Trilogy Metals (TMQ) 8-K NI 43-101 Technical Report on Pre-Feasibility Study

Exhibit 99.1

 

TRILOGY METALS INC.

Arctic Project,

Northwest Alaska, USA

NI 43-101 Technical Report on Pre-Feasibility Study

Report Prepared For:	Trilogy Metals Inc. 609 Granville Street, Suite 1150 Vancouver, BC V7Y 1G5 Canada Tel: 604-638-8088 Fax: 604-638-0644 www.trilogymetals.com
Report Prepared By:	Paul Staples, P.Eng., Ausenco Engineering Canada Inc. Justin Hannon, P.Eng., Ausenco Engineering Canada Inc. Antonio Peralta Romero, PhD, P.Eng., Amec Foster Wheeler Americas Ltd. Bruce Davis, FAusIMM, BD Resource Consulting, Inc. John J. DiMarchi, CPG, Core Geoscience Inc.
	Jeffrey B. Austin, P.Eng., International Metallurgical & Environmental Inc. Robert Sim, P.Geo., SIM Geological Inc. Calvin Boese, P.Eng., M.Sc., SRK Consulting (Canada) Inc. Bruce Murphy, P.Eng., SRK Consulting (Canada) Inc. Tom Sharp, PhD, P.Eng., SRK Consulting (Canada) Inc.
Effective Date:	February 20, 2018
Release Date:	April 6, 2018

Trilogy Metals Inc.

NI 43-101 Technical Report on the Arctic Project,

Northwest Alaska

 

Important Notice

This report was prepared as National Instrument 43-101 Technical Report for Trilogy Metals Inc. (Trilogy Metals or Trilogy) by Ausenco Engineering Canada Inc., Core Geoscience Inc., Amec Foster Wheeler Americas Limited, BD Resource Consulting, Inc., International Metallurgical & Environmental Inc., SIM Geological Inc., and SRK Consulting (Canada) Inc., collectively the “Report Authors”. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Trilogy Metals subject to the respective terms and conditions of its contracts with the individual Report Authors. Except for the purposes legislated under Canadian provincial securities law, any other uses of this report by any third party is at that party’s sole risk.

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Table of Contents

1.0	Summary			1-16
	1.1	Introduction		1-16
	1.2	Management Property Description and Location		1-16
	1.3	Geology and Mineralization		1-17
	1.4	Drilling and Sampling		1-18
	1.5	Mineral Processing and Metallurgical Testing		1-19
	1.6	Mineral Resource Estimate		1-21
	1.7	Mining Reserves and Mining Methods		1-22
	1.8	Recovery Methods		1-23
	1.9	Project Infrastructure		1-24
	1.10	Market Studies		1-26
	1.11	Environmental, Permitting, Social and Closure Considerations		1-27
		1.11.1	Environmental Considerations	1-27
		1.11.2	Permitting Considerations	1-27
		1.11.3	Social Considerations	1-28
		1.11.4	Closure Planning	1-29
	1.12	Capital Costs		1-29
	1.13	Operating Costs		1-30
	1.14	Economic Analysis		1-31
	1.15	Sensitivity Analysis		1-32
	1.16	Interpretations and Conclusions		1-32
	1.17	Recommendations		1-32
2.0	Introduction			2-1
	2.1	Terms of Reference		2-1
	2.2	Units of Measurement		2-2
	2.3	Qualified Persons		2-2
	2.4	Site Visit		2-2
	2.5	Effective Dates		2-3
	2.6	Information Sources		2-4
	2.7	Previous Technical Reports		2-4
3.0	Reliance on Other Experts			3-1
	3.1	Introduction		3-1
	3.2	Mineral Tenure, Surface Rights, Royalties, Property Agreements		3-1
	3.3	Legal Considerations		3-1
	3.4	Taxation		3-2
	3.5	Marketing and Contracts		3-2
	3.6	Metal Prices and Exchange Rates		3-2
4.0	Property Description and Location			4-1
	4.1	Location		4-1
	4.2	Ownership		4-1
	4.3	Mineral Tenure		4-1
	4.4	Royalties, Agreements and Encumbrances		4-6
		4.4.1	Kennecott Agreements	4-6
		4.4.2	NANA Agreement	4-6

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		4.4.3	South32 Agreement	4-8
		4.4.4	State Royalty	4-8
	4.5	Surface Rights		4-8
	4.6	Environmental Considerations		4-8
	4.7	Permits		4-9
	4.8	Social Considerations		4-9
	4.9	Comment on Section 4		4-9
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	Road	5-1
		5.1.3	Water	5-1
	5.2	Climate		5-1
	5.3	Local Resources and Infrastructure		5-2
	5.4	Physiography		5-2
	5.5	Comment on Section 5		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	6-9
		6.2.3	Geophysics	6-9
		6.2.4	Drilling	6-10
		6.2.5	Specific Gravity	6-10
		6.2.6	Petrology, Mineralogy, and Research Studies	6-10
		6.2.7	Geotechnical, Hydrological and Acid-Base Accounting Studies	6-11
		6.2.8	Metallurgical Studies	6-12
	6.3	Development Studies		6-12
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
	7.6	Deposits and Prospects		7-17
8.0	Deposit Types			8-1
9.0	Exploration			9-1
	9.1	Grids and Surveys		9-3
	9.2	Geological Mapping		9-3
	9.3	Geochemistry		9-6
	9.4	Geophysics		9-9
	9.5	Bulk Density		9-11

