operating Cost ($/t milled)
Mining*	28.40
Processing**	19.86
G&A	8.92
Surface Services	3.48
Road Toll	3.27
Total	63.93

Notes:^*Excluding preproduction cost.
^**Including tailings management operation cost.

Figure 21.1 LOM Average Operating Cost Distribution

21.5.2 MINING OPERATING COST ESTIMATE

Tetra Tech prepared the operating cost estimate based on price information for consumables, maintenance and labour costs from other similar projects in the area.

RELEVANT CONSUMABLE PRICES

Table 21.10 shows consumable pricing used in the calculation of equipment operating costs.

Table 21.10 Relevant Consumables Prices

Description	Unit	Price
Fuel	$/L	1.182
Lube	$/L	1.725
Emulsion	$/kg	1.826
ANFO	$/kg	1.704

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LABOUR

The mine indirect manpower includes personnel allocated to mine management, mine operations, maintenance services, technical services and mine geology.

Labour costs were calculated using the yearly cost per labour category equal to an average of salaries from two comparatively similar Alaskan mining operations.

BLASTING SERVICE

The mine will contract out this service including the supply of mix trucks and the trained personnel who will carry out the delivery of the explosive mix to the drill hole and the blasting operation. The fixed cost of this service is estimated at $0.05/t mined not including consumables.

MINING OPERATING COST SUMMARY

The operating cost of the total mine movement was calculated based on all information described in this section. Table 21.11 summarizes the mining operating costs per activity for the production period and displays those same costs as unit costs in $/t mined.

Table 21.11 Mining LOM and Unit Operating Cost Summary

Production	LOM Cost ($ million)	Unit Cost ($/t mined)
Loading	51	0.16
Hauling	212	0.67
Drilling	48	0.15
Blasting	197	0.62
Ancillary	16	0.05
Labour	370	1.16
Support	82	0.26
Dewatering	16	0.05
Other	22	0.07
Total Costs	1,013	3.18

21.5.3 PROCESSING OPERATING COSTS

The estimated LOM average process operating cost is 19.86$ /t milled or $59.05 million per year. The estimate is based on a total mill feed of 35.68 Mt (LOM) or an average annual process rate of 2,973,435 t/a or an average daily process rate of 8,146 t/d and 365 d/a.

A summary of the plant operation costs is shown in Table 21.12.

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Project, Ambler Mining District, Northwest Alaska

Table 21.12 Summary of Process Operating Cost

Description	Labour Force	Annual Cost ($)	Unit Cost ($/t milled)
Labour Force
Operating Staff	14	2,013,000	0.68
Operating Labour	52	4,073,000	1.37
Maintenance	63	5,611,000	1.89
Metallurgy Lab and Quality Control	13	1,230,000	0.41
Subtotal Labour Force	142	12,927,000	4.35
Supplies
Maintenance Supplies	-	17,528,000	5.89
Operating Supplies	-	2,106,000	0.71
Subtotal Supplies	-	19,634,000	6.60
Power Supply (Mill and Tailings)	-	26,485,000	8.91
Subtotal Power	-	26,485,000	8.91
Total (Process)	142	59,046,000	19.86

The estimated labour force cost is $4.35/t milled. A total of 142 people are estimated for the process operation, including 14 staff for management and professional services, 52 operators for operating, 63 personnel for maintenance and 13 personnel for metallurgical lab and assaying. The estimate is based on 12 hours per shift, 24 h/d, and 365 d/a.

The operating cost for the operating supplies include mill and crusher liners, mill grinding media and reagents. Reagent consumptions are estimated from the laboratory test results and comparable operations. The consumable metal costs and reagent costs are from the current budget prices from potential suppliers or Tetra Tech’s database.

The maintenance supplies are estimated at $5.89/t milled, operating supplies are estimated at $0.71/t milled and the unit power cost is estimated to be $8.91/t milled. The power cost is estimated based on the average power requirement of 82.25 MWh/a and a unit electric energy price of $0.322/kWh for the electric power generated on site.

All operating cost estimates exclude taxes, permitting costs, or other government imposed costs, unless otherwise noted. The estimate includes:

• labour force, including supervision, operation, and maintenance; salary/wage levels are based on average labour rates for similar operations; a benefit burden rate of 41% including holiday and vacation payment, pension plan, various other benefits, and tool allowance

• power supply from the electricity generating on site

• crusher/mill liner and mill grinding media consumptions estimated from the BWi and the Tetra Tech’s in-house database

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• maintenance supply costs, including building maintenance costs, based on approximately 7% of major equipment capital costs

• laboratory supplies, service vehicles consumables and other costs based on Tetra Tech’s in-house database and industry experience

• reagent costs based on the consumption rates from the test results and quoted budget prices or Tetra Tech’s in-house database.

21.5.4 GENERAL AND ADMINISTRATIVE COSTS AND SURFACE SERVICES COSTS

The LOM average G&A costs are estimated to be $8.92/t milled. Tetra Tech and NovaCopper developed these costs which include:

• labour cost for administrative personnel

• services expenses related to general administration, human resources, safety and security

• allowances for insurance and licenses

• costs for camp and personnel transport

• sustainability, including environment, community liaison and engineering consulting.

A summary of the G&A costs are provided in Table 21.13.

Table 21.13 G&A Operating Costs

	Labour Force	Total Cost ($/a)	Unit Cost ($/t milled)
G&A Labour Force
G&A	36	4,091,000	1.38
G&A Hourly Personnel	14	1,398,000	0.47
Subtotal G&A Labour Force	50	5,489,000	1.85
G&A Expense
General Office Expense	-	250,000	0.08
Computer Supplies including Software	-	100,000	0.03
Communications	-	500,000	0.17
Travel	-	150,000	0.05
Audit	-	100,000	0.03
Consulting/External Assays	-	200,000	0.07
Head Office Allowance: Marketing	-	200,000	0.07
Environmental	-	600,000	0.20
Insurance	-	1,500,000	0.50
Regional Taxes and Licenses Allowance	-	300,000	0.10
Legal Services	-	100,000	0.03
Warehouse	-	200,000	0.07

table continues…

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	Labour Force	Total Cost ($/a)	Unit Cost ($/t milled)
Entertainment/Membership	-	50,000	0.02
Recruiting	-	100,000	0.03
Medicals and First Aid	-	100,000	0.03
Relocation Expense	-	50,000	0.02
Training/Safety	-	200,000	0.07
Accommodation/Camp Cost	-	7,745,000	2.61
Crew Transportation (flight and bus)		7,376,000	2.48
Liaison Committee/Sustainability	-	500,000	0.17
Small Vehicles	-	200,000	0.07
Others	-	500,000	0.17
Subtotal G&A Expense	-	21,021,000	7.07
Total	50	26,510,000	8.92

The surface service costs are estimated at $3.48/t milled and are detailed in Table 21.14. The costs include:

• labour costs for surface service personnel and maintenance workshop

• portable water and waste management

• general maintenance including yards, roads, fences, and building maintenance

• electrical power for site services, including lighting

• building heating

• avalanche control

• airport maintenance

• water treatment

• off-site operation expense including office heating and electricity.

Table 21.14 Surface Services Operating Costs

Surface Service	Labour Force	Total Cost ($/a)	Unit Cost ($/t milled)
Surface Service Personnel	29	2,355,000	0.79
Potable Water and Waste Management	-	230,000	0.08
Water Treatment Plant	-	1,672,000	0.56
General Supplies	-	200,000	0.07
Building Maintenance	-	320,000	0.11
Building Heating – Mine Site	-	200,000	0.07
Electrical Power – Excluding WTPs	-	2,821,000	0.95
Road Maintenance	-	690,000	0.23

table continues…

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Surface Service	Labour	Total Cost	Unit Cost
Avalanche Control	-	350,000	0.12
Airport Maintenance	-	400,000	0.13
Others – Mobile Equipment	-	300,000	0.10
Off-Site Operation Expense	-	800,000	0.27
Total	29	10,338,000	3.48

21.5.5 TAILINGS STORAGE FACILITY COST

The LOM operating costs for the tailings pond construction, including lining, are estimated to be $112.8 million. The costs for tailings delivery from the plant to the TSF and reclaim water pumping are included into the process operating costs.

21.5.6 ROAD TOLL COST

Tetra Tech has not estimated the road toll cost that NovaCopper will pay to use the AMDIAR proposed to be built by the Government of Alaska. Since this cost is determined by confidential negotiations between NovaCopper and the AIDEA, a State-owned private corporation, Tetra Tech has relied on NovaCopper management to provide the road toll cost.

There is currently no developed surface access to the Arctic Project area. Access to the Arctic Project is proposed to be via a road approximately 340 km (211 miles) long, extending west from the Dalton Highway where it would connect with the proposed Arctic Project area. Although the capital costs of the road are not yet known, NovaCopper has been in discussions with AIDEA. The working assumption of this PEA study is that AIDEA would arrange financing in the form of a public-private partnership to construct and arrange for the construction and maintenance of the access road. AIDEA would charge a toll to multiple mining and industrial users (including NovaCopper’s Arctic Project) in order to pay back the costs of financing the AMDIAR. This model is very similar to what AIDEA undertook when the DeLong Mountain Transportation System (also known as the Red Dog Mine Road and Port facilities) were constructed in the 1980s. The amount paid in tolls by any user will be affected by the cost of the road, its financing structure, and the number of mines that would use the AMDIAR to ship concentrates to a port in Alaska. For the purposes of this PEA study, it has been assumed that a toll would be based on a $150 million 30-year bond at a 5% interest rate, which would result in the Arctic Project paying approximately $9.7 million each year for its 12-year mine life. The toll payments are assumed in the PEA to commence when the Arctic Project has reached commercial production.

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22.0 ECONOMICANALYSIS

A PEA should not be considered to be a prefeasibility or feasibility study, as the economics and technical viability of the Arctic Project have not been demonstrated at this time. The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Furthermore, there is no certainty that the conclusions or results as reported in the PEA will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Tetra Tech prepared an economic evaluation of the Arctic Project based on a pre-tax financial model. The NPV was estimated at the beginning of the two-year construction period.

As of May 30, 2013, the long-term metal price assumptions used in the economic analysis are as follows:

• copper: $2.90/lb

• lead: $0.90/lb

• zinc: $0.85/lb

• gold: $1,300.00/oz

• silver: $22.70/oz.

The pre-tax financial results are:

• 22.8% IRR

• 4.6-year payback on the $717.7 million initial capital costs

• $927.7 million NPV at an 8% discount rate.

NovaCopper engaged the Canadian firm Ernst & Young LLP (EY) in Vancouver, BC to prepare tax calculations for use in the post-tax economic evaluation of the Arctic Project with the inclusion of US Federal and Alaska State income taxes, and the Alaska Mining License tax (Section 22.4).

The following post-tax financial results were calculated:

• 17.9% IRR

• 5.0-year payback on the $717.7 million initial capital costs

• $537.2 million NPV at an 8% discount rate.

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Sensitivity analyses were conducted to analyze the sensitivity of the Arctic Project merit measures (NPV, IRR and payback periods) to the main inputs.

22.1	PRE-TAXMODEL
22.1.1	MINE/METALPRODUCTION INFINANCIALMODEL

The life-of-project average material tonnages, grades and metal production are shown in Table 22.1.

Table 22.1Mine/Metal Production from the Arctic Mine

Description	Value
Total Tonnes to Mill ('000)	35,681
Average Annual Tonnes to Mill ('000)	2,973
Total Waste Tonnes Mined ('000)	299,354
Stripping Ratio (LOM)	8.39
Stripping Ratio (excluding pre-stripping)	7.94
LOM (years)	12
Average Head Grade
Copper (%)	2.28
Lead (%)	0.53
Zinc (%)	3.13
Gold (g/t)	0.50
Silver (g/t)	36.91
Total Production
Copper ('000'lb)	1,560,829
Lead ('000'lb)	309,006
Zinc ('000'lb)	2,134,509
Gold ('000'oz)	369
Silver ('000'oz)	34,023
Average Annual Production
Copper ('000'lb)	130,069
Lead ('000'lb)	25,751
Zinc ('000'lb)	177,876
Gold ('000'oz)	31
Silver ('000'oz)	2,835

22.1.2 BASIS OFFINANCIALEVALUATION

The production schedule has been incorporated into the 100% equity pre-tax financial model to develop annual recovered metal production from the relationships of tonnage processed, head grades, and recoveries.

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Project, Ambler Mining District, Northwest Alaska

Payable metal values were calculated using base case metal prices. Net invoice value was calculated each year by subtracting the applicable refining and smelting charges from the payable metal value. At-mine revenues are then estimated by subtracting transportation and insurance costs. Annual operating costs for mining, processing, site services, G&A and others were deducted from the revenues to derive the annual operating cash flows.

Initial and sustaining capital costs as well as working capital have been incorporated on a year-by-year basis over the LOM. Mine reclamation costs are also applied to the capital expenditure. Capital expenditures are then deducted from the operating cash flow to determine the net cash flow before taxes.

Initial capital expenditures include costs accumulated prior to first production of concentrates and sustaining capital includes expenditures for mining and processing additions, replacement of equipment, and tailings embankment construction.

The pre-production construction period is assumed to be two years. The NPV of the Project was calculated at the beginning of this two-year construction period.

Working capital is assumed to be three months of the annual on-site operating cost and fluctuates from year to year based on the annual operating cost. The working capital is recovered at the end of the LOM.

Tetra Tech has not estimated the road toll cost that NovaCopper will pay to use the AMDIAR proposed to be built by the Government of Alaska. Since this cost is determined by confidential negotiations between NovaCopper and the AIDEA, a State-owned private corporation, Tetra Tech has relied on NovaCopper management to provide the road toll cost. For the purposes of this PEA study, it has been assumed that a toll would be paid based on a $150 million 30-year bond at a 5% interest rate, which would result in the Arctic Project paying approximately $9.7 million each year for its 12-year mine life. The toll payments are assumed in the PEA to commence when the Arctic Project has reached commercial production. Based on these assumptions, the total LOM road toll cost paid by NovaCopper cost is $116.4 million which has been applied in the economic analysis as an annual operating cost

Initial and sustaining capital costs were estimated at $717.7 million and $164.4 million, respectively. Mine closure and reclamation costs were estimated at $81.6 million and will be funded by annual payments throughout the mine life.

The undiscounted annual net cash flow (NCF) and cumulative net cash flow (CNCF) are illustrated in Figure 22.1.

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Project, Ambler Mining District, Northwest Alaska

Figure 22.1 Pre-tax Undiscounted Annual and Cumulative NCF

22.2 SUMMARY OFFINANCIALRESULTS

The pre-tax financial model was established on a 100% equity basis, excluding debt financing, and loan interest charges. In addition to the base case, Tetra Tech evaluated four alternative cases by changing the copper price only and holding the rest of parameters constant. The pre-tax financial results for the base case and the alternative cases are presented in Table 22.2.

Table 22.2 Summary of Pre-tax Financial Results

Description	Base Case	Alternate Case 1	Alternate Case 2	Alternate Case 3	Alternate Case 4
Copper Price ($/lb)	2.90	2.50	2.75	3.25	3.50
Lead Price ($/lb)	0.90	0.90	0.90	0.90	0.90
Zinc Price ($/lb)	0.85	0.85	0.85	0.85	0.85
Gold Price ($/oz)	1,300.00	1,300.00	1,300.00	1,300.00	1,300.00
Silver Price ($/oz)	22.70	22.70	22.70	22.70	22.70
Recovered Metal Value ($ million)	7,870.9	7,246.5	7,636.8	8,417.2	8,807.4
Off Site Costs and Deductions ($ million)	2,169.7	2,138.7	2,158.1	2,196.8	2,216.1
On Site Operating Costs ($ million)	2,280.5	2,280.5	2,280.5	2,280.5	2,280.5
Operating Cash Flow ($ million)	3,420.7	2,827.3	3,198.2	3,939.9	4,310.8
Initial Capital Expenditure ($ million)	717.7	717.7	717.7	717.7	717.7
Sustaining Capital ($ million)	164.4	164.4	164.4	164.4	164.4
Mine Closure and Reclamation ($ million)	81.6	81.6	81.6	81.6	81.6
Total Capital Expenditure ($ million)	963.7	963.7	963.7	963.7	963.7
NCF ($ million)	2,457.0	1,863.6	2,234.5	2,976.2	3,347.1

table continues…

NovaCopper Inc.	22-4	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Description	Base Case	Alternate Case 1	Alternate Case 2	Alternate Case 3	Alternate Case 4
Discounted Cash Flow NPV ($ million) at 5%	1,352.9	963.1	1,206.7	1,694.0	1,937.7
Discounted Cash Flow NPV ($ million) at 8%	927.7	618.9	811.9	1,198.0	1,391.0
Discounted Cash Flow NPV ($ million) at 10%	710.1	443.7	610.2	943.2	1,109.7
Payback (years)	4.6	5.1	4.8	4.3	4.1
IRR (%)	22.8	18.5	21.2	26.3	28.7
Cash Cost, Net of By-product Credit ($/lb Cu payable)	0.62	0.62	0.62	0.62	0.63
Capital Cost ($/lb Cu payable)	0.64	0.64	0.64	0.64	0.64
Total Cost ($/lb Cu payable)	1.26	1.26	1.26	1.27	1.27

A PEA should not be considered to be a prefeasibility or feasibility study, as the economics and technical viability of the Arctic Project have not been demonstrated at this time. The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Furthermore, there is no certainty that the conclusions or results as reported in the PEA will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Under the NANA Agreement, NANA has the right, following a construction decision, to elect to purchase a 16% to 25% direct interest in the Project or, alternatively, to receive a 15% Net Proceeds Royalty (NPR). This PEA was prepared on a 100% ownership basis and does not include the impact on NovaCopper if NANA elects to purchase an interest in the Project under the NANA Agreement or, alternatively, the impact on NovaCopper and the Project if the NPR becomes applicable. The PEA does include the 1.0% NSR to be granted to NANA under the NANA Agreement in exchange for a surface use agreement. Section 4.3.2 provides additional information on the NANA Agreement.