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	9.6	Petrology, Mineralogy and Research Studies		9-11
	9.7	Geotechnical, Hydrogeological and Acid Base Accounting Studies		9-11
		9.7.1	Geotechnical and Hydrogeological Assessments	9-11
		9.7.2	Acid-Base Accounting Studies	9-19
		9.7.3	Geochemical Kinetic Studies	9-20
10.0	Drilling			10-1
	10.1	Introduction		10-1
	10.2	Drill Companies		10-3
	10.3	Drill Core Procedures		10-5
		10.3.1	Kennecott	10-5
		10.3.2	NovaGold/Trilogy Metals	10-5
	10.4	Geotechnical Drill Hole Procedures		10-6
	10.5	Collar Surveys		10-7
		10.5.1	Kennecott	10-7
		10.5.2	NovaGold/Trilogy Metals	10-7
	10.6	Downhole Surveys		10-7
	10.7	Recovery		10-8
		10.7.1	Kennecott	10-8
		10.7.2	NovaGold/Trilogy Metals	10-8
	10.8	Drill Intercepts		10-8
	10.9	Prospect Drilling		10-9
11.0	Sample Preparation, Analyses, and Security			11-1
	11.1	Sample Preparation		11-1
	11.2	Core		11-1
		11.2.1	Kennecott and BCMC	11-1
		11.2.2	NovaGold/Trilogy Metals	11-1
	11.3	Acid-Base Accounting Sampling		11-3
	11.4	Density Determinations		11-3
	11.5	Security		11-4
	11.6	Assaying and Analytical Procedures		11-4
	11.7	Quality Assurance/Quality Control		11-5
		11.7.1	Core Drilling Sampling QA/QC	11-5
		11.7.2	Acid-Base Accounting Sampling QA/QC	11-15
		11.7.3	Density Determinations QA/QC	11-15
	11.8	QP’s Opinion		11-18
12.0	Data Verification			12-1
	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		12-2
	12.6	QP Opinion		12-12
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	Historical Test Work Review	13-3
		13.1.3	Mineralogical and Metallurgical Test Work – 2012 to 2017	13-8
	13.2	Recommended Test Work		13-25

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14.0	Mineral Resource Estimate			14-1
	14.1	Introduction		14-1
	14.2	Sample Database and Other Available Data		14-1
		14.2.1	ABA Data	14-4
	14.3	Geologic Model		14-6
		14.3.1	Summary of Geologic Domains	14-9
	14.4	Compositing		14-10
	14.5	Exploratory Data Analysis		14-11
		14.5.1	As-Logged Geology and Domain Statistics	14-11
		14.5.2	Interpreted Lithology and MinZone Domain Statistics	14-13
		14.5.3	Contact Profiles	14-19
		14.5.4	Modeling Implications	14-21
	14.6	Treatment of Outlier Grades		14-22
	14.7	Specific Gravity Data		14-23
	14.8	Variography		14-24
	14.9	Model Setup and Limits		14-28
	14.10	Interpolation Parameters		14-29
	14.11	Block Model Validation		14-32
		14.11.1	Visual Inspection	14-32
		14.11.2	Model Checks for Change of Support	14-33
		14.11.3	Comparison of Interpolation Methods	14-35
		14.11.4	Swath Plots (Drift Analysis)	14-38
	14.12	Resource Classification		14-42
	14.13	Mineral Resource Estimate		14-43
	14.14	Grade Sensitivity Analysis		14-46
15.0	Mineral Reserve Estimates			15-1
	15.1	Overview		15-1
	15.2	Pit Optimization		15-1
	15.3	Dilution and Ore Losses		15-4
	15.4	Mineral Reserve Statement		15-5
	15.5	Factors Affecting Mineral Reserves		15-6
16.0	Mining Methods			16-1
	16.1	Overview		16-1
	16.2	Mine Design		16-1
	16.3	Waste Rock Facilities and Stockpile Designs		16-3
	16.4	Production Schedule		16-5
	16.5	Waste Material Handling		16-7
	16.6	Operating Schedule		16-8
	16.7	Mining Equipment		16-9
		16.7.1	Blasting	16-10
		16.7.2	Drilling	16-11
		16.7.3	Loading	16-13
		16.7.4	Hauling	16-13
		16.7.5	Support	16-14
		16.7.6	Auxiliary	16-16
	16.8	Open Pit Water Management		16-17
	16.9	Geotechnical Review		16-17
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

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		17.1.4	Grinding and Classification	17-5
		17.1.5	Flotation	17-6
	17.2	Plant Process Control		17-14
		17.2.1	Overview	17-14
18.0	Project InfrastrUcture			18-1
	18.1	Introduction		18-1
	18.2	Access Roads		18-4
		18.2.1	Ambler Mining District Industrial Access Project Road	18-4
		18.2.2	Arctic Access Road	18-5
	18.3	Airstrip		18-5
	18.4	Camps		18-6
		18.4.1	Bornite Exploration Camp	18-6
		18.4.2	Temporary Construction Camp	18-6
		18.4.3	Construction / Operations Permanent Camp	18-6
	18.5	Fuel Supply, Storage and Distribution		18-6
	18.6	Power Generation		18-7
	18.7	Electrical System		18-7
	18.8	Surface Water Management		18-7
		18.8.1	Process Water Supply	18-9
		18.8.2	Water Management Infrastructure	18-9
		18.8.3	Waste Rock Collection Pond	18-10
		18.8.4	Site Water and Load Balance	18-10
	18.9	Water Treatment Plant		18-11
	18.10	Tailings Management Facility		18-11
		18.10.1	General Description	18-11
		18.10.2	Design Criteria	18-12
		18.10.3	Overburden Geotechnical Investigation	18-13
		18.10.4	Site Selection	18-13
		18.10.5	Starter Dam	18-13
		18.10.6	Dam Raises and Final Dam	18-14
		18.10.7	TMF Water Pool and Water Return	18-15
	18.11	Tailings Delivery and Return System		18-15
	18.12	Waste Rock Dump and Overburden Stockpiles		18-15
		18.12.1	Waste Rock Dump	18-17
		18.12.2	Overburden and Topsoil Stockpiles	18-17
	18.13	Compressed Air Supply		18-17
	18.14	Site Communications		18-18
	18.15	Fire Protection		18-18
	18.16	Plant Buildings		18-18
		18.16.1	Gatehouse	18-18
		18.16.2	Mine Infrastructure Area	18-18
		18.16.3	Laboratory	18-19
		18.16.4	Administration Building	18-19
		18.16.5	Mill Dry Facility	18-19
		18.16.6	Plant Workshop and Warehouse	18-19
		18.16.7	Primary Crushing	18-19
		18.16.8	Fine Ore Stockpile	18-19
		18.16.9	Process Plant	18-20
		18.16.10	Concentrate Loadout	18-20
		18.16.11	Reagent Storage and Handling	18-20
		18.16.12	Raw Water Supply	18-20
	18.17	Concentrate Transportation		18-20