22.3 SENSITIVITYANALYSIS

Tetra Tech investigated the sensitivity of NPV, IRR and payback period to the key Arctic Project variables. Using the base case as a reference, each of key variables was changed between -30% and +30% at a 10% interval while holding the other variables constant. The following key variables were investigated:

• copper price

• lead price

• zinc price

• gold price

• silver price

• capital costs

• on-site operating costs

• off-site operating costs.

The Arctic Project’s pre-tax NPV, calculated at an 8% discount rate, is most sensitive to copper price and, in decreasing order, on site operating costs, off site operating costs, capital costs, zinc price, silver price, gold price and lead price (Figure 22.2).

NovaCopper Inc.	22-5	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Figure 22.2 Pre-tax NPV Sensitivity Analysis

As shown in Figure 22.3, the Arctic Project’s pre-tax IRR is most sensitive to the copper price followed by capital costs, on site operating costs, off site operating costs, zinc price, silver price, gold price and lead price.

Figure 22.3 Pre-tax IRR Sensitivity Analysis

The payback period (Figure 22.4) is sensitive to copper price followed by on site operating costs, off site operating costs and capital costs and less sensitive to the rest of variables.

NovaCopper Inc.	22-6	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Figure 22.4 Pre-tax Payback Period Sensitivity Analysis

22.4 POST-TAXFINANCIALANALYSIS

NovaCopper engaged EY in Vancouver, BC to prepare tax calculations for use in the post-tax economic evaluation of the Arctic Project with the inclusion of US Federal and Alaska State income taxes, and the Alaska Mining License tax.

The following tax regime was provided by EY and recognized as applicable at the time of report writing. Taxes deducted in the post-tax analysis include US Federal Tax, Alaska State Tax (AST) and Alaska Mining License Tax (AMLT). Taxes are calculated based on currently enacted United States and State of Alaska tax laws and regulations.

22.4.1 US FEDERALTAX

US Federal taxes are determined annually as the higher of “regular” corporate income tax and Alternative Minimum Tax (AMT).

Regular US Federal tax is 35% of taxable income, where taxable income is calculated as revenues less allowable costs. In addition to other allowable costs, Alaska State Income, AMLT, tax depreciation and the greater of the cost depletion or the percentage depletion can be deducted. Depreciation is calculated based on actual capital costs using the applicable tax methods allowable. Cost depletion is based on sales of minerals. Percentage depletion is the lower of a specified percentage of net metal revenues (gross revenues net of royalties) and 50% of net taxable income from mining activities before depletion. Relevant specified percentages are 15% for gold, silver and copper, and 22% for lead and zinc. Losses in a given year may be carried back two years and carried forward for 20 years, except losses incurred as a result of the reclamation costs, which may be carried back 10 years.

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Project, Ambler Mining District, Northwest Alaska

AMT is calculated as 20% of AMT income, which is regular taxable income plus certain adjustments for AMT “preferences”. Depreciation for AMT purposes is calculated at lower initial rates using the same asset class life as for regular tax. AMT percentage depletion in each year is limited to the tax basis of the mineral property. The excess of AMT over regular tax is pooled and credited against the future regular tax in excess of AMT.

For the purposes of this report, as a stand-alone project, it was assumed that the initial adjusted cost base of the depletable and depreciable property was zero and that the initial loss carry-forwards were zero. NovaCopper US Inc. has over $100 million of non-capital losses accumulated to date. Some of these losses may be utilized in the future to reduce NovaCopper’s taxable income from the Arctic Project. Further analysis would be required to determine the actual impact on taxes.

22.4.2 ALASKASTATETAX

AST is determined on a basis similar to US federal tax, with taxes calculated under both a “regular” corporate tax regime and an Alaska AMT. Regular AST is calculated as 9.4 % of taxable income, where taxable income is calculated on the same basis as US federal tax (except that state tax is not deductible). Alaska AMT is calculated as 18% of federal AMT.

22.4.3 ALASKAMININGLICENSETAX

AMLT is an income-based tax, calculated as 7% of the mining income calculated for regular AST purposes, before any deduction of AMLT, except the percentage depletion is the lower of 15% of net metal revenues and 50% of net income before depletion. No loss carry-forwards or carry-backs are applied when calculating income subject to AMLT. No AMLT tax is charged for the first 3.5 years following commencement of production. In each year, AMLT can be reduced by up to 50% through the application of “Exploration Incentive Credits” (EICs).

For the purposes of this report, as a stand-alone project, it was assumed that the initial EIC balance is zero even though the Company has a history of exploration at the Project. It was also assumed that no EICs would be earned over the life of the Arctic Project.

22.4.4 TAXES ANDPOST-TAXRESULTS

At the base case metal prices used for this study, the total estimated taxes payable on the Arctic Project profits are $811.6 million over the 12-year mine life. The components of the various taxes that will be payable for the base case and the alternate copper price cases are shown in Table 22.3.

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Project, Ambler Mining District, Northwest Alaska

Table 22.3Components of the Various Taxes

Tax Component	Base Case	Alternate Case 1	Alternate Case 2	Alternate Case 3	Alternate Case 4
Copper Price ($/lb)	2.90	2.50	2.75	3.25	3.50
Lead Price ($/lb)	0.90	0.90	0.90	0.90	0.90
Zinc Price ($/lb)	0.85	0.85	0.85	0.85	0.85
Gold Price ($/oz)	1,300	1,300	1,300	1,300	1,300
Silver Price ($/oz)	22.70	22.70	22.70	22.70	22.70
Alaska Mining License Tax ($ million)	115.4	87.9	105.1	139.4	156.6
Alaska State Income Tax ($ million)	158.0	126.2	146.1	189.0	213.9
Federal Income Tax ($ million)	538.2	429.9	497.6	640.7	724.7
Total Taxes ($ million)	811.6	644.0	748.7	969.1	1,095.2

The base case and the alternate cases post-tax financial results are summarized in Table 22.4.

Table 22.4 Summary of Post-tax Financial Results

Description	Base Case	Alternate Case 1	Alternate Case 2	Alternate Case 3	Alternate Case 4
Copper Price ($/lb)	2.90	2.50	2.75	3.25	3.50
Lead Price ($/lb)	0.90	0.90	0.90	0.90	0.90
Zinc Price ($/lb)	0.85	0.85	0.85	0.85	0.85
Gold Price ($/oz)	1,300	1,300	1,300	1,300	1,300
Silver Price ($/oz)	22.70	22.70	22.70	22.70	22.70
NCF ($ million)	1,645.5	1,219.6	1,485.8	2,007.1	2,251.9
Discounted Cash Flow NPV ($ million) at 8%	537.2	312.4	453.0	727.6	857.3
Payback (years from start of mill operations)	5.0	5.6	5.2	4.6	4.4
IRR (%)	17.9	14.0	16.5	20.9	22.9

22.5 ROYALTIES

Tetra Tech has relied on NovaCopper on the royalties applicable to the Arctic project used in the economic analysis and the related text below.

As contemplated in the NANA Agreement, a surface use agreement will be executed with NANA upon a mine construction decision for NovaCopper to use NANA lands for access and for mining facilities such as, but not limited to, buildings, tailings ponds, waste dumps, structure landing areas and laydown yards. In exchange for the surface use agreement NovaCopper will grant NANA a 1% NSR royalty on the Arctic Project. NANA’s 1% royalty was incorporated into the Project’s economic analysis.

In addition, under the NANA Agreement, NANA has the right, following a construction decision, to elect to purchase a 16% to 25% direct interest in the Project or, alternatively, to receive a 15% NPR. This PEA was prepared on a 100% ownership basis and does not include the impact on NovaCopper if NANA elects to purchase an interest in the Project under the NANA Agreement or, alternatively, the impact on NovaCopper and the Project if the NPR becomes applicable. Section 4.3.2 provides additional information on the NANA Agreement.

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Project, Ambler Mining District, Northwest Alaska

There is an existing 1% NSR royalty on the Arctic Project which is held by a third party. This royalty was granted when NovaGold purchased a 100% interest in the Ambler lands, which include the Arctic Deposit. NovaCopper, which is the owner of the Property, can purchase back this royalty, at any time, for a one-time payment of $10.0 million. More information on this royalty can be found in NovaCopper’s 2012 annual financial statements. For the purposes of the PEA, given the likelihood of NovaCopper repurchasing this royalty before the commencement of mine construction, this royalty wasnot incorporated into the economic analysis.

22.6 SMELTERTERMS

Typical smelter terms have been applied for the delivery of copper, lead and zinc concentrates to an East Asian smelter based on Tetra Tech’s in-house database of similar projects.

Copper concentrate contracts will generally include payment terms as follows:

• copper – pay 100% of content less 1.0 unit at the London Metal Exchange (LME) price for Grade A copper less a refining charge of $0.07 per accountable pound

• treatment charge – $70/dmt of concentrate delivered

• gold credit – pay 95% of content, refining charge is $10/payable oz

• silver credit – pay 90% of content, refining charge is $0.60/payable oz

• penalty charge – $0.02 per each ppm above 5 ppm mercury; $1.50 per each 1% above 1% lead; $1.50 per each 1% above 3% zinc.

Lead concentrate contracts will generally include payment terms as follows:

• lead – pay 100% of content less 3.0 units at the LME price

• treatment charge – $180/dmt of concentrate delivered

• gold credit – pay 95% of content, refining charge is $10/payable oz

• silver credit – pay 90% of content, refining charge is $0.60/payable oz

• penalty charge – $1.50 per each 0.1% above 0.2% arsenic; $1.50 per each 0.1% above 0.1% antimony; $1.00 per each 1% above 8% zinc.

Zinc concentrate contracts will generally include payment terms as follows:

• zinc – pay 100% of content less 8 units at the LME price

• treatment charge – $260/dmt of concentrate delivered

• gold credit – none

• silver credit – none

• penalty charge – none.

NovaCopper Inc.	22-10	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

22.7 TRANSPORTATIONLOGISTICS

Transportation costs for the concentrates are as follows:

•	trucking – $64.98/wmt
•	port storage and handling – $20.00/wmt
•	ocean transport to Asian port – $63.00/wmt
•	moisture content – 9%
•	concentrate losses due to the different handling operations is assumed to be
	0.42% of concentrate weight.

22.8 INSURANCE

An insurance rate of 0.15% was applied to the provisional invoice value of the concentrates.

22.9 REPRESENTATION ANDMARKETING

An allowance of $2.50/t of concentrate was applied in the economic analysis to account for the representation and marketing.

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Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

23.0 ADJACENTPROPERTIES

Adjacent to NovaCopper’s land holdings, which encompass the Arctic Deposit and a number of VMS prospects of the Ambler Sequence, are two geologically similar deposits: the Sun Deposit owned by Andover Mining Corp. (Andover) and the Smucker Deposit owned by Teck Resources Ltd. (Teck). Both prospects are located in the Ambler Schist belt (Figure 23.1). Neither contains a current mineral resource estimate.

Figure 23.1 Adjacent Properties and Land Status

23.1 SUN DEPOSIT

Andover has a 100% interest (subject to 1.5% production royalty) in the Sun Property, also referred to as the Hot prospect, located in the Ambler Schist Belt, roughly 56 km (35 miles) east of NovaCopper’s Arctic Deposit. The Sun Property consists of a total of 230 claims, representing a land position of 36,800 acres.

The Sun Property includes silver-copper-lead-zinc-gold mineralization on the Main Sun Deposit, S.W. Sun Deposit, and a number of other prospects (Figure 23.2). In total, 96 drill holes totalling 19,285 m have been completed on the Sun Property. Andover completed 47 holes during 2007, 2011 and 2012, with 49 drill holes completed by previous operators Anaconda, Noranda, Cominco and Bear Creek. A historical prefeasibility study on the Sun Deposit conducted for Anaconda Copper Mining Company in 1977 by Canadian Mine Services Ltd., Kilborn Engineering (B.C.) Ltd., and Swan Wooster Engineering Co. Ltd, projected the following resources shown in Table 23.1.

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Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Andover maintains a 20 person camp at the Sun project along with a 1,500 ft airstrip built in 2007. The camp consists of living quarters, core-logging facilities, geological office, mess facility, showers, laundry facilities, generator and tool storage, and indoor and outdoor core storage (Figure 23.2).

The historic resource estimates are considered relevant but not necessarily reliable. In each case, a significant amount of resampling and verification is needed to upgrade or verify the historical estimate as current mineral resources or mineral reserves. Please note that a QP has not done sufficient work to classify any of the the historical estimates as current mineral resources or mineral reserves and the issuer is not treating the historical estimates as current mineral resources or mineral reserves.

This historical estimate is not catgorized, and it is not known what the equivalent CIM category would be.

Table 23.1 Sun Deposit Historical Resources

	Resource (t)	Status	Ag (oz/t)	Cu (%)	Pb (%)	Zn (%)
Open Pit	2,399,000	Historical	2.39	1.93	1.20	4.51
Underground	17,891,000	Historical	2.37	1.91	1.18	4.46

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Project, Ambler Mining District, Northwest Alaska

Figure 23.2 Sun Project Prospect Location Map

23.2 SMUCKER DEPOSIT

Teck owns a 100% interest in the Smucker Property, located 40 km (24 miles) west-northwest of the Arctic Deposit in the same terrane and lithological sequence as the Arctic Deposit. Like the Arctic Deposit, the Smucker Deposit is described as a polymetallic copper-lead-zinc-gold-silver VMS prospect. Currently in target delineation stage, the Smucker Property does not have a current NI 43-101 compliant resource estimate.

Significant drilling by Anaconda in the 1970s intersected precious metal-rich VMS mineralization analogous to the other prospects of the Ambler Sequence (Ambler Schist belt). Table 23.2 shows the reported historical resource estimates for the Smucker Deposit.

The historic resource estimates are considered relevant but not necessarily reliable. In each case, a significant amount of resampling and verification is needed to upgrade or verify the historical estimate as current mineral resources or mineral reserves. Please note that a QP has not done sufficient work to classify any of the the historical estimates as current mineral resources or mineral reserves and the issuer is not treating the historical estimates as current mineral resources or mineral reserves.

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This historical estimate is not catgorized, and it is not known what the equivalent CIM category would be.

Table 23.2 Smucker Deposit Historical Resources

Deposit	Resource (Mt)	Status	Cu (%)	Zn (%)	Pb (%)	Ag (g/t)	Au (g/t)
Smucker	7.2	Historical	0.5	4.9	1.7	156.0	1.1

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Project, Ambler Mining District, Northwest Alaska

24.0 OTHER RELEVANTDATAANDINFORMATION

24.1 BORNITE DEPOSIT

In addition to the Arctic Deposit, NovaCopper, through its agreement with NANA, is actively exploring the Bornite Deposit in the Bornite Carbonate Sequence roughly 17 km southwest of the Arctic Deposit (Figure 23.1). Tetra tech has been unable to verify the information and that the information is not necessarily indicative of the mineralization on the Property that is the subject of the technical report.

In 2013, NovaCopper released an updated NI 43-101 technical report on the resource estimation of the South Reef and Ruby Creek zones of the Bornite Deposit. Results of that estimation are tabulated in Table 24.1.

A total of 217 drill holes, totaling 70,003 m, are considered to be targeted in proximity to the Bornite Deposits. These holes have been completed during 20 different annual campaigns dating from 1957 through 2012. Three underground exploration programs including 51 drill holes and targeting the No.1 Ore Body at Ruby Creek were completed between 1966 and 1968. All of the drill campaigns, with the exception of the 2011 NovaGold campaign and the 2012 NovaCopper campaign, 14 and 22 drill holes, respectively, were completed by Kennecott.

Table 24.1 Bornite NI 43-101 Resources

Deposit	Resource (Mt)	Cut- off (%)	Class	Cu (%)	Cu (Mlb)
Bornite – Ruby Creek Zone	6.8	0.5	Indicated	1.19	179
Bornite – Ruby Creek Zone	47.7	0.5	Inferred	0.84	883
Bornite – South Reef Zone	43.1	1.0	Inferred	2.54	2,409

Note: Mineral resources that are not mineral reserves do not have demonstrated economic viability. Inferred resources have a great amount of uncertainty as to their existence and whether they can be mined legally or economically. It cannot be assumed that all or any part of the Inferred resources will ever be upgraded to a higher category.