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19.0	Market Studies and Contracts			19-1
	19.1	Metal Prices		19-1
	19.2	Markets and Contracts		19-1
	19.3	Smelter Term Assumptions		19-2
	19.4	Transportation and Logistics		19-3
	19.5	Insurance		19-3
	19.6	Representation and Marketing		19-3
	19.7	QP’s Opinion		19-4
20.0	Environmental Studies, permitting and social or community Impact			20-1
	20.1	Environmental Studies		20-1
		20.1.1	Hydrology	20-1
		20.1.2	Water Quality	20-3
		20.1.3	Wetlands Data	20-3
		20.1.4	Aquatic Life Data	20-4
		20.1.5	Hydrogeology Data	20-6
		20.1.6	Cultural Resources Data	20-7
		20.1.7	Subsistence Data	20-8
		20.1.8	Endangered Species, Migratory Birds, and Bald and Golden Eagle protection	20-9
		20.1.9	Acid Base Accounting Data	20-10
		20.1.10	Additional Baseline Data requirements	20-10
	20.2	Permitting		20-11
		20.2.1	Exploration Permits	20-11
		20.2.2	Major Mine Permits	20-11
	20.3	Social or Community Considerations		20-13
	20.4	Mine Reclamation and Closure		20-15
		20.4.1	Reclamation and Closure Plan	20-15
		20.4.2	Reclamation and Closure Financial Assurance	20-20
21.0	Capital and Operating Costs			21-1
	21.1	Capital Cost Summary		21-1
		21.1.1	Capital Cost Introduction	21-1
		21.1.2	Project Execution	21-2
		21.1.3	Work Breakdown Structure	21-2
		21.1.4	Estimate Summary	21-2
		21.1.5	Definition	21-4
		21.1.6	Road Construction	21-5
		21.1.7	Basis of Mining Capital Cost Estimate	21-6
		21.1.8	Mining Capital and Sustaining Capital Costs	21-6
		21.1.9	Tailings Management Facility	21-7
		21.1.10	Water Treatment Plant	21-7
		21.1.11	Sustaining Capital and Closure Costs Summary	21-8
	21.2	Operating Cost Estimate		21-8
		21.2.1	Operating Cost Summary	21-8
		21.2.2	Mining Operating Cost Estimate	21-9
		21.2.3	Processing Operating Cost Estimate	21-10
		21.2.4	General and Administrative and Surface Services Cost Estimates	21-10
		21.2.5	Road Toll Cost Estimate	21-12
22.0	Economic Analysis			22-1
	22.1	Introduction		22-1
	22.2	Cautionary Statement and Forward Looking Information		22-1
	22.3	Inputs to the Cash Flow Model		22-1

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	22.4	Basis of Pre-Tax Financial Evaluation		22-3
	22.5	Pre-Tax Financial Results		22-4
	22.6	Post-Tax Financial Analysis		22-6
		22.6.1	US Federal Tax	22-6
		22.6.2	Alaska State Tax	22-6
		22.6.3	Alaska Mining License Tax	22-7
		22.6.4	Post-Tax Financial Results	22-7
	22.7	Cash Flow		22-8
	22.8	Sensitivity Analysis		22-9
	22.9	Copper and Zinc Metal Price Scenarios		22-13
23.0	Adjacent Properties			23-1
24.0	Other Relevant Data and Information			24-1
25.0	Interpretation and Conclusions			25-1
	25.1	Introduction		25-1
	25.2	Mineral Tenure, Surface Rights, Royalties and Agreements		25-1
	25.3	Geology and Mineralization		25-2
	25.4	Exploration, Drilling and Analytical Data Collection in Support of Mineral Resource Estimation		25-2
	25.5	Metallurgical Testwork		25-2
	25.6	Mineral Resource Estimates		25-3
	25.7	Mineral Reserves and Mine Planning		25-3
	25.8	Recovery Plan		25-3
	25.9	Project Infrastructure		25-4
	25.10	Environmental, Permitting and Social		25-4
	25.11	Markets and Contracts		25-4
	25.12	Capital Costs		25-4
	25.13	Operating Costs		25-5
	25.14	Economic Analysis		25-5
	25.15	Conclusions		25-5
26.0	Recommendations			26-1
	26.1	Introduction		26-1
	26.2	Work Program		26-1
27.0	References			27-1
28.0	Certificates of Qualified Persons			28-1
	28.1	Paul Staples, P.Eng.		28-1
	28.2	Justin Hannon, P.Eng.		28-2
	28.3	Antonio Peralta Romero, P.Eng.		28-3
	28.4	Bruce M. Davis, FAusIMM		28-4
	28.5	John Joseph DiMarchi, CPG		28-5
	28.6	Jeffery B. Austin, P.Eng.		28-6
	28.7	Robert Sim, P.Geo.		28-7
	28.8	Calvin Boese, PEng.		28-8
	28.9	Bruce Murphy, PEng.		28-9
	28.10	Tom Sharp, PEng.		28-10