At Bornite, mineralization is hosted in deformed Devonian clastic and carbonate sedimentary rocks lying immediately south of the Ambler Schist Belt rocks across the Ambler lowlands. Widespread hydrothermal dolomitization is characteristic of the belt and hosts the associated copper mineralization. Copper mineralization at Bornite is comprised of chalcopyrite, bornite, and chalcocite as stringers, veinlets, and breccia fillings distributed in stacked, roughly stratiform zones exploiting favourable stratigraphy.

In 2013, NovaCopper expects to continue expansion of the Bornite deposit. The 2013 program included four drill rigs and approximately 8,200 m of drilling targeted at expanding the Ruby Creek and South Reef resources.

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Project, Ambler Mining District, Northwest Alaska

25.0 INTERPRETATIONANDCONCLUSIONS

25.1 GENERAL

Tetra Tech has prepared a PEA for the Arctic Project and a technical report prepared in accordance with NI 43-101 and Form 43-101F1. The PEA describes the potential technical and economic viability of establishing a conventional open pit mine, a mill complex and related infrastructure to process massive and semi-massive copper-zinc-lead-silver-gold mineralization from the Arctic Deposit. Based on the work carried out in this PEA and the resultant economic evaluation, this study should be followed by further technical and economic studies leading to a prefeasibility study.

25.2 PROPERTY DESCRIPTION AND LOCATION

Information from NovaCopper supports that the mining tenure held is valid and is sufficient to support declaration of mineral resources. Tenures have been surveyed in accordance with appropriate regulatory requirements. Annual claim-holding fees have been paid to the relevant regulatory authority and all applicable permits for current work activities are valid. There are no known environmental liabilities due to previous operators or ongoing NovaCopper exploration activities at the Property.

25.3 GEOLOGY

The Arctic Deposit is considered to be an example of a polymetallic (copper-zinc-lead-silver-gold) VMS-style occurrence, common within the Ambler Sequence stratigraphy of the Ambler District. Knowledge of the deposit settings, lithologies, structural and alteration controls on mineralization and SG, and the mineralization style and setting is sufficient to support mineral resource estimation.

The exploration and drilling programs completed to date are appropriate to the type of the deposit. The exploration, drilling, geological modelling and research work supports ore genesis interpretations.

Tetra Tech has reviewed the quantity and quality of the data collected by NovaCopper, and previously by NovaGold, and found it reliable and suitable for the purpose of this study. Review of the QA/QC dataset and reports found NovaCopper’s sample preparation, analysis and security procedures to meet or exceed industry best practices.

It is Tetra Tech’s opinion that the drill database for the Arctic Deposit is reliable and sufficient to support the purpose of this technical report and a current mineral resource estimate.

NovaCopper Inc.	25-1	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

25.4 MINERAL RESOURCE ESTIMATION

The mineral resource estimate, as seen in Table 25.1 and which formed the basis of the PEA, was completed by Mr. Michael F. O’Brien, M.Sc., Pr.Sci.Nat, FGSSA, FAusIMM, FSAIMM and an independent QP as set forth by NI 43-101. The overall effective date of this resource estimate is July 30, 2013. Estimations of mineral resources for the Arctic Project conform to industry best practices and meet requirements of CIM (2010).

Table 25.1Resource estimate for the Arctic Project (NSR Cut-off of $35/t)

Category	Mt	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)	Cu (Mlb)	Zn (Mlb)	Pb (Mlb)	Au (Moz)	Ag (Moz)
Indicated	23.848	3.26	4.45	0.76	0.71	53.2	1,713	2,338	400.9	0.55	40.8
Inferred	3.363	3.22	3.84	0.58	0.59	41.5	239	285	43.2	0.06	4.5

Notes:	1.	These resource estimates have been prepared in accordance with NI 43-101 and the CIM Definition Standards. Mineral resources that are not mineral reserves do not have demonstrated economic viability. Inferred resources have a great amount of uncertainty as to their existence and whether they can be mined legally or economically. It cannot be assumed that all or any part of the Inferred resources will ever be upgraded to a higher category.
	2.	Mineral Resources are reported within mineralization wireframes contained within an Indicated and Inferred pit design using an assumed copper price of US$2.90/lb, zinc price of US$0.85/lb, lead price of US$0.90/lb, silver price of US$22.70/oz, and gold price of US$1,300/oz.
	3.	Appropriate mining costs, processing costs, metal recoveries and inter ramp pit slope angles were used to generate the pit design.
	4.	The $35.01/t milled cut-off is calculated based on a process operating cost of $19.03/t, G&A of $7.22/t and site services of $8.76/t. NSR equals payable metal values, based on the metal prices outlined in Note 2 above, less applicable treatment, smelting, refining costs, penalties, concentrate transportation costs, insurance and losses and royalties.
	5.	The LOM strip ratio is 8.39.
	6.	Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content.
	7.	Tonnage and grade measurements are in metric units. Contained copper, zinc and lead pounds are reported as imperial pounds, contained silver and gold ounces as troy ounces.

Factors which may affect the Mineral Resource estimate include advancements/changes to the geological, geotechnical and geometallurgical models, infill drilling to convert mineral resources to a higher classification, extension drilling to test for deep extensions to known resources, and collection of additional SG data. Additional factors which may affect the open pit extraction scenario used to constrain the estimates are commodity price and exchange rate assumptions, assumptions used to estimate metallurgical recoveries, pit slope angles, and variations in the planned infrastructure layout. It should be noted that all factors pose potential risks and opportunities to the current mineral resource. Recommendations to address these risks and opportunities are provided in Section 26.0.

25.5 METALLURGICAL TEST WORK AND PROCESS DESIGN

Metallurgical test work and associated analytical procedures were performed by recognized testing facilities, and the tests performed were appropriate to the mineralization type. The samples selected for the 2012 testing were representative of the various types and styles of mineralization present at the Arctic Deposit. The results and conclusions of this study are mainly derived from the 2012 test work.

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In general, the 2012 test results indicated that the samples responded well to the tested flowsheet consisting of talc pre-flotation, conventional copper-lead bulk flotation and zinc flotation, followed by copper and lead separation. The 2012 test work showed that the mineralization from different mineralization zones have slightly different requirements in terms of flotation reagents. These requirements are driven primarily by changes in base metal ratios in the process feed as well as the overall sulphide content of tested samples. According to the test work results, the concentrates produced are expected to be saleable, especially the high-grade copper and zinc concentrates. Further tests are required to optimize the process conditions and flowsheet, especially copper and lead separation, talc removal and test condition optimization for Zone 1 & 2 mineralization.

The grindability test results showed that the mineralized samples are neither resistant nor abrasive to ball mill grinding. The materials are considered to be soft or very soft in terms of grinding requirements.

A 10,000 t/d process plant has been designed for the Arctic Project to process massive/semi-massive sulphide mineralization for the production of separate copper, lead and zinc concentrates. The proposed process plant will consist of one stage of crushing, primary grinding in a SAG mill/ball mill (SAB) grinding circuit, talc pre-flotation, copper-lead rougher and cleaner flotation, zinc rougher and cleaner flotation, and copper and lead separation flotation. The operating philosophy and selected equipment is in line with industry expectations.

Table 25.2 summarizes estimated recoveries and concentrate grades based on the projected mill feed grades.

Further plant design optimization is required to incorporate site conditions and updated test work results to maximize project economics.

Table 25.2 Projected Recoveries and Concentrate Grades

Head Grades (based on proposed mine plan)
Copper	2.28%
Zinc	3.13%
Lead	0.53%
Silver	37 g/t
Gold	0.50 g/t
Metal Recoveries
To Copper Concentrate
Copper	87.1%
Silver	40.2%
Gold	57.9%
To Zinc Concentrate
Zinc	86.8%
To Lead Concentrate
Lead	74.0%

table continues…

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Silver	40.2%
Gold	6.8%
Concentrate Grades
Copper Concentrate	29.0% Copper
Zinc Concentrate	56.0% Zinc
Lead Concentrate	50.0% Lead

25.6 MINING METHODS

The Arctic Project’s total mine life is 13 years including 1 year of pre-stripping followed by 12 years of production.

The mine will utilize a conventional truck-and-shovel fleet and the proposed open pit operations will provide process plant feed at a nominal rate of 10,000 t/d. The pit design incorporates a bench height of 5 m and an inter-ramp angle of 45°. Pit optimization and design results are based on scoping level geotechnical and operational information.

Due to operational and geometric constraints, the mine will require a ramp-up period of three years before achieving full mill capacity. The mining schedule does not currently consider a low-grade stockpiling option but this can be assessed in more detail in future studies.

Factors which may affect the mine plan include advancements/changes to the geotechnical, hydrological and ML/ARD/SG waste characterization models. Additional factors which may affect the open pit extraction scenario are commodity price assumptions, assumptions used to estimate metallurgical recoveries, pit slope angles, and variations in the planned infrastructure layout. Trade-off studies to optimize the productions schedule, including further work on evaluating equipment size (particularly for the waste material) may prove beneficial.

It should be noted that all factors pose potential risks and opportunities to the current mine plan. Recommendations to address these risks and opportunities are provided in Section 26.0.

25.7 PROJECT INFRASTRUCTURE

The Arctic Project infrastructure includes building locations which are based on the assumption that the geotechnical data and the geohazards are favourable, however further investigation is required which could affect planned infrastructure layout.

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25.8 TAILINGS AND WATER MANAGEMENT

The work completed to date suggests that the current tailings and water management concept is practicable and should be carried forward to the next level of design.

The tailings and water management concept presented in this study has the potential to be optimized. The surficial geology and geotechnical conditions should be investigated and the design basis and operating criteria should be further refined.

25.9 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

The Arctic Project will be subject to a mine permitting process typical for a mine of its size in Alaska. NovaCopper will need to apply for appropriate mining operations-related permits under NWAB, state and federal laws to allow proposed mining operations. Exploration activities completed to date have been conducted under the relevant permits.

In order to support this process, NovaCopper will have to broaden their existing baseline environmental program and complete a number of additional studies that will support the permit applications and mine plan. NovaCopper has signed a cooperative agreement with NANA, which formalized an Oversight Committee and a Subsistence Committee, a Communications Subcommittee and a Workforce Development Subcommittee. NovaCopper is formally engaging the Arctic Project stakeholders in a deliberate effort to make the Project directly beneficial to them throughout the life of the project. NovaCopper will be required to develop a mine plan that is protective of the environment during mining operations as well as a reclamation and closure plan that ensures the environment is protected well after mine closure. There are no characteristics of the Arctic Project from an environmental, permitting or community impact perspective that should prevent NovaCopper from advancing this project to the next stage.

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25.10 ECONOMIC ANALYSIS

Tetra Tech prepared an economic evaluation of the Arctic Project based on a pre-tax financial model. NPV was estimated at the beginning of the two-year construction period. NovaCopper engaged the Canadian firm of EY in Vancouver, BC to prepare tax calculations for use in the post-tax economic evaluation of the Arctic Project with the inclusion of US federal and Alaska income taxes and Alaska Mining License tax. Economic merit measures such as NPV, IRR and payback period were estimated on both pre-tax and post-tax basis for the base case and four alternate metal price cases. Sensitivity analyses were conducted to analyze the sensitivity of the Arctic Project merit measures (NPV, IRR and payback periods) to the main inputs. Using the base case as a reference, each of key variables was changed between -30% and +30% at a 10% interval while holding the other variables constant. The Arctic Project’s pre-tax NPV, calculated at an 8% discount rate, was found to be most sensitive to copper price and, in decreasing order, on site operating costs, off site operating costs, capital costs, zinc price, silver price, gold price and lead price. The pre-tax IRR was found to be most sensitive to the copper price followed by capital costs, on site operating costs, off site operating costs, zinc price, silver price, gold price and lead price. The payback period was found to be sensitive to copper price followed by on site operating costs, off site operating costs and capital costs and less sensitive to the rest of variables. Using base case long-term metal prices of $2.90/lb for copper, $0.85/lb for zinc, $0.90 for lead, $22.70/oz for silver and $1,300/oz for gold, the pre-tax NPV at an 8% discount rate was $927.7 million calculated at the beginning of the two-year construction period and IRR was 22.8%. After-tax NPV was $537.2 million and after-tax IRR was 17.9%. The estimated payback period was 4.6 years before-tax and 5.0 years after-tax.

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26.0 RECOMMENDATIONS

26.1 GENERAL

Tetra Tech recommends the following actions and activities in support of advancing the Arctic Project should NovaCopper determine that a prefeasibility study is warranted. As part of the recommended work program, additional exploration and geotechnical drilling will be required to support the following studies/permits: geotechnical studies, metallurgical studies, engineering studies, waste-rock characterization studies, additional baseline studies and environmental permitting activities, marketing studies and numerous trade-off studies. Additional areas for work are also likely to be identified as activities progress. The total estimated direct program costs are approximately $6 million to $6.5 million, which excludes site costs such as camp support, overhead and other indirect costs.

26.2 GEOLOGY

Tetra Tech makes the following recommendations for future work:

• Model Integration: The geological model for the various mineralized zones (based on interpretation of geological and metal sampling data), geotechnical model, ARD model and the density model (based on SG measurements and barite occurrence) should be integrated into a single block model and correlating wireframes to provide a single repository of deposit data.

• Bulk Density Characterization: Historically, drill hole data was collected on the assumption of eventual underground exploitation. Consequently, only limited SG data was collected for the overlying rocks above the mineralized zones. The bias in the existing SG dataset in favor of the rock adjacent to the mineralized zones is likely to positively bias the estimate of the SG of the waste material to be stripped. A more representative set of SG measurements of the shallow schistose and potentially more weathered near surface material, represents an opportunity to decrease the stripping ratio for proposed open pit mining. At least 500 additional measurements, from representative intervals of the hanging wall units from existing cores in good condition, would provide better spatial representation and should be added to the database. An integrated geological model would also support better SG definition along domain boundaries.

• Diamond Drilling: Material currently classified as an Inferred resources on the deeper southern edges of the deposit should be drilled to support conversion to Indicated Resources (75 m centres). The proposed drilling would total 4,400 m and it is anticipated that approximately 700 m would require assaying.

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The estimated cost of the recommended drilling, assaying and SG characterization work is $3.52 million.

26.3 MINERAL PROCESSING AND METALLURGICAL TESTING

In general, the flowsheet developed in the 2012 test program is feasible for the Arctic Deposit mineralization; however, Tetra Tech recommends further metallurgical test work on representative samples to optimize the flowsheet and process conditions, and to determine the design-related parameters required for further studies. Tetra Tech makes the following recommendations for future work:

• Additional metallurgical test work and mineralogical evaluations should be conducted to confirm the metallurgical response of the representative samples to the established process flowsheet. The tests should focus on 1) the impact of talc on copper and lead bulk flotation, 2) zinc mineral suppression reagent regime optimization for copper and lead bulk flotation, and 3) copper and lead separation.

• Additional variability test work on samples generated from different mineralization zones with different head grades, and samples representing the initial three to five years of mill feed, should be undertaken.

• Locked-cycle test work, or a mini-pilot plant, using bulk copper-lead concentrate produced from a larger bulk sample should be undertaken. Optimization of the copper and lead separation process is warranted as it will directly affect the saleable nature of the copper and lead concentrates and limited work in this area has been completed to date.

• Process design-related parameters should be determined, including material bulk density, grindability index, concentrate settling rates and filtration rates. As well, simulations for the proposed grinding circuit should be conducted. The destruction of residual cyanide in the copper and lead concentrate slurries should be studied.

The estimated cost of the mineral processing and metallurgical test work is estimated to be $780,000.

26.4 PROCESS PLANT DESIGN

Further optimizations on plant designs including primary comminution circuits, flotation circuits, regrinding circuits and layout are recommended including incorporating updated test work results to optimize the process flowsheet. The costs associated with the optimizations will be part of the costs for the next phase of study.

The mine plan should be optimized to evaluate if increasing the mill feed rate for the initial mine life is achievable.

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26.5 MINING METHODS

Tetra Tech makes the following recommendations for future work:

• Conduct geotechnical studies to better define the appropriate design pit slope angles and to investigate the fractured area to the south-east of the proposed open pit.

• Perform a trade-off study on the optimum mill capacity.

• Perform a trade-off study to investigate low-grade stockpiling.

• Perform a trade-off study to investigate different mine fleet options.

The estimated cost for the proposed mining work will be approximately $250,000.

26.6 PROJECT INFRASTRUCTURE

Additional studies outlined below are recommended for advancing the Arctic Project infrastructure. The definition of a unified project footprint with appropriately augmented topography is recommended.

26.6.1 PROCESS PLANT AND INFRASTRUCTURE LOCATION

Additional studies for the location of the process plant and related infrastructure will be required. An investigation of the soil conditions needs to be performed in order to simplify the design of the mill building and major equipment foundations. Tetra Tech makes the following recommendations for future work:

• Drilling: A total of six drill holes (30 m deep) will be required at the primary crusher (one drill hole), mill building (two drill holes), stockpile (one drill hole), waste dump (two drill holes). The cost of the work is estimated to be $60,000. For the open pit area, a total of eight drill holes (approximately 250 m deep) will be required within the pit. The cost of the work is estimated to be $600,000.