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List of Appendices

Appendix A – List of Claims

List of Tables

Table 1-1	Summary of Overall Metal Recovery – Arctic Project	1-21
Table 1-2	Mineral Resource Estimate for the Arctic Deposit	1-22
Table 1-3	Mineral Reserve Estimates for the Arctic Deposit	1-22
Table 1-4	Initial Capital Costs	1-29
Table 1-5	Sustaining Capital and Closure Costs	1-30
Table 1-6	Sustaining Capital and Closure Costs	1-30
Table 6-1	Known Mapping, Geochemical, and Geophysical Programs Targeting VMS Prospects in the Ambler Mining District	6-4
Table 6-2	Mining and Technical Studies	6-12
Table 9-1	Summary of Trilogy/NovaGold Exploration Activities Targeting VMS-style Mineralization in the Ambler Sequence Stratigraphy and the Arctic Deposit	9-2
Table 9-2	TDEM Loops and Locations	9-9
Table 9-3	Summary of derived rock mass parameter values per rock mass domain.	9-15
Table 9-4	Selected acceptance criteria	9-17
Table 9-5	Summary of slope modelling results	9-18
Table 10-1	Companies, Campaigns, Drill Holes and Metres Drilled at the Arctic Deposit	10-1
Table 10-2	Summary of Trilogy/NovaGold Drilling	10-2
Table 10-3	Drill Contractors, Drill Holes, Meterage and Core Sizes by Drill Campaign at the Arctic Deposit	10-4
Table 10-4	Recovery and RQD 2004 to 2008 Arctic Drill Campaigns	10-8
Table 10-5	Drill, Meterage and Average Drill Depth for Trilogy Ambler Sequence VMS Targets	10-9
Table 10-6	Trilogy Metals Exploration Drilling – Ambler Schist Belt	10-11
Table 11-1	Analytical Laboratories Used by Operators of the Arctic Project	11-5
Table 13-1	Metallurgical Test Work Programs	13-1
Table 13-2	Metallic Mineral Identified in Arctic Project Samples	13-3
Table 13-3	Bond Ball Mill Work Index	13-5
Table 13-4	Head Analyses	13-6
Table 13-5	Flotation Test on Ambler Low Talc Composite	13-7
Table 13-6	Head Grades – Composite Samples – 2012	13-9
Table 13-7	Head Grade 2017 Variability Samples and Pilot Plant Composite	13-10
Table 13-8	Mineral Modal Abundance for Composite Samples – 2012	13-11
Table 13-9	Bond Ball Mill Grindability and Abrasion Index Test Results	13-13
Table 13-10	Locked Cycle Metallurgical Test Results	13-17
Table 13-11	SGS Open Circuit Copper and Lead Separation Test Results	13-20
Table 13-12	ALS Metallurgy Locked Cycle Copper-Lead Separation Test Results	13-22
Table 13-13	Summary of Overall Metal Recovery – Arctic Deposit	13-23

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Table 13-14	Multi-element Assay Results –Lead Concentrate and Copper Concentrate	13-24
Table 13-15	Multi-element Assay Results – Zinc Concentrate	13-24
Table 14-1	Summary of Sample Data Used to Develop the Resource Block Model	14-6
Table 14-2	Summary of Lithology Domains	14-9
Table 14-3	Summary of Mineralized Zone (MinZone) Domains	14-10
Table 14-4	Summary of Geotech, Alteration, Talc and Weathering Domains	14-10
Table 14-5	Summary of Estimation Domains	14-22
Table 14-6	Summary of Treatment of Outlier Sample Data	14-23
Table 14-7	Copper Correlogram Parameters	14-24
Table 14-8	Lead Correlogram Parameters	14-25
Table 14-9	Zinc Correlogram Parameters	14-25
Table 14-10	Gold Correlogram Parameters	14-26
Table 14-11	Silver Correlogram Parameters	14-26
Table 14-12	Sulphur Correlogram Parameters	14-27
Table 14-13	AP Correlogram Parameters	14-27
Table 14-14	NP Correlogram Parameters	14-28
Table 14-15	Block Model Limits	14-28
Table 14-16	Interpolation Parameters for Copper	14-29
Table 14-17	Interpolation Parameters for Lead	14-29
Table 14-18	Interpolation Parameters for Zinc	14-30
Table 14-19	Interpolation Parameters for Gold	14-30
Table 14-20	Interpolation Parameters for Silver	14-30
Table 14-21	Interpolation Parameters for Sulphur	14-31
Table 14-22	Interpolation Parameters for AP	14-31
Table 14-23	Interpolation Parameters for NP	14-31
Table 14-24	Interpolation Parameters for Specific Gravity	14-32
Table 14-25	Parameters Used to Generate a Resource-Limiting Pit Shell	14-44
Table 14-26	Mineral Resource Estimate for the Arctic Deposit	14-45
Table 14-27	Sensitivity of Mineral Resource to Cut-off Grade	14-47
Table 15-1	Optimization Inputs	15-2
Table 15-2	Mineral Reserve Statement	15-5
Table 16-1	Mine Design Parameters	16-1
Table 16-2	Production Schedule	16-6
Table 16-3	Gross Operating Hours per Year	16-8
Table 16-4	Productive Utilization Ramp-up	16-9
Table 16-5	Availability and Productive Utilization Post Ramp-up	16-9
Table 16-6	Blasting Design Input	16-10
Table 16-7	Blast Designs	16-11
Table 16-8	Rock Type Weight and UCS	16-11
Table 16-9	PV271 Drill Penetration Rates	16-12
Table 16-10	Drill Requirements and Performance	16-12
Table 16-11	Truck Requirements & Productivity Statistics	16-14
Table 16-12	Support Equipment	16-16
Table 16-13	Auxiliary Equipment	16-16
Table 17-1	Processing Facility Design Criteria	17-4
Table 18-1	TMF Design Parameters and Design Criteria	18-12
Table 18-2	Mode of Transport and Distances for Concentrate Shipping	18-21
Table 19-1	Concentrate Transport Costs	19-3
Table 20-1	Major Mine Permits Required for the Arctic Project	20-13
Table 20-2	Summary of Closure and Reclamation Costs	20-19
Table 21-1	Estimate Summary Level 1 Major Facility	21-3
Table 21-2	Initial Estimate by Major Discipline	21-4
Table 21-3	Estimate Exchange Rates	21-5
Table 21-4	Mine Capital Costs	21-7

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Table 21-5	Sustaining Capital and Closure Costs	21-8
Table 21-6	Overall Operating Cost Estimate	21-9
Table 21-7	Life of Mine Mining Cost	21-9
Table 21-8	Summary of Processing Operating Cost Estimates	21-10
Table 21-9	G&A Cost Estimates	21-11
Table 21-10	Surface Services Cost Estimates	21-12
Table 22-1	Mine and Payable Metal Production for the Arctic Mine	22-3
Table 22-2	Summary of Pre-Tax Financial Results	22-5
Table 22-3	Summary of Post-Tax Financial Results	22-8
Table 22-4	Pre and Post-Tax Arctic Project Production and Cash Flow Forecast	22-9
Table 22-5	Pre-tax Copper Price Scenarios	22-13
Table 22-6	Pre-tax Zinc Price Scenarios	22-13