• A geohazard evaluation study including avalanche paths and probable impact. The cost of the work is estimated to be $60,000.

• Field investigations to examine the maximum flow at freshette. The cost of the work is estimated to be $40,000.

• Ground investigation by a geotechnical engineer. The cost of the work is estimated to be $20,000.

The total estimated cost for the process plant and infrastructure location studies is $780,000.

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26.6.2 ACCESS ROAD

Further geotechnical detail and aggregate sourcing data will be required to support access road design. Tetra Tech makes the following recommendations for future work:

• Six drill holes 10 to 15 m deep at the proposed three bridges abutments - The cost of the work is estimated to be $20,000.

• Ten test pits (2 to 3 m deep) for borrow material every 2 km along the access road borrow sources. The cost of the work is estimated to be $50,000.

• Ten drill holes (10 to 15 m deep) required every 500 m for the last 5 km of the alignment to confirm the subgrade type. The cost of the work is estimated to be $120,000.

The total estimated cost for the access road test work is $190,000.

26.6.3 ELECTRICAL POWER GENERATION

In July 2013, AIDEA released a feasibility report on the Interior Energy Project (IEP). The proposed project would include a natural gas liquefaction plant located on the North Slope of Alaska near Deadhorse. From the Plant, liquefied natural gas (LNG) would be trucked almost 800 km (500 miles) south to Fairbanks, primarily along the Dalton Highway. Based on the economic and financial analysis conducted in the feasibility, the IEP is projected to be feasible, with natural gas delivered to residential and commercial customers in the Fairbanks North Star Borough at prices sufficiently below current energy costs (AIDEA 2013). Tetra Tech recommends conducting a trade-off study to investigate optimized on-site power generation by LNG compared to diesel. The cost for the study is expected to be $50,000.

26.6.4 AIRSTRIP

Tetra Tech makes the following recommendations for future work:

• Two drill holes 10 to 15 m deep required to test for permafrost. The cost of the work is estimated to be $10,000.

• Refine alignment, apron sizing/requirements. The estimated cost is $30,000.

The total estimated cost for the airstrip evaluation is $40,000.

26.6.5 TAILINGS AND WASTE MANAGEMENT

TAILINGS

Tetra Tech makes the following recommendations for future work:

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• Conduct geotechnical investigation of embankment and spillway sites to characterize potential construction and borrow materials. The cost of the work is estimated to be $60,000.

• Further define plant information including: tailings production rates (average, maximum, minimum), range of slurry solids concentrations, and required capacity of reclaim water system (mill water demand). The cost of the work is estimated to be $30,000.

• Complete additional tailings characterization studies including: tailings rheology, geochemical and mineralogical characteristics, water release and recovery rates, supernatant hydrochemistry, shear strength, dispersive tendency, average beach slope, tailings surface accessibility with time, drying and strength gain of deposited tailings, and crust development and dust generation. The cost of the work is estimated to be $80,000.

• Evaluate embankment stability to include additional data from a detailed geotechnical investigation including: seismic deformation analysis, foundation preparation requirements, optimization of embankment geometry, embankment critical section, evaluation of potential embankment raises, tailings liquefaction potential. The cost of the work is estimated to be $10,000.

• Conduct a water balance. The cost of the work is estimated to be $10,000.

• Further evaluate tailings co-disposal with waste rock to optimize operation and mitigate acid generation. The cost of the work is estimated to be $15,000.

• Complete a detailed geohazard evaluation for the tailings site. The cost of the work is estimated to be $10,000.

The total estimated cost to design the TSF at the prefeasibility level is $215,000.

WATER MANAGEMENT

Tetra Tech makes the following recommendations for future work:

• Baseline work should be conducted to support design parameters that will support the design of the water management structures. This will include detailed field data acquisition and laboratory testing programs to characterize hydrological, hydrogeological and geo-environmental conditions for the proposed project. The cost of the work is estimated to be $20,000.

• Hydrology and hydraulic analyses, including a detailed water balance, will need to be completed to determine the types, locations and sizes of the water management structures as well as the times and quantities of discharges. The cost of the work is estimated to be $10,000.

• Geochemical modelling must be carried out to establish potential acidity of surface run-off over the temporary waste dumps and pit, in order to assess the need for collecting and treating contact water or discharging it directly to existing natural streams. The cost of the work is estimated to be $135,000.

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The total estimated cost to design the storm water management plan at the prefeasibility level is $185,000.

26.7 ENVIRONMENTAL

Additional environmental baseline studies are required to advance the Arctic Project through the prefeasibility stage. Existing hydrology and aquatic life monitoring should be broadened to cover additional streams in the Arctic Project area and sampling frequency may need to be increased. In addition, the baseline program should be broadened to include additional studies such as ground water hydrology, meteorology, geochemistry – waste rock characterization, wetlands delineation and functional assessment, wildlife and subsistence surveys. An accurate water balance model is needed for developing robust mine and closure plans. Sufficient meteorological and hydrology data (in terms of frequency and duration in years) are required in developing that water balance model. It is beyond the scope of the document to make more specific recommendations. Annual costs for environmental monitoring and studies could range from less than $100,000 to more than $500,000 in certain years.

26.8 ECONOMIC ANALYSIS

A concentrate marketing study is required to investigate the marketability of the three concentrates and better define the applicable transportation, insurance, marketing, refining and smelting costs. The cost of the work is estimated to be $50,000.

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27.0 REFERENCES

27.1 GEOLOGY

Alaska Climate Summaries: Kobuk (1971-2000) http://www.wrcc.dri.edu/cgi-bin/cliMAIN.pl?akkobu.

Aleinikoff, J. N., Moore, T. E., Walter, M., and Nokleberg, W. J. (1993). U-Pb ages of zircon, monazite, and sphene from Devonian Metagranites and Metafelsites, Central Brooks Range, Alaska: U.S. Geological Survey Bulletin, v. B 2068, p. 59-70.

BD Resource Consulting (2013). Technical Report for the Bornite Property, South Reef and Ruby Creek Zones, Northwest Alaska, USA. Prepared by BD Resource Consulting Inc. for NovaCopper. Report Date February 8, 2013, 149 pages.

Bearing Marine Lines (2007). Personal Communication with Company Representative.

BGC Engineering Inc. (2012). Rock Mechanics and Hydrogeology Study, Ambler project, Arctic Deposit, SubArctic Creek, Alaska draft report. October 2012.

Brown, H. (2011). Memo to Ambler Staff “Ambler District Database Audit”: Internal NovaGold Resources Inc. Memo, May 19, 2011.

Dillon, J. T., Pessel, G. H., Chen, J. H., and Veach, N. C. (1980). Middle Paleozoicmagmatism and Orogenesis in the Brooks Range, Alaska: Geology, v. 8, p. 338-343.

Dillon, J. T., Tilton, G. R., Decker, J., and Kelley, M. J. (1987). Resource Implications of Magmatic and Metamorphic Ages for Devonian Igneous Rocks in the Brooks Range, in Tailleur, I. L., and Weimer, P., Alaskan North Slope Geology, Pacific Section, Society of Economic Paleontologists and Mineralogists, p. 713-723.

Dodd, S. P., Lindberg, P. A., Albers, D. F., Robinson, J. D., Prevost, R. (2004). Ambler Project, 2004 Summary Report, Unpublished Internal Report, Alaska Gold Company.

Gottschalk, R. R., and Oldow, J. S. (1988). Low-angle Normal Faults in the South-central Brooks Range Fold and Thrust Belt, Alaska, Geology, 16, p. 395-399.

Hitzman, M. W. (1982). Geology of the Ruby Creek Copper Deposit, Southwestern Brooks Range, Alaska: Economic Geology v. 81, p. 1644-1674.

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Hitzman, M. W., Smith, T. E., and Proffett, J. M. (1982). Bedrock Geology of the Ambler District, Southwestern Brooks Range, Alaska: Alaska Division of Geological and Geophysical Surveys Geologic Report 75, scale 1:250,000.

Hitzman, M. W., Proffett, J. M., Schmidt, J. M., and Smith, T. E. (1986). Geology and Mineralization of the Ambler District, Northwestern Alaska: Economic Geology v. 81, p. 1592-1618.

Kennecott Research Center (September 1968). Amenability Testing of Diamond Drill Core Samples from Arctic, Alaska Project, TR 68-20.

Kennecott Research Center (August 1972). Amenability Testing of Samples from Bear Creek Mining Company’s Arctic Deposit, TR 72-12.

Kennecott Research Center (September 1976). Recovery of Mineral Values Arctic Prospect, RTR 76¬22.

Kennecott (1977). Annual Report Arctic Deposit: Unpublished in-house report.

Kennecott (1998). Arctic Deposit and Ambler District Field Report: Unpublished in-house report.

Kennecott Research Center (January 1997). Process Selection for Arctic Deposit, Technical Report RTR 77-4.

Kobuk Valley National Park (2007). www.kobuk.valley.national-park.com/info.htm#env.

Lakefield Research Limited (January 7, 1999). An Investigation of the Recovery of Lead, Zinc & Precious Metals from Samples of the Arctic Project Ore submitted by Kennecott Minerals, Progress Report No.1.

Lindberg, P. A. (2004). Structural Geology of the Arctic Cu-Zn-Pb-Ag Sulfide Deposit: Alaska Gold Company Unpublished Report.

Michaelson, S. D. (March 11, 1970). Memo to H. L. Bauer.

Moore, T. E., Wallace, W. K., Bird, K. J., Karl, S. M., Mull, C. G., and Dillon, J. T. (1994). Geology of Northern Alaska, in Plafker, G., and Berg, H. C., eds., The Geology of Alaska: Boulder, Colorado, Geological Society of America, The Geology of North America, v. G1, p. 49-140.

Mull, C. G. (1982). Tectonic Evolution and Structural Style of the Brooks Range, Alaska; an Illustrated Summary, in Geologic Studies of the Cordilleran Thrust Belt, Rocky Mt. Assoc. Geol., Denver, CO, United States (USA), p. 1-45.

Mull, C. G. (1985). Cretaceous Tectonics, Depositional Cycles, and the Nanushuk Group, Brooks Range and Arctic Slope, Alaska, U.S. Geol. Soc. Bull., 1614, p. 7-36.

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NANA/DOWL Engineers and CH2M Hill (March 2005). Ambler District Access Study, Phase I Summary Route Identification and Screening Analysis, p. 51.

Nauman, C. R. (December 29, 1994). Memo Arctic: Status of Development of Ore Processing.

Oldow, J. S., Seidensticker, C. M., Phelps, J. C., Julian, F. E., Gottschalk, R. R., Boler, K. W., Handschy, J. W., and Ave Lallemant, H. G. (1987). Balanced Cross Sections Through the Central Brooks Range and North Slope, Arctic Alaska, AAPG, p. 19, 8 plates.

Otto, B. R. (2006). Personal Communication.

Proffett, J. M. (1999). Summary of Conclusions on Geology of the Arctic Deposit, AK: Kennecott Minerals Company Unpublished Report.

Randolph, M. P. (August 29, 1990). Internal Kennecott Memo to T. J. Stephenson, Arctic Deposit.

Robertson Geoconsultants Inc. (December 1998). Initial Assessment of Geochemical and Hydrological Conditions at Kennecott’s Arctic Project.

Russell, R. H. (1977). Annual Report, Arctic Deposit: Bear Creek Mining Company Unpublished Report.

Russell, R. H. (1995). Arctic Project 1995 Evaluation Report, Geologic Report: Kennecott Corporation Unpublished Report.

Sawyer, Roger J. (January 15, 1999). Memo to J. Earnshaw, Kennecott Minerals, Arctic-Metallurgy Projections.

Schmidt, J. M. (1983). Geology and Geochemistry of the Arctic Prospect, Ambler District, Alaska: Unpublished Ph.D. dissertation, Stanford University.

Schmidt, J. M. (1986). Stratigraphic Setting and Mineralogy of the Arctic Volcanogenic Massive Sulfide Prospect, Ambler District, Alaska: Economic Geology v. 81. p. 1619-1643.

Schmidt, J. M. (1988). Mineral and Whole Rock Compositions of Seawater-Dominated Hydrothermal Alteration at the Arctic Volcanogenic Massive Sulfide Prospect, Alaska: Economic Geology v.83, p. 822-842.

Shaw, Stone & Webster Management Consultants, Inc. (2006). Mine Power Study Arctic Project – Ambler Mining District, Unpublished Report for Alaska Gold Company, p. 73.

SRK (2012). NI 43-101 Preliminary Economic Assessment, Ambler Project, Kobuk, AK. Prepared by SRK Consulting (U.S.) Inc. for NovaCopper. Report Date March 9, 2012. 276 pages.

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Till, A. B., Schmidt, J. M., and Nelson, S. W. (1988). Thrust Involvement of Metamorphic Rocks, Southwestern Brooks Range, Alaska: Geology, v. 16, p. 930-933.

URSA Engineering (1998). Arctic Project Rock Mass Characterization, Prepared for: Kennecott Minerals, Co., Unpublished Report, p. 49.

Vallat, C. (2013a). Memo to NovaCopper Inc. “Historic Pre-NovaGold(2004) Arctic Assay Validations and Updates Within NovaCopper Database”: GeoSpark Consulting Inc. April 22, 2013.

Vallat, C. (2013b). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. Arctic Project, pre-NovaGold (pre-2004)”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. April 22, 2013.

Vallat, C. (2013c). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. 2004 Arctic Project”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. June 6, 2013.

Vallat, C. (2013d). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. 2005 Arctic Project”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. June 6, 2013.

Vallat, C. (2013e). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. 2006 Arctic Project”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. June 6, 2013.

Vallat, C. (2013f). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. 2007 Arctic Project”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. June 6, 2013.

Vallat, C. (2013g). “Quality Assurance and Quality Control Review on Analytical Results Related to the NovaCopper Inc. 2008 Arctic Project”: GeoSpark Consulting Inc. unpublished report for NovaCopper Inc. June 6, 2013.

Vallat, C. (2013h). Memo to NovaCopper Inc. “Arctic Projects 2011 Drill Program Assays Reported in 2013 – QAQC” GeoSpark Consulting Inc. May 31, 2013.

Vogl, J. J., Calvert, A. J., Gans, P. B. (2002). Mechanisms and Timing of Exhumation of Collision-Related Metamorphic Rocks, Southern Brooks Range, Alaska: Insights from Ar (40)/ Ar (39) Thermochronology, Tectonics, v 21, No 3, p. 1-18.

West, A. (2013). Memo to E. Workman & GeoSpark Consulting Inc. “NovaCopper Arctic Project Database Verification”: Internal NovaCopper Inc. Memo, January 8, 2013

Williams, A. (2000). Opportunities in the NANA Region, in: Mining Alaska National Interest Lands Conservation Act (ANILCA)-Twenty Years Later-Abstracts, Alaska Miners Association 2000 Annual Convention, p. 25.

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Zieg, G. A., et al. (2005). Ambler Project 2005 Progress Report, Unpublished Internal Report, Alaska Gold Company.

WEBSITE:
http://alaskamininghalloffame.org/inductees/berg.php, February 14, 2012.

27.2 ENVIRONMENTAL

ADF&G (2011). Caribou Management Report- Alaska Department of Fish and Game, Division of Wildlife Conservation, Patricia Harper Editor, 2011.

Craig, C. (2012). 2012 UKMP Water Quality Report, Internal Report by NovaCopper US Inc., January, 2012.

Dowl HKM (2013). NovaCopper Final Year-End Report 2012, Consultant Report prepared for NovaCopper under Work Order 60816, March 2013.

Shaw Alaska Inc. (2008). Hydraulics Data Report July 2008 Event Final, Shaw Environmental, Inc.

Shaw Environmental Inc. (2007). Ambler Project, 2007 Environmental Baseline Sampling, Alaska Gold, Inc.

Shaw Environmental Inc. (2008). Water Quality Report July 2008 Event Final, Shaw Environmental, Inc.

Shaw Environmental Inc. (2009). Hydraulics Data Report July 2009 Event Draft, Shaw Environmental, Inc.

Shaw Environmental Inc. (2009). Water Quality Report July 2009 Event Final, Shaw Alaska, Inc.

SRK (2012). NI 43-101 Preliminary Economic Assessment Ambler Project Kobuk, AK, SRK Consulting (U.S.), Inc., March 9, 2012.

Stephen R. Braund & Associates (2012). Ambler Mining District Access Road Subsistence Data Gap Memo, Prepared for the Alaska Department of Transportation and Public Facilities, May 2012.

Tetra Tech (2010a). Arctic Deposit Access Environmental Baseline Data Collection - Hydrology, Ambler Mining District, Alaska, December 1, 2010.

Tetra Tech (2010b). Arctic Deposit Access Environmental Baseline Data Collection - Wetlands & Vegetation, Ambler Mining District, Alaska, November 24, 2010.

Tetra Tech (2011). Arctic Deposit Access Environmental Baseline Data Collection – Aquatics, Ambler Mining District, Alaska, January 20, 2011.