List of Figures

Figure 2-1	Property Location Map (Tetra Tech, 2013)	2-1
Figure 4-1	Upper Kobuk Mineral Projects Lands (Trilogy Metals, 2017)	4-2
Figure 4-2	Arctic Project Mineral Tenure Plan (Trilogy Metals, 2017)	4-3
Figure 4-3	Mineral Tenure Layout Plan (Trilogy Metals, 2018)	4-4
Figure 4-4	Arctic Deposit Location (Trilogy Metals, 2018)	4-5
Figure 7-1	Geologic Terranes of the Southern Brooks Range (Trilogy Metals, 2017)	7-2
Figure 7-2	Geology of the Ambler Mining District (Trilogy Metals, 2017)	7-5
Figure 7-3	Ambler Sequence Stratigraphy in the Arctic Deposit Area (Trilogy Metals, 2017)	7-7
Figure 7-4	Generalized Geology of the Central Ambler District (Trilogy Metals, 2017)	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 (Trilogy Metals, 2017)	7-11
Figure 7-7	Typical Massive Sulphide Mineralization at the Arctic Deposit	7-16
Figure 7-8	Major Prospects of the Ambler Mining District (Trilogy Metals, 2017)	7-17
Figure 9-1	Mapping Campaigns in and around the Arctic Deposit (Trilogy Metals, 2017)	9-4
Figure 9-2	Arctic Deposit Area Geology (Trilogy Metals, 2017)	9-5
Figure 9-3	2016 Updated Arctic Surface Geology Map (Trilogy Metals, 2017)	9-6
Figure 9-4	Copper Distribution in Silt and Soil Samples in the Dead Creek Area	9-7
Figure 9-5	Zinc Distribution in Silt and Soil Samples in the Dead Creek Deposit Area	9-8
Figure 9-6	TDEM Loops and Contoured Resistivity – Dead Creek Prospect	9-10
Figure 9-7	SRK Structural Model used in the Slope Stability Analysis	9-13
Figure 9-8	Six structural and geomechanical domains were identified.	9-14
Figure 9-9	SRK Preliminary Hydrology Model – talc Confinement Model	9-16
Figure 9-10	SRK Design Sectors and the Recommended Range of Inter-Ramp Angles to be used in the Slope Design	9-19
Figure 10-1	Plan Map of Drill Holes in the Vicinity of the Arctic Deposit (Trilogy Metals, 2017)	10-3
Figure 10-2	Collar Locations and Principal Target Areas – Ambler District (Trilogy Metals, 2017)	10-10
Figure 10-3	Sunshine Prospect and Drill Hole Locations (Trilogy Metals, 2017)	10-11
Figure 11-1	Spatial Availability of QA/QC Data (Trilogy Metals, 2017)	11-6
Figure 11-2	Graph Showing Good Agreement between Wet-dry Measured Specific Gravity and Pycnometer Measured Specific Gravity	11-16

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Figure 11-3	Measured versus Stoichiometric Specific Gravities	11-17
Figure 11-4	Scatter Plot Showing the Measured Specific Gravity versus Multiple (Copper, Iron, Zinc, Barium) Regression Estimate	11-18
Figure 12-1	Distribution of the Differences Between GPS Elevations and the DTM	12-1
Figure 13-1	Proposed Copper-Lead-Zinc Flowsheet Showing Talc Pre-float	13-2
Figure 14-1	Isometric View of Copper Grades in Drill Holes (Sim, 2017)	14-3
Figure 14-2	Isometric View of Copper Grades in Drill Holes (Sim, 2017)	14-4
Figure 14-3	Isometric Views of Available AP and NP Data (Sim, 2017)	14-5
Figure 14-4	Isometric View of Available Sulphur Data (Sim,2017)	14-5
Figure 14-5	Cross Section 613250E Showing Lithology Domains at Arctic (Sim, 2017)	14-7
Figure 14-6	Cross Section 7453000N Showing Lithology Domains at Arctic (Sim, 2017)	14-7
Figure 14-7	Isometric View of Geotechnical Domains (Sim, 2017)	14-8
Figure 14-8	Isometric Views of Talc Domains (Sim, 2017)	14-8
Figure 14-9	Isometric Views of Alteration Domains (Sim, 2017)	14-9
Figure 14-10	Boxplots of Copper by Logged Lithology Type (Sim, 2017)	14-12
Figure 14-11	Boxplots of Copper by Lithology Domain (Sim, 2017)	14-13
Figure 14-12	Boxplots of Gold by Lithology Domain (Sim, 2017)	14-14
Figure 14-13	Boxplots of Copper by MinZone Domain (Sim, 2017)	14-15
Figure 14-14	Boxplots of AP, NP and Sulphur by Lithology Domain (Sim, 2017)	14-17
Figure 14-15	Boxplots of AP, NP and Sulphur by Talc Domain (Sim, 2017)	14-18
Figure 14-16	Boxplots of AP, NP and Sulphur by Weathered Domain (Sim, 2017)	14-18
Figure 14-17	Boxplots of SG by MinZone and Lithology Group Domains (Sim, 2017)	14-19
Figure 14-18	Contact Profiles of Copper Between MinZone and other Lithology Domain Groups (Sim, 2017)	14-20
Figure 14-19	Contact Profile of AP, NP and Sulphur Between Weathered and Fresh Rocks (Sim, 2017)	14-20
Figure 14-20	Contact Profile of AP, NP and Sulphur Inside / Outside of the Talc Domains (Sim, 2017)	14-21
Figure 14-21	North-South Vertical Section of Copper Estimates in the Block Model (Section 613250E) (Sim, 2017)	14-32
Figure 14-22	West-East Vertical Section of Copper Estimates in the Block Model (Section 7453000N) (Sim, 2017)	14-33
Figure 14-23	Herco and Model Grade / Tonnage Plots for Copper in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-34
Figure 14-24	Herco and Model Grade / Tonnage Plots for Lead in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-34
Figure 14-25	Herco and Model Grade / Tonnage Plots for Zinc in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-34
Figure 14-26	Herco and Model Grade / Tonnage Plots for Gold in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-35
Figure 14-27	Herco and Model Grade / Tonnage Plots for Silver in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-35
Figure 14-28	Comparison of Copper Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-36
Figure 14-29	Comparison of Lead Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-36
Figure 14-30	Comparison of Zinc Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-37
Figure 14-31	Comparison of Gold Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-37
Figure 14-32	Comparison of Silver Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-38
Figure 14-33	Swath Plot of Copper in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-39

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Figure 14-34	Swath Plot of Lead in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-39
Figure 14-35	Swath Plot of Zinc in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-40
Figure 14-36	Swath Plot of Gold in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-40
Figure 14-37	Swath Plot of Silver in MinZone Domains 1, 3 and 5 (Sim, 2017)	14-41
Figure 14-38	Swath Plot of AP in Rocks Outside of the MinZone Domains (Sim, 2017)	14-41
Figure 14-39	Swath Plot of NP in Rocks Outside of the MinZone Domains (Sim, 2017)	14-42
Figure 14-40	Swath Plot of Sulphur Rocks Outside of the MinZone Domains (Sim, 2017)	14-42
Figure 14-41	Isometric Views of Arctic Mineral Resource (Sim, 2017)	14-46
Figure 15-1	Pit-by-Pit Analysis	15-3
Figure 15-2	Selected Pit Shell	15-4
Figure 15-3	Contact Dilution Estimation Procedure	15-5
Figure 16-1	Ultimate Pit Design	16-2
Figure 16-2	Section 1 Showing Mine Design and Selected Pit Shell (looking West)	16-2
Figure 16-3	Section 2 Showing Mine Design and Selected Pit Shell (looking North-West)	16-3
Figure 16-4	Waste Rock Dump	16-4
Figure 16-5	Ore Stockpile	16-5
Figure 16-6	Production Schedule	16-6
Figure 16-7	Scheduled Cu Feed Grade	16-7
Figure 16-8	Stockpile Balance	16-7
Figure 17-1	Simplified Process Flowsheet	17-3
Figure 18-1	Proposed Location of the Processing Plant and Other Buildings (Ausenco, 2018)	18-2
Figure 18-2	Proposed Site Layout (Ausenco, 2018)	18-3
Figure 18-3	Proposed Route of AMDIAP Road (Ambler Access Website 2018)	18-4
Figure 18-4	Arctic Access Road (AllNorth 2017)	18-5
Figure 18-5	Surface Water Management Plan during Operations	18-8
Figure 18-6	Cross Section through the TMF Starter Dam to Elevation 830 m	18-14
Figure 18-7	Cross Section of the TMF and WRD at Final Design Elevation	18-15
Figure 18-8	General Location of WRD and Stockpiles	18-16
Figure 20-1	Current Water Quality and Hydrology Stations Location Map	20-6
Figure 20-2	2016 Aquatics Survey Sample Sites	20-10
Figure 22-1	Pre-tax NPV Sensitivity Analysis (Ausenco, 2018)	22-2
Figure 22-2	Pre-tax IRR Sensitivity Analysis (Ausenco, 2018)	22-2

Glossary

Acme Analytical Laboratories Ltd.	AcmeLabs
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 and Export Authority	AIDEA
Alaska Native Claims Settlement Act	ANCSA
Alaska Native Regional Corporations	ANCSA Corporations
Ambler Mining District Industrial Access Project	AMDIAP

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Andover Mining Corp.	Andover
Annual Hardrock Exploration Activity	AHEA
atomic absorption	AA
atomic absorption spectroscopy	AAS
atomic emission spectroscopy	ICP_AES
Audio-Frequency Magneto-Telluric	AMT
BD Resource Consulting, Inc.	BDRC
Bear Creek Mining Corporation	BCMC
Arctic Property	the Property
Canadian Institute of Mining, Metallurgy, and Petroleum	CIM
Circular corrugated steel pipes	CSP
complex resistivity induced polarization	CRIP
Controlled Source Audio-frequency Magneto-Telluric	CSAMT
Electromagnetic	EM
Environmental Impact Statement	EIS
Environmental Protection Agency	EPA
Exploration Agreement and Option to Lease	NANA Agreement
Fugro Airborne Surveys	Fugro
GeoSpark Consulting Inc.	GeoSpark
General and Administrative.	G&A
High density sludge	HDS
inductively coupled plasma	ICP
inductively coupled plasma-mass	ICP-MS
Internal Rate of Return	IRR
International Organization for Standardization	ISO
Kennecott Exploration Company and Kennecott Arctic Company	Kennecott
Kennecott Research Centre	KRC
LiDAR	Light Detection and Ranging
liquefied natural gas	LNG
life of mine	LOM
Mine Development Associates	MDA
meters above sea level	masl
NANA Regional Corporation, Inc.	NANA
National Environmental Policy Act	NEPA
National Instrument 43-101	NI 43-101
natural source audio-magnetotelluric	NSAMT
naturally occurring asbestos	NOA
net present value	NPV
net smelter return	NSR
North American Datum	NAD
Northern Land Use Research Inc.	NLUR Inc.
Northwest Arctic Borough	NWAB
Trilogy Metals Inc.	Trilogy Metals
NovaGold Resources Inc.	NovaGold
Polarized Light Microscopy	PLM
Quality Assurance/Quality Control	QA/QC

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SIM Geological Inc.	SGI
single point	SP
Teck Resources Ltd.	Teck
Tailings management facility	TMF
Universal Transverse Mercator	UTM
US Army Corps of Engineers	USACE
US Geological Survey	USGS
volcanogenic massive sulphide	VMS
WH Pacific, Inc.	WHPacific
Waste rock collection pond	WCRP
Waste rock dump	WRD
Water treatment plant	WTP
Zonge International Inc.	Zonge

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1.0 Summary

1.1 Introduction

Trilogy Metals Inc. (Trilogy Metals or Trilogy) commissioned Ausenco Engineering Canada Inc. (Ausenco) to compile a Technical Report (the Report) on the results of a Pre-Feasibility Study on the Arctic Deposit, part of the Arctic Project (the Project) in the Ambler Mining District of Northwest Alaska

The Report supports disclosure by Trilogy Metals in the news release dated 20 February 2018, entitled “Trilogy Metals Announces Pre-Feasibility Study Results and Reserves for the Arctic Project, Alaska”.