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27.3 MINING

EBA, A Tetra Tech Company (2013). Arctic Project PEA – Geotechnical Open Pit Design Criteria. June 17, 2013.

27.4 CAPITAL COST ESTIMATE

Adventurine Engineering Inc., Mine and Mill Equipment Costs, An Estimators Guide (2012) published by InfoMine USA Inc., CostMine Division.

Page, John S., Estimator’s General Construction Man-Hour Manual, Second Edition, 1999, Butterworth-Hinemann, ISBN-12: 978-0-87201-320-9.

Page, John S., Estimator’s Equipment Installation Man-Hour Manual, Third Edition, 1999, Gulf Professional Publishing, ISBN-13: 978-0-88415-287-3.

Page, John S., Estimator’s Piping Man-Hour Manual, Fifth Edition, 1999, Butterworth-Heinemann/(originally published by Gulf Publishing Company), ISBN-13: 978-0-88415-259-0.

RSMeans Costworks, Reed Construciton Data, Heavy Constuciton Cost Data (2013 Cost Data) Electronic Database.

Salzere, Krist Noyes (2010), Canadian Mines Salaries, Wages & Benefits, 2010 Survey Results, InfoMine USA, Inc.

Stephen R. Braund & Associates (2012). Ambler Mining District Access Road Subsistence Data Gap Memo, Prepared for the Alaska Department of Transportation and Public Facilities, May 2012.

27.5 ECONOMIC ANALYSIS

NovaCopper Inc. Second Quarter 2013 Management’s Discussion & Analysis, May 31, 2013.http://www.novacopper.com/i/pdf/financials/2013-Q2-MDA.pdf

27.6 ELECTRICAL POWER GENERATION

Interior Energy Project Feasibility Report. Proposed Project – North Slope LNG Plant July 2013. Report prepared by the Alaska Industrial Development and Export Authority (AIDEA) and Alaska Energy Authority.

NovaCopper Inc.	27-6	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.0 CERTIFICATESOFQUALIFIEDPERSONS

28.1 MICHAELF.O’BRIEN, M.SC., PR.SCI.NAT, FGSSA, FAUSIMM,FSAIMM

I, Michael F. O’Brien, M.Sc., Pr. Sci. Nat., FGSSA, FAusIMM, FSAIMM, of Vancouver, British Columbia, do hereby certify:

• I am a Chief Geologist with Tetra Tech WEI Inc. with a business address at 800- 555 West Hastings Street, Vancouver, British Columbia, V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of the University of Natal (B.Sc Hons. Geology, 1978). I am a registered Professional Natural Scientist (Geological Scientist) in good standing of the South African Council for Natural Scientific Professions (South Africa, 400295/87). My relevant experience is 33 years of experience in operations, mineral project assessment and I have the experience relevant to Mineral Resource estimation of metal deposits. I have estimated Mineral Resources for greenstone-hosted gold (Cuiaba in Minas Gerais Brazil), diatreme complex epithermal gold deposits (Cripple Creek in Colorado), porphyry copper-gold (Pebble Mine in Alaska) and shear zone-hosted deposits (Mongbwalu in the Dominican Republic of Congo). I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”) under the Accepted Foreign Associations and Membership Designations (Appendix A).

• My most recent personal inspection of the Property was June 19, 2013 for three days.

• I am responsible for Sections 1.1 to 1.4, 1.13, 1.14, 2.0, 3.0, 4.0 to 12.0, 14.0, 23.0, 24.0, 25.1 to 25.4, 26.1, 26.2, 27.1, and 28.1 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

NovaCopper Inc.	28-1	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

Original document signed and sealed by
   Michael F. O’Brien, M.Sc., Pr.Sci.Nat.,
          FGSSA, FAusIMM, FSAIMM                              
Michael F. O’Brien, M.Sc.,
Pr.Sci.Nat., FGSSA,
FAusIMM, FSAIMM
Chief Geologist
Tetra Tech WEI Inc.

NovaCopper Inc.	28-2	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.2 JIANHUI(JOHN) HUANG, PH.D., P.ENG.

I, Jianhui (John) Huang, Ph.D., P.Eng., of Burnaby, British Columbia, do hereby certify:

• I am a Senior Metallurgist with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of North-East University (B.Eng., 1982), Beijing General Research Institute for Non-ferrous Metals (M.Eng., 1988), and Birmingham University (Ph.D., 2000). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #30898). My relevant experience with respect to mineral engineering includes more than 30 years of involvement in mineral process for metal recovery from various base metal, gold, silver and rare metal ores. I have relevant experience in copper, lead, zinc, silver and gold recovery from various ores, including from multi-metal massive and semi-massive sulphide mineralization. Projects include the Fuping Gold-Silver-Copper-Lead-Zinc Massive/Semi-Massive Ore Project, the Tulsequah Chief Project and the Bellekeno Project. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 1.5, 1.7, 1.11, 13.0, 17.0, 18.3, 19.0, 21.1, 21.5.1, 21.5.3, 21.5.4, 25.5, 26.3, 26.4, 26.6.3, 27.6, and 28.2 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

Original document signed and sealed by
   Jianhui (John) Huang, Ph.D., P.Eng.       
Jianhui (John) Huang, Ph.D., P.Eng.
Senior Metallurgist
Tetra Tech WEI Inc.

NovaCopper Inc.	28-3	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.3 SABRYABDELHAFEZ, PH.D., P.ENG.

I, Sabry Abdel Hafez, Ph.D., P.Eng., of Vancouver, British Columbia, do hereby certify:

• I am a Senior Mining Engineer with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of Assiut University (B.Sc Mining Engineering, 1991; M.Sc. in Mining Engineering, 1996; Ph.D. in Mineral Economics, 2000). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #34975). My relevant experience is in mine evaluation. I have more than 19 years of experience in the evaluation of mining projects, advanced financial analysis, and mine planning and optimization. My capabilities range from the conventional mine planning and evaluation to the advanced simulation-based techniques that incorporate both market and geological uncertainties. I have been involved in the technical studies of several base metals, gold, coal, and aggregate mining projects in Canada and abroad. I have recently been involved in the technical reports for the Copper Fox’s Schaft Creek project feasibility study, Pretium Resources’ Brucejack project feasibility study, AQM’s Zafranal PEA, Castle Resources’ Granduc project PEA study and Seabridge’s KSM project prefeasibility study. I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

• My most recent personal inspection of the Property was June 19, 2013 for three days.

• I am responsible for Sections 1.6, 1.12, 15.0, 16.0, 18.7, 21.4, 21.5.2, 21.5.6, 22.0, 25.6, 25.10, 26.5, 26.8, 27.3, 27.5, and 28.3 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

NovaCopper Inc.	28-4	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Original document signed and sealed by
      Sabry Abdel Hafez, Ph.D., P.Eng.             
Sabry Abdel Hafez, Ph.D., P.Eng.
Senior Mining Engineer
Tetra Tech WEI Inc.

28.4 HASSANGHAFFARI, P.ENG.

I, Hassan Ghaffari, P.Eng., of Vancouver, British Columbia, do hereby certify:

• I am the Director of Metallurgy with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of the University of Tehran (M.A.Sc., Mining Engineering, 1990) and the University of British Columbia (M.A.Sc., Mineral Process Engineering, 2004). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #30408). My relevant experience includes 23 years of experience in mining and plant operation, project studies, management, and engineering. As the lead metallurgist for the Pebble Copper/Gold Moly Project in Alaska, I am coordinating all metallurgical test work and preparing and peer reviewing the technical report and the operating and capital costs of the plant and infrastructure for both the scoping and prefeasibility studies. For the Ajax Copper-Gold Project in BC, I was the Project Manager responsible for the process, infrastructure and overall management of the 60,000 t/d mill. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 1.8 (except 1.8.2), 18.1, 18.4 to 18.6, 18.8, 18.12 to 18.14, 25.7, 26.6 (except 26.6.2 to 26.6.5) and 28.4 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NovaCopper Inc.	28-5	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

Original document signed and sealed by
           Hassan Ghaffari, P.Eng.                    
Hassan Ghaffari, P.Eng.
Director of Metallurgy
Tetra Tech WEI Inc.

NovaCopper Inc.	28-6	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.5 HARVEYWAYNESTOYKO, P.ENG.

I, Harvey Wayne Stoyko, P.Eng., of Vancouver, British Columbia, do hereby certify:

• I am the Manager of Estimating with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of the University of Saskatchewan (B.Sc. Mechanical Engineering, 1985). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #17092). My relevant experience with respect to mine development and costing includes over 25 years of combined mining experience. This includes capital cost engineering/cost control for both greenfield and brownfield studies, and acquisitions/mergers or development of properties (construction) with Placer Dome. I have also been involved as an Owners representative with the planning, costing/cost control and execution of mine/concentrate handling facilities including plant, road, rail and port with the Port of Vancouver (Kinder Morgan) for Comino’s Red Dog Project. At Tetra Tech, I have directed the preparation of capital cost estimates for technical studies and I am responsible for project controls for ongoing studies and EPCM projects. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 1.10, 21.2 to 21.3, 26.6, 27.4 and 28.5 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

Original document signed and sealed by 
          Harvey Wayne Stoyko, P.Eng.           
Harvey Wayne Stoyko, P.Eng.
Manager of Estimating
Tetra Tech WEI Inc.

NovaCopper Inc.	28-7	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.6 MICHAELCHIN, P.ENG.

I, Michael Chin, P.Eng., of Vancouver, British Columbia, do hereby certify:

• I am a Civil Engineer with Tetra Tech WEI Inc. with a business address at 800- 555 West Hastings Street, Vancouver, British Columbia, V6B 1M1.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of the University of Alberta (Bachelor of Science in Civil Engineering, 1985). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #17172). My relevant experience is 26 years of civil engineering design and construction for mines, power plants, highways, and other heavy civil projects. Projects include the Zafranal greenfield copper gold mine in Peru and the Brucejack greenfield underground gold mine in Stewart, BC, Canada. I was also the Civil Project Engineer for the Eskay Creek Mine in Smithers, BC, Canada and the Willow Creek Mine in Chetwynd, BC, Canada. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 18.2.1, 18.2.3, and 28.6 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

Original document signed and sealed by 
                Michael Chin, P.Eng.                  
Michael Chin, P.Eng.
Civil Engineer
Tetra Tech WEI Inc.

NovaCopper Inc.	28-8	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.7 GRAHAMWILKINS, P.ENG.

I, Graham Wilkins, P.Eng., of Vancouver, British Columbia, do hereby certify:

• I am a Senior Transportation Engineer with EBA, A Tetra Tech Company with a business address at Oceanic Plaza, 9^thFloor, 1066 West Hastings Street, Vancouver, British Columbia, V6E 3X2.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of Carleton University (B.Eng., Civil Engineering, 1986). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia, License #16255. I have 27 years relevant transportation experience. I have worked on projects from the initial planning stages right through to detailed design and construction completion. My transportation roles are comprehensive including construction supervision, design, specs and project management. I have worked on resource and mining projects throughout BC, Alaska, Yukon, NWT and Nunavut. I was involved in the planning and design of the transportation infrastructure to the Kiggavik Mine near Baker Lake, Nunavut. Elements of the infrastructure included 100 km of an all-weather road and an option of a 100 km long winter road, 400 m bridge, port facility and an airstrip. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• My most recent personal inspection of the Property was in the fall of 2011 for one day.

• I am responsible for Sections 18.2.2, 18.2.4, 26.6.2, 26.6.4, and 28.7 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have prior involvement with the Property that is the subject of the Technical Report. I was a Qualified Person for NovaGold’s 2011 technical report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September, 2013 at Vancouver, British Columbia

NovaCopper Inc.	28-9	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

Original document signed and sealed by 
              Graham Wilkins, P.Eng.               
Graham Wilkins, P.Eng.
Senior Transportation Engineer
EBA, A Tetra Tech Company

NovaCopper Inc.	28-10	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.8 MARVINSILVA, PH.D., PE, P.ENG.

I, Marvin Silva, Ph.D., PE, P.Eng., of Tucson, Arizona, do hereby certify:

• I am a Senior Geotechnical Engineer with Tetra Tech Inc. with a business address at 3031 West Ina Road, Tucson, Arizona 85741.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of the National Autonomous University of Nicaragua (B.Sc Agricultural Engineering, 1981); the Institute of Odessa in Ukraine (M.Sc. Water Resources Engineering, 1985); and the University of Alberta (Ph.D. Geoenvironmental Engineering, 1999). I am a member in good standing of the Association of Professional Engineers and Geoscientists of Alberta (License #52477). My relevant experience with respect to geotechnical engineering includes 21 years of experience in mining and infrastructure projects in Canada and the United States of America, including 7.5 years dedicated to research of mine waste tailings and design of tailings storage facilities. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 1.8.2, 18.9 to 18.11, 21.5.5, 25.8, 26.5.5, and 28.8 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September 2013 at Tucson, Arizona

Original document signed and sealed by 
         Marvin Silva, Ph.D., PE, P.Eng.          
Marvin Silva, Ph.D., PE, P.Eng.
Senior Geotechnical Engineer
Tetra Tech Inc.

NovaCopper Inc.	28-11	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

28.9 JACKDIMARCHI, CPG

I, Jack DiMarchi, CPG, Fairbanks, Alaska, do hereby certify:

• I am a Project Manager/Senior Scientist with Tetra Tech Inc. with a business address at 615 Bidwell Avenue, Box 8, Fairbanks, Alaska, 99701.

• This certificate applies to the technical report entitled “Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska”, dated September 12, 2013 (the “Technical Report”).

• I am a graduate of Colorado State University (BS Geology 1978). I am a member in good standing of the American Institute of Professional Geologists, Certification #9217. My relevant experience is having served in several capacities in mine development, permitting and regulating mines including 10 years as Chief Geologist with Tech Resources on the Pogo Project, 3.5 years as Large Mine Coordinator with the State of Alaska Department of Natural Resources Office of Permitting and Project Management and 1.5 years with Tetra Tech responsible for authoring Environmental Chapters in other 43-101 reports including the Sun Project, Oracle Ridge Project and Rosemont Copper Project and other environmental permitting work on projects that include the Livengood Project. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

• I have not completed a personal inspection of the Property.

• I am responsible for Sections 1.9, 20.0, 25.9, 26.7, 27.2, and 28.9 of the Technical Report.

• I am independent of NovaCopper Inc. as defined by Section 1.5 of the Instrument.

• I have no prior involvement with the Property that is the subject of the Technical Report.

• I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument.

• As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and dated this 12^thday of September, 2013 at Fairbanks, Alaska

Original document signed and sealed by
                 Jack DiMarchi, CPG                  
Jack DiMarchi, CPG
Project Manager/Senior Scientist Tetra Tech Inc.