The firms and consultants who are providing Qualified Persons (QPs) responsible for the content of this Report, which is based on the Pre-Feasibility Study completed in 2018 (the 2018 PFS) and supporting documents prepared for the 2018 PFS, are, in alphabetical order, Amec Foster Wheeler Americas Ltd. (Amec Foster Wheeler); BD Resource Consulting, Inc., (BDRC); SRK Consulting (Canada) Inc. (SRK), and SIM Geological Inc. (SIM).

All amounts are in US dollars unless otherwise stated.

1.2 Management Property Description and Location

The Arctic Project (Property or Project) is located in the Ambler mining district (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 Project is located 270 km east of the town of Kotzebue, 36 km north of the village of Kobuk, and 260 km west of the Dalton Highway, an all-weather state-maintained highway.

NovaGold Resources Inc. (NovaGold) acquired the Arctic Project from Kennecott Exploration Company and Kennecott Arctic Company (collectively, Kennecott) in 2004. In 2012, NovaGold transferred all copper projects to NovaCopper Inc. NovaCopper Inc. subsequently underwent a name change to Trilogy Metals Inc. in 2016. The Project comprises approximately 46,336 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 1,386 contiguous claims, including 883 40-acre State claims, 503 160-acre State claims, and eighteen Federal patented claims comprising 272 acres (110 ha) held in the name of NovaCopper US Inc., a wholly owned subsidiary of Trilogy Metals.

Surface use of the private land held as Federal patented claims is limited only by reservations in the patents and by generally-applicable environmental laws. Surface use of State claims allows the owner of the mining claim to make such use of the surface as is “necessary for prospecting for, extraction of, or basic processing of minerals.”

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Under the Kennecott Purchase and Termination Agreement, Kennecott retained a 1% net smelter return (NSR) royalty that is purchasable at any time by Trilogy Metals for a one-time payment of $10 million.

The NANA Regional Corporation, Inc. (NANA) controls lands granted under the Alaska Native Claims Settlement Act (ANCSA) to the south of the Project boundary. Trilogy Metals and NANA have entered into an agreement (the NANA Agreement) that consolidates Trilogy Metals’ 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 has a term of 20 years, with an option in favour of Trilogy Metals to extend the term for an additional 10 years. If, following receipt of a feasibility study and the release for public comment of a related draft environmental impact statement, Trilogy Metals decides to proceed with construction of a mine on the lands subject to the NANA Agreement, 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 Trilogy Metals from such project. 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 electing to participate between 16% to 25%, and Trilogy Metals owning the balance of the interest in the joint venture. If Trilogy Metals 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 Trilogy Metals which will afford Trilogy Metals access to the project along routes approved by NANA. In consideration for the grant of such surface use rights, Trilogy Metals will grant NANA a 1% net smelter royalty on production and provide an annual payment on a per acre basis.

Trilogy Metals has entered into an option agreement with South32 Limited (South32) whereby South32 has the right to form a 50/50 Joint Venture with respect to the Trilogy Metals’ Alaskan assets including the Arctic Project.  Upon exercise of the option, Trilogy Metals will transfer its Alaskan assets, including the Arctic Project, and South32 will contribute a minimum of $150 million, to a newly formed joint venture.

1.3 Geology and Mineralization

The Arctic Deposit is considered to be a volcanogenic massive sulphide (VMS) deposit.

The Ambler mining district is located on the southern margin of the Brooks Range. Within the VMS belt, several 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.

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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 occurs within eight modelled SMS and MS 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 is predominately coarse-grained sulphides consisting mainly of chalcopyrite, sphalerite, galena, tetrahedrite-tennantite, pyrite, arsenopyrite, and pyrrhotite. Trace amounts of electrum are also present.

1.4 Drilling and Sampling

Drilling at the Arctic Deposit and within the Ambler District has been ongoing since its initial discovery in 1967. Approximately 56,480 m of drilling has been completed within the Ambler District, including 39,323 m of drilling in 174 drill holes at the Arctic deposit or on potential extensions in 27 campaigns spanning 50 years. Drill programs were completed by Kennecott and its subsidiaries, Anaconda, and Trilogy Metals and its predecessor companies.

Core recoveries are acceptable. Geological and geotechnical logging is in line with industry generally-accepted practices. Drill collar and downhole survey data were collected using industry-recognised instrumentation and methods.

Between 2004 and 2006, NovaGold conducted a systematic drill core re-logging and re-sampling campaign of Kennecott and BCMC era drill holes. 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. During the Trilogy Metals campaigns, 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.

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Gold assays were determined using fire analysis followed by an atomic absorption spectroscopy (AAS) finish. An additional 49-element suite was assayed by inductively coupled plasma-mass spectroscopy (ICP-MS) methodology, following nitric acid aqua regia digestion. The copper, zinc, lead, and silver analyses were completed by atomic absorption (AA), following a triple acid digest, when overlimits.

Standard reference materials, blanks, duplicates and check samples have been regularly submitted at a combined level of 20% of sampling submissions for all NovaGold/NovaCopper/Trilogy Metals era campaigns. BDRC 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.

Specific gravity (SG) measurements have been conducted on 3,023 samples in the database and range from a minimum of 2.43 to a maximum of 4.99 and average 3.08. The distribution of SG data is considered sufficient to support estimation in the resource model.

An aerial LiDAR survey was completed to support pre-feasibility level resource estimation, engineering design, environmental studies, and infrastructure layout evaluations. Agreement between surveyed drill hole collar elevations and a LIDAR topographic surface verifies the correctness of the digital topography for use in estimation.

It was concluded that the drill database and topographic surface for the Arctic Deposit is reliable and sufficient to support the current estimate of mineral resources.

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%.

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· 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.

· 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.  In this test work program, it appeared that most of the gold reported to the copper concentrate and on average the silver was equally recovered into the copper and lead concentrates. Subsequent test work to better define the copper and lead separation process was conducted in 2017, including a more detailed evaluation of the precious metal deportment in the copper and lead separation process.

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 indicate 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.

In 2017, ALS Metallurgy conducted detailed copper and lead separation flotation test work using a bulk sample of copper-lead concentrate produced from the operation of a pilot plant.  This test work confirmed high lead recoveries in locked cycle testing of the copper-lead separation process and confirmed precious metal recoveries into the representative copper and lead concentrates. This test work indicated a clear tendency of the gold values to follow the lead concentrate, giving it a significant gold grade and value.