NovaCopper Inc.	28-12	1297650100-REP-R0002-03
Preliminary Economic Assessment Report on the Arctic
Project, Ambler Mining District, Northwest Alaska

A P P E N D I X  A

CLAIMS MAP AND CLAIMS LIST

Claim No.	Claim Name	Status	Grant Date	Expiry Date	Acres	Govt Office	T.	R.	Sec.	Qtr.
540543	Arctic 40A	Granted	August 29, 1990	November 30, 2013	2	Kotzebue	21N	11E	35	SW
540544	Arctic 496A	Granted	August 29, 1990	November 30, 2013	2	Kotzebue	21N	11E	34	SE
540545	Arctic 1001	Granted	July 16, 1984	November 30, 2013	8	Kotzebue	21N	11E	34	SE
540546	Arctic 1002	Granted	July 16, 1984	November 30, 2013	8	Kotzebue	21N	11E	34	SE & SW
540549	Arctic 1005	Granted	September 2, 1990	November 30, 2013	6	Kotzebue	21N	11E	35	SW
546144	SC 24	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	16	SW & SE
546145	SC 25	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	16	SW, SE, NW & NE
546146	SC 26	Granted	June 28, 1995	November 30, 2013	40	Kotzebue	21N	10E	16	NW & NE
546147	SC 34	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	16	SE
546148	SC 35	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	16	SE & NE
546149	SC 36	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	16	NE
546150	SC 44	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15; 16	SW; SE
546151	SC 45	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15; 16	SW & NW; SE & NE
546152	SC 46	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15; 16	NW; NE
546153	SC 54	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15	SW
546154	SC 55	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15	SW & NW
546155	SC 56	Granted	June 29, 1995	November 30, 2013	40	Kotzebue	21N	10E	15	NW
546156	SC 64	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15	SW & SE
546157	SC 65	Granted	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15	SW, SE, NW & NE
546158	SC 66	Granted	June 29, 1995	November 30, 2013	40	Kotzebue	21N	10E	15	NW & NE
590853	AM 63-165	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NW
590854	AM 63-166	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NW
590855	AM 63-167	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NE
590856	AM 63-168	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NE
590857	AM 63-169	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NW
590858	AM 63-170	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NW
590859	AM 63-171	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NE
590860	AM 63-172	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NE
590874	AM 64-165	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NW
590875	AM 64-166	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NW
590876	AM 64-167	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NE
590877	AM 64-168	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	14	NE
590878	AM 64-169	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NW
590879	AM 64-170	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NW
590880	AM 64-171	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NE
590881	AM 64-172	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	13	NE
590895	AM 65-165	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SW
590896	AM 65-166	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SW
590897	AM 65-167	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SE
590898	AM 65-168	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SE
590899	AM 65-169	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SW
590900	AM 65-170	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SW
590901	AM 65-171	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SE
590902	AM 65-172	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SE
590916	AM 66-165	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SW
590917	AM 66-166	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SW
590918	AM 66-167	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SE
590919	AM 66-168	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	SE
590920	AM 66-169	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SW
590921	AM 66-170	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SW
590922	AM 66-171	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SE
590923	AM 66-172	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	SE
590940	AM 67-165	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	NW
590941	AM 67-166	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	NW
590942	AM 67-167	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	NE
590943	AM 67-168	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	11	NE

590944	AM 67-169	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	NW
590945	AM 67-170	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	NW
590946	AM 67-171	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	NE
590947	AM 67-172	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	9E	12	NE
590998	AM 56-186	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	27	NW
590999	AM 56-187	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	27	NE
591000	AM 56-188	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	27	NE
591001	AM 56-189	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	26	NW
591002	AM 56-190	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	26	NW
591003	AM 56-191	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	26	NE
591004	AM 56-192	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	26	NE
591005	AM 56-193	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	25	NW
591006	AM 56-194	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	25	NW
591007	AM 56-195	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	25	NE
591008	AM 57-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	19	SE
591009	AM 57-177	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SW
591010	AM 57-178	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SW
591011	AM 57-179	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SE
591012	AM 57-180	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SE
591013	AM 57-181	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SW
591014	AM 57-182	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SW
591015	AM 57-183	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SE
591016	AM 57-184	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SE
591017	AM 57-185	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SW
591018	AM 57-186	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SW
591019	AM 57-187	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SE
591020	AM 57-188	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SE
591021	AM 57-189	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	23	SW
591022	AM 57-190	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	23	SW
591023	AM 57-191	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	23	SE
591024	AM 57-192	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	23	SE
591025	AM 57-193	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	24	SW
591026	AM 57-194	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	24	SW
591027	AM 57-195	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	24	SE
591028	AM 58-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	19	SE
591029	AM 58-177	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SW
591030	AM 58-178	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SW
591031	AM 58-179	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SE
591032	AM 58-180	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	SE
591033	AM 58-181	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SW
591034	AM 58-182	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SW
591035	AM 58-183	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SE
591036	AM 58-184	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	SE
591037	AM 58-185	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SW
591038	AM 58-186	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SW
591039	AM 58-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	SE
591040	AM 58-188	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	22	SE
591041	AM 58-189	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	SW
591042	AM 58-190	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	SW
591043	AM 58-191	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	SE
591044	AM 58-192	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	SE
591045	AM 58-193	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	SW
591046	AM 58-194	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	SW
591047	AM 59-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	19	NE
591048	AM 59-177	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NW
591049	AM 59-178	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NW
591050	AM 59-179	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NE

591051	AM 59-180	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NE
591052	AM 59-181	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NW
591053	AM 59-182	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NW
591054	AM 59-183	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NE
591055	AM 59-184	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NE
591056	AM 59-185	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NW
591057	AM 59-186	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NW
591058	AM 59-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NE
591059	AM 59-188	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	22	NE
591060	AM 59-189	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NW
591061	AM 59-190	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NW
591062	AM 59-191	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NE
591063	AM 59-192	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NE
591064	AM 59-193	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NW
591065	AM 60-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	19	NE
591066	AM 60-177	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NW
591067	AM 60-178	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NW
591068	AM 60-179	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NE
591069	AM 60-180	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	20	NE
591070	AM 60-181	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NW
591071	AM 60-182	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NW
591072	AM 60-183	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NE
591073	AM 60-184	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	21	NE
591074	AM 60-185	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NW
591075	AM 60-186	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NW
591076	AM 60-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	22	NE
591077	AM 60-188	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	22	NE
591078	AM 60-189	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NW
591079	AM 60-190	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NW
591080	AM 60-191	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NE
591081	AM 60-192	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	23	NE
591082	AM 60-193	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NW
591083	AM 61-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	SE
591084	AM 61-177	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SW
591085	AM 61-178	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SW
591086	AM 61-179	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SE
591087	AM 61-180	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SE
591088	AM 61-181	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SW
591089	AM 61-182	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SW
591090	AM 61-183	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SE
591091	AM 61-184	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SE
591092	AM 61-185	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	SW
591093	AM 61-186	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	SW
591094	AM 61-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	SE
591095	AM 61-188	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	15	SE
591096	AM 61-189	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	14	SW
591097	AM 61-190	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	14	SW
591098	AM 61-191	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	14	SE
591099	AM 61-192	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	14	SE
591100	AM 61-193	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	13	SW
591101	AM 62-176	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	SE
591102	AM 62-177	Granted	September 9, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SW
591103	AM 62-178	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SW
591104	AM 62-179	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SE
591105	AM 62-180	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	SE
591106	AM 62-181	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SW
591107	AM 62-182	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	SW

591108	AM 62-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	SE
591109	AM 62-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	SE
591110	AM 62-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	SW
591111	AM 62-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	SW
591112	AM 62-191	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	SE
591113	AM 62-192	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	SE
591114	AM 62-193	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	13	SW
591115	AM 63-173	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NW
591116	AM 63-174	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NW
591117	AM 63-175	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NE
591118	AM 63-176	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NE
591119	AM 63-177	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NW
591120	AM 63-178	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NW
591121	AM 63-179	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NE
591122	AM 63-180	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NE
591123	AM 63-181	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NW
591124	AM 63-182	Granted	September 5, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NW
591125	AM 63-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NE
591126	AM 63-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NE
591127	AM 63-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NW
591128	AM 63-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NW
591129	AM 63-191	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NE
591130	AM 63-192	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NE
591131	AM 63-193	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	13	NW
591132	AM 64-173	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NW
591133	AM 64-174	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NW
591134	AM 64-175	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NE
591135	AM 64-176	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	18	NE
591136	AM 64-177	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NW
591137	AM 64-178	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NW
591138	AM 64-179	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NE
591139	AM 64-180	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	17	NE
591140	AM 64-181	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NW
591141	AM 64-182	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NW
591142	AM 64-183	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NE
591143	AM 64-184	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	16	NE
591144	AM 64-185	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NW
591145	AM 64-186	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NW
591146	AM 64-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NE
591147	AM 64-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	15	NE
591148	AM 64-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NW
591149	AM 64-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NW
591150	AM 64-191	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NE
591151	AM 64-192	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	14	NE
591152	AM 64-193	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	13	NW
591153	AM 65-173	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SW
591154	AM 65-174	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SW
591155	AM 65-175	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SE
591156	AM 65-176	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SE
591157	AM 65-177	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SW
591158	AM 65-178	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SW
591159	AM 65-179	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SE
591160	AM 65-180	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SE
591161	AM 65-181	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SW
591162	AM 65-182	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SW
591163	AM 65-183	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SE
591164	AM 65-184	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SE

591165	AM 65-185	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SW
591166	AM 65-186	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SW
591167	AM 65-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SE
591168	AM 65-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SE
591169	AM 65-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SW
591170	AM 65-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SW
591171	AM 65-191	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SE
591172	AM 65-192	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SE
591173	AM 65-193	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	SW
591174	AM 66-173	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SW
591175	AM 66-174	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SW
591176	AM 66-175	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SE
591177	AM 66-176	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	SE
591178	AM 66-177	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SW
591179	AM 66-178	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SW
591180	AM 66-179	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SE
591181	AM 66-180	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	SE
591182	AM 66-181	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SW
591183	AM 66-182	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SW
591184	AM 66-183	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SE
591185	AM 66-184	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	SE
591186	AM 66-185	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SW
591187	AM 66-186	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SW
591188	AM 66-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SE
591189	AM 66-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	SE
591190	AM 66-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SW
591191	AM 66-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SW
591192	AM 66-191	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SE
591193	AM 66-192	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	SE
591194	AM 66-193	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	SW
591195	AM 67-173	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	NW
591196	AM 67-174	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	NW
591197	AM 67-175	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	NE
591198	AM 67-176	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	7	NE
591199	AM 67-177	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	NW
591200	AM 67-178	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	NW
591201	AM 67-179	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	NE
591202	AM 67-180	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	8	NE
591203	AM 67-181	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	NW
591204	AM 67-182	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	NW
591205	AM 67-183	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	NE
591206	AM 67-184	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	9	NE
591207	AM 67-185	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	NW
591208	AM 67-186	Granted	September 6, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	NW
591209	AM 67-187	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	NE
591210	AM 67-188	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	10	NE
591211	AM 67-189	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	NW
591212	AM 67-190	Granted	September 10, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	NW
591213	AM 67-191	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	NE
591214	AM 67-192	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	11	NE
591215	AM 67-193	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	NW
591216	AM 67-194	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	NW
591217	AM 67-195	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	NE
591218	AM 67-196	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	10E	12	NE
591219	AM 49-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SW
591220	AM 49-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SE
591221	AM 49-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SE

591222	AM 49-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	SW
591223	AM 49-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	SW
591224	AM 49-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	SW
591225	AM 49-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	SE
591226	AM 49-216	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	35	SE
591227	AM 49-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SW
591228	AM 49-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SW
591229	AM 49-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SE
591230	AM 49-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SE
591231	AM 50-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SW
591232	AM 50-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SE
591233	AM 50-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	SE
591234	AM 50-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	SW
591235	AM 50-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	SW
591236	AM 50-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	SE
591237	AM 50-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	SW
591238	AM 50-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	SW
591239	AM 50-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	SE
591240	AM 50-216	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	35	SE
591241	AM 50-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SW
591242	AM 50-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SW
591243	AM 50-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SE
591244	AM 50-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	SE
591245	AM 51-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NW
591246	AM 51-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NE
591247	AM 51-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NE
591248	AM 51-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	NW
591249	AM 51-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	NW
591250	AM 51-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	NE
591251	AM 51-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	NE
591252	AM 51-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	NW
591253	AM 51-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	NW
591254	AM 51-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	NE
591255	AM 51-216	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	NE
591256	AM 51-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NW
591257	AM 51-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NW
591258	AM 51-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NE
591259	AM 51-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NE
591260	AM 52-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NW
591261	AM 52-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NE
591262	AM 52-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	33	NE
591263	AM 52-209	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	34	NW
591264	AM 52-210	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	34	NW
591265	AM 52-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	34	NE
591266	AM 52-212	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	34	NE
591267	AM 52-213	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	35	NW
591268	AM 52-214	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	35	NW
591269	AM 52-215	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	35	NE
591270	AM 52-216	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	35	NE
591271	AM 52-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NW
591272	AM 52-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NW
591273	AM 52-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NE
591274	AM 52-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	36	NE
591275	AM 53-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SW
591276	AM 53-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SE
591277	AM 53-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SE
591278	AM 53-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SW

591279	AM 53-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SW
591280	AM 53-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SE
591281	AM 53-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SE
591282	AM 53-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SW
591283	AM 53-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SW
591284	AM 53-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SE
591285	AM 53-216	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	26	SE
591286	AM 53-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SW
591287	AM 53-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SW
591288	AM 53-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SE
591289	AM 53-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SE
591290	AM 54-206	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SW
591291	AM 54-207	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SE
591292	AM 54-208	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	SE
591293	AM 54-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SW
591294	AM 54-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SW
591295	AM 54-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SE
591296	AM 54-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	SE
591297	AM 54-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SW
591298	AM 54-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SW
591299	AM 54-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	SE
591300	AM 54-216	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	26	SE
591301	AM 54-217	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SW
591302	AM 54-218	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SW
591303	AM 54-219	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SE
591304	AM 54-220	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	SE
591305	AM 55-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	NW
591306	AM 55-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	NE
591307	AM 55-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	28	NE
591308	AM 55-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NW
591309	AM 55-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NW
591310	AM 55-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NE
591311	AM 55-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NE
591312	AM 55-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NW
591313	AM 55-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NW
591314	AM 55-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NE
591315	AM 55-216	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	26	NE
591316	AM 55-217	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NW
591317	AM 55-218	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NW
591318	AM 55-219	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NE
591319	AM 55-220	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NE
591320	AM 56-206	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	28	NW
591321	AM 56-207	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	28	NE
591322	AM 56-208	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	28	NE
591323	AM 56-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NW
591324	AM 56-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NW
591325	AM 56-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NE
591326	AM 56-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	27	NE
591327	AM 56-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NW
591328	AM 56-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NW
591329	AM 56-215	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	26	NE
591330	AM 56-216	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	26	NE
591331	AM 56-217	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NW
591332	AM 56-218	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NW
591333	AM 56-219	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NE
591334	AM 56-220	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	25	NE
591335	AM 57-206	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SW

591336	AM 57-207	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SE
591337	AM 57-208	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SE
591338	AM 57-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SW
591339	AM 57-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SW
591340	AM 57-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SE
591341	AM 57-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SE
591342	AM 57-213	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	23	SW
591343	AM 57-214	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	23	SW
591344	AM 57-215	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SE
591345	AM 57-216	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SE
591346	AM 57-217	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SW
591347	AM 57-218	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SW
591348	AM 57-219	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SE
591349	AM 57-220	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SE
591350	AM 58-206	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SW
591351	AM 58-207	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SE
591352	AM 58-208	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	21	SE
591353	AM 58-209	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SW
591354	AM 58-210	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SW
591355	AM 58-211	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SE
591356	AM 58-212	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	22	SE
591357	AM 58-213	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SW
591358	AM 58-214	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SW
591359	AM 58-215	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SE
591360	AM 58-216	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	SE
591361	AM 58-217	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SW
591362	AM 58-218	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SW
591363	AM 58-219	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SE
591364	AM 58-220	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	SE
591365	AM 59-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NW
591366	AM 59-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NE
591367	AM 59-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NE
591368	AM 59-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NW
591369	AM 59-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NW
591370	AM 59-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NE
591371	AM 59-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NE
591372	AM 59-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NW
591373	AM 59-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NW
591374	AM 59-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NE
591375	AM 59-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NE
591376	AM 59-213	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NW
591377	AM 59-214	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NW
591378	AM 59-215	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NE
591379	AM 59-216	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NE
591380	AM 59-217	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	NW
591381	AM 59-218	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	NW
591382	AM 60-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NW
591383	AM 60-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NE
591384	AM 60-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	20	NE
591385	AM 60-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NW
591386	AM 60-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NW
591387	AM 60-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NE
591388	AM 60-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	21	NE
591389	AM 60-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NW
591390	AM 60-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NW
591391	AM 60-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NE
591392	AM 60-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	22	NE

591393	AM 60-213	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NW
591394	AM 60-214	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NW
591395	AM 60-215	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NE
591396	AM 60-216	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	23	NE
591397	AM 60-217	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	NW
591398	AM 60-218	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	24	NW
591399	AM 61-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SW
591400	AM 61-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SE
591401	AM 61-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SE
591402	AM 61-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SW
591403	AM 61-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SW
591404	AM 61-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SE
591405	AM 61-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SE
591406	AM 61-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SW
591407	AM 61-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SW
591408	AM 61-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SE
591409	AM 61-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SE
591410	AM 61-213	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SW
591411	AM 61-214	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SW
591412	AM 61-215	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SE
591413	AM 61-216	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SE
591414	AM 61-217	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	13	SW
591415	AM 61-218	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	13	SW
591416	AM 62-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SW
591417	AM 62-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SE
591418	AM 62-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	SE
591419	AM 62-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SW
591420	AM 62-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SW
591421	AM 62-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SE
591422	AM 62-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	SE
591423	AM 62-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SW
591424	AM 62-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SW
591425	AM 62-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SE
591426	AM 62-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	SE
591427	AM 62-213	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SW
591428	AM 62-214	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SW
591429	AM 62-215	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SE
591430	AM 62-216	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	14	SE
591431	AM 62-217	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	13	SW
591432	AM 62-218	Granted	September 19, 1997	November 30, 2013	40	Kotzebue	21N	11E	13	SW
591433	AM 63-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NW
591434	AM 63-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NE
591435	AM 63-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NE
591436	AM 63-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NW
591437	AM 63-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NW
591438	AM 63-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NE
591439	AM 63-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NE
591440	AM 63-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NW
591441	AM 63-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NW
591442	AM 63-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NE
591443	AM 63-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NE
591444	AM 64-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NW
591445	AM 64-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NE
591446	AM 64-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	17	NE
591447	AM 64-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NW
591448	AM 64-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NW
591449	AM 64-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NE

591450	AM 64-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	16	NE
591451	AM 64-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NW
591452	AM 64-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NW
591453	AM 64-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NE
591454	AM 64-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	15	NE
591455	AM 65-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SW
591456	AM 65-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SE
591457	AM 65-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SE
591458	AM 65-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SW
591459	AM 65-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SW
591460	AM 65-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SE
591461	AM 65-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SE
591462	AM 65-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SW
591463	AM 65-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SW
591464	AM 65-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SE
591465	AM 65-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SE
591466	AM 66-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SW
591467	AM 66-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SE
591468	AM 66-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	SE
591469	AM 66-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SW
591470	AM 66-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SW
591471	AM 66-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SE
591472	AM 66-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	SE
591473	AM 66-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SW
591474	AM 66-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SW
591475	AM 66-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SE
591476	AM 66-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	SE
591477	AM 67-197	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	7	NW
591478	AM 67-198	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	7	NW
591479	AM 67-199	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	7	NE
591480	AM 67-200	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	7	NE
591481	AM 67-201	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	NW
591482	AM 67-202	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	NW
591483	AM 67-203	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	NE
591484	AM 67-204	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	8	NE
591485	AM 67-205	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	NW
591486	AM 67-206	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	NW
591487	AM 67-207	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	NE
591488	AM 67-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	NE
591489	AM 67-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NW
591490	AM 67-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NW
591491	AM 67-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NE
591492	AM 67-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NE
591493	AM 68-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	9	NE
591494	AM 68-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NW
591495	AM 68-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NW
591496	AM 68-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NE
591497	AM 68-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	10	NE
591498	AM 69-208	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	4	SE
591499	AM 69-209	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	3	SW
591500	AM 69-210	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	3	SW
591501	AM 69-211	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	3	SE
591502	AM 69-212	Granted	September 8, 1997	November 30, 2013	40	Kotzebue	21N	11E	3	SE
591503	AM 49-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SW
591504	AM 49-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SW
591505	AM 49-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SE
591506	AM 49-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SE

591507	AM 49-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SW
591508	AM 49-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SW
591509	AM 49-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SE
591510	AM 49-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SE
591511	AM 49-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	SW
591512	AM 49-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	SW
591513	AM 50-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SW
591514	AM 50-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SW
591515	AM 50-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SE
591516	AM 50-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	SE
591517	AM 50-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SW
591518	AM 50-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SW
591519	AM 50-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SE
591520	AM 50-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	SE
591521	AM 50-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	SW
591522	AM 50-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	SW
591523	AM 51-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NW
591524	AM 51-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NW
591525	AM 51-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NE
591526	AM 51-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NE
591527	AM 51-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NW
591528	AM 51-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NW
591529	AM 51-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NE
591530	AM 51-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NE
591531	AM 51-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	NW
591532	AM 51-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	NW
591533	AM 52-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NW
591534	AM 52-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NW
591535	AM 52-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NE
591536	AM 52-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	31	NE
591537	AM 52-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NW
591538	AM 52-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NW
591539	AM 52-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NE
591540	AM 52-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	32	NE
591541	AM 52-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	NW
591542	AM 52-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	33	NW
591543	AM 53-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SW
591544	AM 53-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SW
591545	AM 53-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SE
591546	AM 53-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SE
591547	AM 54-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SW
591548	AM 54-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SW
591549	AM 54-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SE
591550	AM 54-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	SE
591551	AM 55-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NW
591552	AM 55-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NW
591553	AM 55-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NE
591554	AM 55-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NE
591555	AM 56-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NW
591556	AM 56-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NW
591557	AM 56-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NE
591558	AM 56-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	21N	12E	30	NE
591575	AM 37-226	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SW
591576	AM 37-227	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SE
591577	AM 37-228	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SE
591578	AM 37-229	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	SW
591579	AM 37-230	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	SW

591590	AM 38-226	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SW
591591	AM 38-227	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SE
591592	AM 38-228	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	SE
591593	AM 38-229	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	SW
591594	AM 38-230	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	SW
591605	AM 39-226	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NW
591606	AM 39-227	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NE
591607	AM 39-228	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NE
591608	AM 39-229	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	NW
591609	AM 39-230	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	NW
591620	AM 40-226	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NW
591621	AM 40-227	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NE
591622	AM 40-228	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	16	NE
591623	AM 40-229	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	NW
591624	AM 40-230	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	15	NW
591635	AM 41-225	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SW
591636	AM 41-226	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SW
591637	AM 41-227	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SE
591638	AM 41-228	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SE
591639	AM 41-229	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	SW
591640	AM 41-230	Granted	September 7, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	SW
591648	AM 42-223	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	SE
591649	AM 42-224	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	SE
591650	AM 42-225	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SW
591651	AM 42-226	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SW
591652	AM 42-227	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SE
591653	AM 42-228	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	SE
591654	AM 42-229	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	SW
591655	AM 42-230	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	SW
591661	AM 43-221	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NW
591662	AM 43-222	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NW
591663	AM 43-223	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NE
591664	AM 43-224	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NE
591665	AM 43-225	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NW
591666	AM 43-226	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NW
591667	AM 43-227	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NE
591668	AM 43-228	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NE
591669	AM 43-229	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	NW
591670	AM 43-230	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	NW
591676	AM 44-219	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	7	NE
591677	AM 44-220	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	7	NE
591678	AM 44-221	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NW
591679	AM 44-222	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NW
591680	AM 44-223	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NE
591681	AM 44-224	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	8	NE
591682	AM 44-225	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NW
591683	AM 44-226	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NW
591684	AM 44-227	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NE
591685	AM 44-228	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	9	NE
591686	AM 44-229	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	NW
591687	AM 44-230	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	10	NW
591693	AM 45-217	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SW
591694	AM 45-218	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SW
591695	AM 45-219	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SE
591696	AM 45-220	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SE
591697	AM 45-221	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SW
591698	AM 45-222	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SW

591699	AM 45-223	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SE
591700	AM 45-224	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SE
591701	AM 45-225	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SW
591702	AM 45-226	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SW
591703	AM 45-227	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SE
591704	AM 45-228	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SE
591705	AM 45-229	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	SW
591706	AM 45-230	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	SW
591712	AM 46-217	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SW
591713	AM 46-218	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SW
591714	AM 46-219	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SE
591715	AM 46-220	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	SE
591716	AM 46-221	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SW
591717	AM 46-222	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SW
591718	AM 46-223	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SE
591719	AM 46-224	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	SE
591720	AM 46-225	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SW
591721	AM 46-226	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SW
591722	AM 46-227	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SE
591723	AM 46-228	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	SE
591724	AM 46-229	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	SW
591725	AM 46-230	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	SW
591731	AM 47-217	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NW
591732	AM 47-218	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NW
591733	AM 47-219	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NE
591734	AM 47-220	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NE
591735	AM 47-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NW
591736	AM 47-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NW
591737	AM 47-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NE
591738	AM 47-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NE
591739	AM 47-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NW
591740	AM 47-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NW
591741	AM 47-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NE
591742	AM 47-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NE
591743	AM 47-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	NW
591744	AM 47-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	NW
591745	AM 48-217	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NW
591746	AM 48-218	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NW
591747	AM 48-219	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NE
591748	AM 48-220	Granted	September 11, 1997	November 30, 2013	40	Kotzebue	20N	12E	6	NE
591749	AM 48-221	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NW
591750	AM 48-222	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NW
591751	AM 48-223	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NE
591752	AM 48-224	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	5	NE
591753	AM 48-225	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NW
591754	AM 48-226	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NW
591755	AM 48-227	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NE
591756	AM 48-228	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	4	NE
591757	AM 48-229	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	NW
591758	AM 48-230	Granted	September 18, 1997	November 30, 2013	40	Kotzebue	20N	12E	3	NW
634110	EDC 1	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SW
634111	EDC 2	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SE
634112	EDC 3	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SE
634113	EDC 4	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SW
634114	EDC 5	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SW
634115	EDC 6	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SE
634116	EDC 7	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SE

634117	EDC 8	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	8	SW
634118	EDC 9	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SW
634119	EDC 10	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SE
634120	EDC 11	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	12	SE
634121	EDC 12	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SW
634122	EDC 13	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SW
634123	EDC 14	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SE
634124	EDC 15	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	7	SE
634125	EDC 16	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	8	SW
634126	EDC 17	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NW
634127	EDC 18	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NE
634128	EDC 19	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NE
634129	EDC 20	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NW
634130	EDC 21	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NW
634131	EDC 22	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NE
634132	EDC 23	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NE
634133	EDC 24	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	17	NW
634134	EDC 25	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NW
634135	EDC 26	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NE
634136	EDC 27	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	NE
634137	EDC 28	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NW
634138	EDC 29	Granted	April 26, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	NW
634139	EDC 30	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	NE
634140	EDC 31	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	NE
634141	EDC 32	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	17	NW
634142	EDC 33	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	SW
634143	EDC 34	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	SE
634144	EDC 35	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	10E	13	SE
634145	EDC 36	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SW
634146	EDC 37	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SW
634147	EDC 38	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SE
634148	EDC 39	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SE
634149	EDC 40	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	17	SW
634150	EDC 41	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	13	SW
634151	EDC 42	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	13	SE
634152	EDC 43	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	13	SE
634153	EDC 44	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SW
634154	EDC 45	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SW
634155	EDC 46	Granted TA	August 19, 2011	November 30, 2013	40	Kotzebue	21N	11E	18	SE
634156	EDC 47	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	18	SE
634157	EDC 48	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	17	SW
634158	EDC 49	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NW
634159	EDC 50	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NE
634160	EDC 51	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NE
634161	EDC 52	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NW
634162	EDC 53	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NW
634163	EDC 54	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NE
634164	EDC 55	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NE
634165	EDC 56	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	NW
634166	EDC 57	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NW
634167	EDC 58	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NE
634168	EDC 59	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	NE
634169	EDC 60	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NW
634170	EDC 61	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NW
634171	EDC 62	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NE
634172	EDC 63	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	NE
634173	EDC 64	Granted	April 27, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	NW

634174	EDC 65	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	SE
634175	EDC 66	Granted TA	July 18, 2006	November 30, 2013	40	Kotzebue	21N	10E	24	SE
634176	EDC 67	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SW
634177	EDC 68	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SW
634178	EDC 69	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SE
634179	EDC 70	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SE
634180	EDC 71	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SW
634181	EDC 72	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SW
634182	EDC 73	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SE
634183	EDC 74	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SE
634184	EDC 75	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	21	SW
634185	EDC 76	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	10E	24	SE
634186	EDC 77	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SW
634187	EDC 78	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SW
634188	EDC 79	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SE
634189	EDC 80	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	19	SE
634190	EDC 81	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SW
634191	EDC 82	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SW
634192	EDC 83	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SE
634193	EDC 84	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	20	SE
634194	EDC 85	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	21	SW
634195	EDC 86	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	10E	25	NE
634196	EDC 87	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	30	NW
634197	EDC 88	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	30	NW
634198	EDC 89	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	30	NE
634199	EDC 90	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	30	NE
634200	EDC 91	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	29	NW
634201	EDC 92	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	29	NW
634202	EDC 93	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	29	NE
634203	EDC 94	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	29	NE
634204	EDC 95	Granted	April 28, 2000	November 30, 2013	40	Kotzebue	21N	11E	28	NW
650291	HOSS 01	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	18	NW
650292	HOSS 02	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	18	NE
650293	HOSS 03	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	17	NW
650294	HOSS 04	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	17	NE
650295	HOSS 05	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	16	NW
650296	HOSS 06	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	16	NE
650297	HOSS 07	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	15	NW
650298	HOSS 08	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	18	SW
650299	HOSS 09	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	18	SE
650300	HOSS 10	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	17	SW
650301	HOSS 11	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	17	SE
650302	HOSS 12	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	16	SW
650303	HOSS 13	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	16	SE
650304	HOSS 14	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	15	SW
650305	HOSS 15	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	19	NW
650306	HOSS 16	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	19	NE
650307	HOSS 17	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	20	NW
650308	HOSS 18	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	20	NE
650309	HOSS 19	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	21	NW
650310	HOSS 20	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	21	NE
650311	HOSS 21	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	22	NW
650312	HOSS 22	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	19	SW
650313	HOSS 23	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	19	SE
650314	HOSS 24	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	20	SW
650315	HOSS 25	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	20	SE
650316	HOSS 26	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	21	SW

650317	HOSS 27	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	21	SE
650318	HOSS 28	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	22	SW
650319	HOSS 29	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	30	NW
650320	HOSS 30	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	30	NE
650321	HOSS 31	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	29	NW
650322	HOSS 32	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	29	NE
650323	HOSS 33	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	28	NW
650324	HOSS 34	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	28	NE
650325	HOSS 35	Granted	July 13, 2005	November 30, 2013	160	Kotzebue	22N	10E	27	NW
651299	GAP 1	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	28	NW
651300	GAP 2	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	28	NE
651301	GAP 3	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	27	NW
651302	GAP 4	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	27	NE
651303	GAP 5	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	26	NW
651304	GAP 6	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	26	NE
651305	GAP 7	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	25	NW
651306	GAP 8	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	25	NE
651307	GAP 9	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	30	NW
651308	GAP 10	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	30	NE
651309	GAP 11	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	29	NW
651310	GAP 12	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	28	SW
651311	GAP 13	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	28	SE
651312	GAP 14	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	27	SW
651313	GAP 15	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	27	SE
651314	GAP 16	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	26	SW
651315	GAP 17	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	26	SE
651316	GAP 18	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	25	SW
651317	GAP 19	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	25	SE
651318	GAP 20	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	30	SW
651319	GAP 21	Granted	September 27, 2005	November 30,��2013	160	Kotzebue	22N	9E	30	SE
651320	GAP 22	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	29	SW
651321	GAP 23	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	29	SE
651322	GAP 24	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	28	SW
651323	GAP 25	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	28	SE
651324	GAP 26	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	27	SW
651325	GAP 27	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	34	NE
651326	GAP 28	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	35	NW
651327	GAP 29	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	35	NE
651328	GAP 30	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	36	NW
651329	GAP 31	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	36	NE
651330	GAP 32	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	31	NW
651331	GAP 33	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	31	NE
651332	GAP 34	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	32	NW
651333	GAP 35	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	32	NE
651334	GAP 36	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	33	NW
651335	GAP 37	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	33	NE
651336	GAP 38	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	34	NW
651337	GAP 39	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	34	NE
651338	GAP 40	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	35	NW
651339	GAP 41	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	35	NE
651340	GAP 42	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	36	NW
651341	GAP 43	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	36	NE
651342	GAP 44	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	31	NW
651343	GAP 45	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	31	NE
651344	GAP 46	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	32	NW
651345	GAP 47	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	32	NE
651346	GAP 48	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	33	NW

651347	GAP 49	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	33	NE
651348	GAP 50	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	34	NW
651349	GAP 51	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	30	SW
651350	GAP 52	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	30	SE
651351	GAP 53	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	29	SW
651352	GAP 54	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	29	SE
651353	GAP 55	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	28	SW
651354	GAP 56	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	28	SE
651355	GAP 57	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	27	SW
651356	GAP 58	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	34	SE
651357	GAP 59	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	35	SW
651358	GAP 60	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	35	SE
651359	GAP 61	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	36	SW
651360	GAP 62	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	8E	36	SE
651361	GAP 63	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	31	SW
651362	GAP 64	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	31	SE
651363	GAP 65	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	32	SW
651364	GAP 66	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	32	SE
651365	GAP 67	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	33	SW
651366	GAP 68	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	33	SE
651367	GAP 69	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	34	SW
651368	GAP 70	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	34	SE
651369	GAP 71	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	35	SW
651370	GAP 72	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	35	SE
651371	GAP 73	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	36	SW
651372	GAP 74	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	9E	36	SE
651373	GAP 75	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	31	SW
651374	GAP 76	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	31	SE
651375	GAP 77	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	32	SW
651376	GAP 78	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	32	SE
651377	GAP 79	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	33	SW
651378	GAP 80	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	33	SE
651379	GAP 81	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	22N	10E	34	SW
651380	GAP 82	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	3	NE
651381	GAP 83	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	2	NW
651382	GAP 84	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	2	NE
651383	GAP 85	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	1	NW
651384	GAP 86	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	1	NE
651385	GAP 87	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	6	NW
651386	GAP 88	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	6	NE
651387	GAP 89	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	2	NW
651388	GAP 90	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	2	NE
651389	GAP 91	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	1	NW
651390	GAP 92	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	1	NE
651391	GAP 93	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	6	NW
651392	GAP 94	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	6	NE
651393	GAP 95	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	5	NW
651394	GAP 96	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	5	NE
651395	GAP 97	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	4	NW
651396	GAP 98	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	4	NE
651397	GAP 99	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	3	NW
651398	GAP 100	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	2	SW
651399	GAP 101	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	2	SE
651400	GAP 102	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	1	SW
651401	GAP 103	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	1	SE
651402	GAP 104	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	6	SW
651403	GAP 105	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	6	SE

651404	GAP 106	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	5	SW
651405	GAP 107	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	5	SE
651406	GAP 108	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	4	SW
651407	GAP 109	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	4	SE
651408	GAP 110	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	3	SW
651409	GAP 111	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	3	SE
651410	GAP 112	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	2	SW
651411	GAP 113	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	10E	2	SE
651412	GAP 114	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	11	NW
651413	GAP 115	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	11	NW
651414	GAP 116	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	11	NE
651415	GAP 117	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	11	NE
651416	GAP 118	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	12	NW
651417	GAP 119	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	12	NW
651418	GAP 120	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	12	NE
651419	GAP 121	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	9E	12	NE
651420	GAP 122	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	7	NW
651421	GAP 123	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	7	NW
651422	GAP 124	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	7	NE
651423	GAP 125	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	7	NE
651424	GAP 126	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	8	NW
651425	GAP 127	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	8	NW
651426	GAP 128	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	8	NE
651427	GAP 129	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	8	NE
651428	GAP 130	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	9	NW
651429	GAP 131	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	9	NW
651430	GAP 132	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	9	NE
651431	GAP 133	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	9	NE
651432	GAP 134	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	10	NW
651433	GAP 135	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	10	NW
651434	GAP 136	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	10	NE
651435	GAP 137	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	10	NE
651436	GAP 138	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	11	NW
651437	GAP 139	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	11	NW
651438	GAP 140	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	11	NE
651439	GAP 141	Granted	September 27, 2005	November 30, 2013	40	Kotzebue	21N	10E	11	NE
651440	GAP 142	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	5	NW
651441	GAP 143	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	5	NE
651442	GAP 144	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	4	NW
651443	GAP 145	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	4	NE
651444	GAP 146	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	3	NW
651445	GAP 147	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	3	NE
651446	GAP 148	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	3	SE
651447	GAP 149	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	2	SW
651448	GAP 150	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	2	SE
651449	GAP 151	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	1	SW
651450	GAP 152	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	1	SE
651451	GAP 153	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	6	SW
651452	GAP 154	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	6	SE
651453	GAP 155	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	5	SW
651454	GAP 156	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	5	SE
651455	GAP 157	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	4	SW
651456	GAP 158	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	4	SE
651457	GAP 159	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	3	SW
651458	GAP 160	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	3	SE
651459	GAP 161	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	10	NE
651460	GAP 162	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	11	NW

651461	GAP 163	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	11	NE
651462	GAP 164	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	12	NW
651463	GAP 165	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	12	NE
651464	GAP 166	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	7	NW
651465	GAP 167	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	7	NE
651466	GAP 168	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	8	NW
651467	GAP 169	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	8	NE
651468	GAP 170	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	9	NW
651469	GAP 171	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	9	NE
651470	GAP 172	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	10	NW
651471	GAP 173	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	10	NE
651472	GAP 174	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	10	SE
651473	GAP 175	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	11	SW
651474	GAP 176	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	11	SE
651475	GAP 177	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	12	SW
651476	GAP 178	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	12	SE
651477	GAP 179	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	7	SW
651478	GAP 180	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	7	SE
651479	GAP 181	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	8	SW
651480	GAP 182	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	8	SE
651481	GAP 183	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	9	SW
651482	GAP 184	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	9	SE
651483	GAP 185	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	10	SW
651484	GAP 186	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	10	SE
651485	GAP 187	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	15	NE
651486	GAP 188	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	14	NW
651487	GAP 189	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	14	NE
651488	GAP 190	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	13	NW
651489	GAP 191	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	8E	13	NE
651490	GAP 192	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	18	NW
651491	GAP 193	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	18	NE
651492	GAP 194	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	17	NW
651493	GAP 195	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	17	NE
651494	GAP 196	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	16	NW
651495	GAP 197	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	16	NE
651496	GAP 198	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	15	NW
651497	GAP 199	Granted	September 27, 2005	November 30, 2013	160	Kotzebue	21N	9E	15	NE
651270	PAL 1	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	10E	25	NE
651271	PAL 2	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	10E	25	NE
651272	PAL 3	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	30	NW
651273	PAL 4	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	30	NW
651274	PAL 5	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	30	NE
651275	PAL 6	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	30	NE
651276	PAL 7	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	29	NW
651277	PAL 8	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	29	NW
651278	PAL 9	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	29	NE
651279	PAL 10	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	29	NE
651280	PAL 11	Granted	September 26, 2005	November 30, 2013	40	Kotzebue	21N	11E	28	NW
651289	PAL 20	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	10E	25	SE
651290	PAL 21	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	30	SW
651291	PAL 22	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	30	SE
651292	PAL 23	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	29	SW
651293	PAL 24	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	29	SE
651294	PAL 25	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	28	SW
651296	PAL 27	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	32	NE
651297	PAL 28	Granted	September 26, 2005	November 30, 2013	160	Kotzebue	21N	11E	33	NW
651152	ZED 1	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	10	NE

651153	ZED 2	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	11	NW
651154	ZED 3	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	11	NE
651155	ZED 4	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	10	SE
651156	ZED 5	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	11	SW
651157	ZED 6	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	11	SE
651158	ZED 7	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	12	SW
651159	ZED 8	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	12	SE
651160	ZED 9	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	15	NE
651161	ZED 10	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	14	NW
651162	ZED 11	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	14	NE
651163	ZED 12	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	13	NW
651164	ZED 13	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	13	NE
651165	ZED 14	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	18	NW
651166	ZED 15	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	18	NE
651167	ZED 16	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	17	NW
651168	ZED 17	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	17	NE
651169	ZED 18	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	16	NW
651170	ZED 19	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	16	NE
651171	ZED 20	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	15	NW
651172	ZED 21	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	15	NE
651173	ZED 22	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	15	SE
651174	ZED 23	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	14	SW
651175	ZED 24	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	14	SE
651176	ZED 25	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	13	SW
651177	ZED 26	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	13	SE
651178	ZED 27	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	18	SW
651179	ZED 28	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	18	SE
651180	ZED 29	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	17	SW
651181	ZED 30	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	17	SE
651182	ZED 31	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	16	SW
651183	ZED 32	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	16	SE
651184	ZED 33	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	15	SW
651185	ZED 34	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	15	SE
651186	ZED 35	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	12E	24	NE
651187	ZED 36	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	19	NW
651188	ZED 37	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	19	NE
651189	ZED 38	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	20	NW
651190	ZED 39	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	20	NE
651191	ZED 40	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	21	NW
651192	ZED 41	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	21	NE
651193	ZED 42	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	22	NW
651194	ZED 43	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	22	NE
651195	ZED 44	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	19	SE
651196	ZED 45	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	20	SW
651197	ZED 46	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	20	SE
651198	ZED 47	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	21	SW
651199	ZED 48	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	21	SE
651200	ZED 49	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	22	SW
651201	ZED 50	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	22	SE
651202	ZED 51	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	23	NW
651203	ZED 52	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	23	NE
651204	ZED 53	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	24	NW
651205	ZED 54	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	24	NE
651206	ZED 55	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	19	NW
651207	ZED 56	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	19	NE
651208	ZED 57	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	20	NW
651209	ZED 58	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	20	NE

651210	ZED 59	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	21	NW
651211	ZED 60	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	21	NE
651212	ZED 61	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	22	NW
651213	ZED 62	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	22	NE
651214	ZED 63	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	23	NW
651215	ZED 64	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	23	NE
651216	ZED 65	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	24	NW
651217	ZED 66	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	24	NE
651218	ZED 67	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	15E	19	NW
651219	ZED 68	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	23	SW
651220	ZED 69	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	23	SE
651221	ZED 70	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	24	SW
651222	ZED 71	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	24	SE
651223	ZED 72	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	19	SW
651224	ZED 73	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	19	SE
651225	ZED 74	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	20	SW
651226	ZED 75	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	20	SE
651227	ZED 76	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	21	SW
651228	ZED 77	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	21	SE
651229	ZED 78	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	22	SW
651230	ZED 79	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	22	SE
651231	ZED 80	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	23	SW
651232	ZED 81	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	23	SE
651233	ZED 82	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	24	SW
651234	ZED 83	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	24	SE
651235	ZED 84	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	15E	19	SW
651236	ZED 85	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	26	NW
651237	ZED 86	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	26	NE
651238	ZED 87	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	25	NW
651239	ZED 88	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	25	NE
651240	ZED 89	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	30	NW
651241	ZED 90	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	30	NE
651242	ZED 91	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	29	NW
651243	ZED 92	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	29	NE
651244	ZED 93	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	28	NW
651245	ZED 94	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	28	NE
651246	ZED 95	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	27	NW
651247	ZED 96	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	27	NE
651248	ZED 97	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	26	NW
651249	ZED 98	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	26	NE
651250	ZED 99	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	25	NW
651251	ZED 100	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	25	NE
651252	ZED 101	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	15E	30	NW
651253	ZED 102	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	26	SW
651254	ZED 103	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	26	SE
651255	ZED 104	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	25	SW
651256	ZED 105	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	13E	25	SE
651257	ZED 106	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	30	SW
651258	ZED 107	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	30	SE
651259	ZED 108	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	29	SW
651260	ZED 109	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	29	SE
651261	ZED 110	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	28	SW
651262	ZED 111	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	28	SE
651263	ZED 112	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	27	SW
651264	ZED 113	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	27	SE
651265	ZED 114	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	26	SW
651266	ZED 115	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	26	SE

651267	ZED 116	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	25	SW
651268	ZED 117	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	14E	25	SE
651269	ZED 118	Granted	September 28, 2005	November 30, 2013	160	Kotzebue	20N	15E	30	SW
655537	ZED 119	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	12E	12	NW
655538	ZED 120	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	12E	12	NE
655539	ZED 121	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	13E	7	NW
655540	ZED 122	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	13E	7	SW
655541	ZED 123	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	13E	7	SE
655542	ZED 124	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	13E	8	SW
655543	ZED 125	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	20N	13E	8	SE
655648	KG 1	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	11	SW
655649	KG 2	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	11	SE
655650	KG 3	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	14	NW
655651	KG 4	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	14	NE
655652	KG 5	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	13	NW
655653	KG 6	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	13	NE
655654	KG 7	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	13	SE
655655	KG 8	Granted	September 8, 2006	November 30, 2013	160	Kotzebue	21N	11E	24	NE
714584	East DH 1	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	15	SE
714585	East DH 2	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	22	NE
714586	East DH 3	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	23	NW
714587	East DH 4	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	22	SE
714588	East DH 5	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	23	SW
714589	COBRE 1	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	2	NW
714590	COBRE 2	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	2	NE
714591	COBRE 3	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	1	NW
714592	COBRE 4	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	1	NE
714593	COBRE 5	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	6	NW
714594	COBRE 6	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	6	NE
714595	COBRE 7	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	5	NW
714596	COBRE 8	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	5	NE
714597	COBRE 9	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	2	SW
714598	COBRE 10	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	2	SE
714599	COBRE 11	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	1	SW
714600	COBRE 12	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	1	SE
714601	COBRE 13	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	6	SW
714602	COBRE 14	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	6	SE
714603	COBRE 15	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	5	SW
714604	COBRE 16	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	5	SE
714605	COBRE 17	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	11	NW
714606	COBRE 18	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	11	NE
714607	COBRE 19	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	12	NW
714608	COBRE 20	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	7E	12	NE
714609	COBRE 21	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	7	NW
714610	COBRE 22	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	7	NE
714611	COBRE 23	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	8	NW
714612	COBRE 24	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	8	NE
714613	COBRE 25	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	9	NW
714614	COBRE 26	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	9	NE
714615	COBRE 27	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	7	SE
714616	COBRE 28	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	8	SW
714617	COBRE 29	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	8	SE
714618	COBRE 30	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	9	SW
714619	COBRE 31	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	9	SE
714620	COBRE 32	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	17	NW
714621	COBRE 33	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	17	NE
714622	COBRE 34	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	16	NW

714623	COBRE 35	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	16	NE
714624	COBRE 36	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	15	NW
714625	COBRE 37	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	15	NE
714626	COBRE 38	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	17	SE
714627	COBRE 39	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	16	SW
714628	COBRE 40	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	16	SE
714629	COBRE 41	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	15	SW
714630	COBRE 42	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	15	SE
714631	COBRE 43	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	21	NW
714632	COBRE 44	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	21	NE
714633	COBRE 45	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	22	NW
714634	COBRE 46	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	22	NE
714635	COBRE 47	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	23	NW
714636	COBRE 48	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	23	NE
714637	COBRE 49	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	21	SW
714638	COBRE 50	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	21	SE
714639	COBRE 51	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	22	SW
714640	COBRE 52	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	22	SE
714641	COBRE 53	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	23	SW
714642	COBRE 54	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	8E	23	SE
714643	West Horse 1	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	12	SW
714644	West Horse 2	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	12	SE
714645	West Horse 3	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	7	SW
714646	West Horse 4	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	7	SE
714647	West Horse 5	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	10E	8	SW
714648	West Horse 6	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	13	NW
714649	West Horse 7	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	13	NE
714650	West Horse 8	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	13	SW
714651	West Horse 9	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	13	SE
714652	West Horse 10	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	24	NW
714653	West Horse 11	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	24	NE
714654	West Horse 12	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	24	SW
714655	West Horse 13	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	24	SE
714656	West Horse 14	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	27	NE
714657	West Horse 15	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	26	NW
714658	West Horse 16	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	26	NE
714659	West Horse 17	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	25	NW
714660	West Horse 18	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	25	NE
714661	West Horse 19	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	27	SE
714662	West Horse 20	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	26	SW
714663	West Horse 21	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	26	SE
714664	West Horse 22	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	25	SW
714665	West Horse 23	Granted	June 21, 2012	November 30, 2013	160	Kotzebue	22N	9E	25	SE
714748	AM 37-225	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	16	SW
714749	AM 38-225	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	16	SW
714750	AM 39-225	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	16	NW
714751	AM 40-225	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	16	NW
714752	AM 41-223	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	8	SE
714753	AM 41-224	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	8	SE
714754	AM 43-219	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	7	NE
714755	AM 43-220	Granted	July 10, 2012	November 30, 2013	40	Kotzebue	20N	12E	7	NE
714756	AM 38-217	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	18	SW
714757	AM 38-219	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	18	SE
714758	AM 38-221	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	17	SW
714759	AM 38-223	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	17	SE
714760	AM 42-219	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	7	SE
714761	AM 42-221	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	8	SW

714762	AM 44-217	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	7	NW
714763	AM 40-217	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	18	NW
714764	AM 40-219	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	18	NE
714765	AM 40-221	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	17	NW
714766	AM 40-223	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	17	NE
714767	AM 42-217	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	20N	12E	7	SW
714768	KG 9	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	21N	12E	19	NW
714769	KG 10	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	21N	12E	19	NE
714770	KG 11	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	21N	12E	19	SW
714771	KG 12	Granted	July 10, 2012	November 30, 2013	160	Kotzebue	21N	12E	19	SE
715147	Cobre 55	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	14	SW
715148	Cobre 56	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	14	SE
715149	Cobre 57	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	13	SW
715150	Cobre 58	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	23	NW
715151	Cobre 59	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	23	NE
715152	Cobre 60	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	24	NW
715153	Cobre 61	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	24	NE
715154	Cobre 62	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	19	NW
715155	Cobre 63	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	23	SE
715156	Cobre 64	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	24	SW
715157	Cobre 65	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	24	SE
715158	Cobre 66	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	19	SW
715159	Cobre 67	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	19	SE
715160	Cobre 68	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	26	NE
715161	Cobre 69	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	25	NW
715162	Cobre 70	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	25	NE
715163	Cobre 71	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	30	NW
715164	Cobre 72	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	30	NE
715165	Cobre 73	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	29	NW
715166	Cobre 74	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	29	NE
715167	Cobre 75	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	26	SE
715168	Cobre 76	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	25	SW
715169	Cobre 77	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	25	SE
715170	Cobre 78	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	30	SW
715171	Cobre 79	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	30	SE
715172	Cobre 80	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	29	SW
715173	Cobre 81	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	29	SE
715174	Cobre 82	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	36	NW
715175	Cobre 83	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	7E	36	NE
715176	Cobre 84	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	31	NW
715177	Cobre 85	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	31	NE
715178	Cobre 86	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	32	NW
715179	Cobre 87	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	32	NE
715180	Cobre 88	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	31	SE
715181	Cobre 89	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	32	SW
715182	Cobre 90	Granted	September 9, 2012	November 30, 2013	160	Kotzebue	23N	8E	32	SE
50-81-0127	USMS2245-1	Issued	July 27, 1981	no expiry	240.02	Kotzebue	21N	11E	34; 35
50-83-0174	USMS2245-2	Issued	June 3, 1983	no expiry	31.91	Kotzebue	20N; 21N	11E	2; 35
					112058
546162	Arctic 3	Relinquished	August 19, 2011		17	Kotzebue	21N	11E	35	SW & NW
F-054378	ARCTIC 5	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	35
F-054379	ARCTIC 6	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	35
F-054381	ARCTIC 8	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	35
F-054382	ARCTIC 12	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	34
F-054383	ARCTIC 14	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	34
F-054384	ARCTIC 16	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	34
F-054385	ARCTIC 18	Relinquished	September 6, 1965		20	Kotzebue	21N	11E	34

F-054386	ARCTIC 20	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054387	ARCTIC 21	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054388	ARCTIC 22	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054392	ARCTIC 33	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	35
F-054418	ARCTIC 300	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054419	ARCTIC 301	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054425	ARCTIC 400	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34
F-054426	ARCTIC 401	Relinquished	September 6, 1965	20	Kotzebue	21N	11E	34