The conclusions of test work conducted both in 2012 and 2017 indicate that the Arctic materials are well-suited to the production of high-quality copper and zinc concentrates using flotation techniques which are industry standard.  Copper and zinc recovery data is reported in the range of 91 to 89% respectively, which reflects the high grade nature of the deposit as well as the coarse grained nature of these minerals.  Lead concentrates have the potential to be of high quality and can also be impacted by zones of very high talc contents which has the potential to dilute lead concentrate grades.  The lead concentrate is also shown to be rich in precious metals, which has some advantages in terms of marketability of this material.

An overall metallurgical balance for the project is summarized in Table 1-1.  This table of metal recoveries is based on an expected average recovery over the entire resource based on grades and detailed results of metallurgical test work conducted in 2012 and 2017. 

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Table 1-1        Summary of Overall Metal Recovery – Arctic Project

		Concentrate Grade					Metal Recoveries
Process stream	Mass %	Cu %	Pb %	Zn %	Au g/t	Ag g/t	Cu %	Pb %	Zn %	Au %	Ag %
Process Feed	100.0	2.31	0.59	3.22	0.49	38
Copper Conc	7.15	29.5	0.75	3.0	0.35	240	91.2	8.7	5.7	5.2	45.1
Lead Conc	1.02	1.7	50.0	0.9	28.0	1300	0.7	80.0	0.3	58.9	34.9
Zinc Conc	4.85	1.7	0.5	59.2	0.55	49.6	3.6	4.0	91.0	5.5	6.3
Process Tailings	86.98	0.12	0.05	0.15	0.17	6	4.5	7.3	3.0	30.5	13.7

1.6 Mineral Resource Estimate

The mineral resource estimate has been prepared by Robert Sim, P.Geo. SIM Geological Inc. and Bruce M. Davis, FAusIMM, BD Resource Consulting, Inc.

Mineral resource estimates are made from a 3D block model based on geostatistical applications using commercial mine planning software (MineSight® v11.60-2). The block model has a nominal block size measuring 10 x 10 x 5 m and utilizes data derived from 152 drill holes in the vicinity of the Arctic deposit. The resource estimate was generated using drill hole sample assay results and the interpretation of a geological model which relates to the spatial distribution of copper, lead, zinc, gold and silver. Interpolation characteristics were defined based on the geology, drill hole spacing, and geostatistical analysis of the data. The effects of potentially anomalous high-grade sample data, composited to two metre intervals, are controlled by limiting the distance of influence during block grade interpolation. The grade models have been validated using a combination of visual and statistical methods. The resources were classified according to their proximity to the sample data locations and are reported, as required by NI 43-101, according to the CIM Definition Standards for Mineral Resources and Mineral Reserves. Model blocks estimated by three or more drill holes spaced at a maximum distance of 100 metres are included in the Indicated category. Inferred blocks are within a maximum distance of 150 metres from a drill hole. The estimate of Indicated and Inferred mineral resources is within a limiting pit shell derived using projected technical and economic parameters.

The mineral resource estimate is listed in Table 1-2.

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Table 1-2        Mineral Resource Estimate for the Arctic Deposit

		Average Grade:					Contained metal:
Class	M tonnes	Cu %	Pb%	Zn%	Au g/t	Ag g/t	Cu Mlbs	Pb Mlbs	Zn Mlbs	Au koz	Ag Moz
Indicated	36.0	3.07	0.73	4.23	0.63	47.6	2441	581	3356	728	55
Inferred	3.5	1.71	0.60	2.72	0.36	28.7	131	47	210	40	3

Notes:

(1) Resources stated as contained within a pit shell developed using metal prices of US$3.00/lb Cu, $0.90/lb Pb, $1.00/lb Zn, $1300/oz Au and $18/oz Ag and metallurgical recoveries of 92% Cu, 77% Pb, 88% Zn, 63% Au and 56% Ag and operating costs of $3/t mining and $35/t process and G&A. The average pit slope is 43 degrees.

(2) The base case cut-off grade is 0.5% copper equivalent. CuEq = (Cu%x0.92)+(Zn%x0.290)+(Pb%x0.231)+(Augptx0.398)+(Aggptx0.005).

(3) The Mineral Resource Estimate is reported on a 100% basis without adjustments for metallurgical recoveries.

(4) The Estimate of Mineral Resources is inclusive of Mineral Reserves. 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 into Mineral Reserves.

(5) Inferred resources have a great amount of uncertainty as to whether they can be mined legally or economically. It is reasonably expected that a majority of Inferred resources will be converted to Indicated resources with additional exploration.

(6) Effective date of the Mineral Resource Estimate is April 25, 2017.

1.7 Mining Reserves and Mining Methods

The Arctic Project is designed as a conventional truck-shovel operation assuming 131 t trucks for waste and 91 t trucks for ore, as well as 17 m³ and 12 m³ shovels for waste and ore respectively. The pit design includes three nested phases to balance stripping requirements while satisfying the concentrator requirements.

The design parameters include a ramp width of 28.5 m, in-pit road grades of 8% and out-pit road grades of 10%, bench height of 5 m, targeted mining width between 70 and 100 m, berm interval of 15 m, variable slope angles by sector and a minimum mining width of 30 m.

The smoothed final pit design contains approximately 43 Mt of ore and 296 Mt of waste for a resulting stripping ratio of 6.9:1. Within the 43 Mt of ore, the average grades are 2.32% Cu, 3.24% Zn, 0.57 % Pb, 0.49 g/t Au and 36.0 g/t Ag.

The Mineral Reserve estimates are shown in Table 1-3.

Table 1-3        Mineral Reserve Estimates for the Arctic Deposit

	Tonnage	Grades
Class	t x 1000	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)
Proven Mineral Reserves	-	-	-	-	-	-
Probable Mineral Reserves	43,038	2.32	3.24	0.57	0.49	36.0
Proven & Probable Mineral Reserves	43,038	2.32	3.24	0.57	0.49	36.0
Waste within Designed Pit	296,444
Total Tonnage within Designed Pit	339,482

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Notes:

(1) Mineral Reserves are estimated assuming open pit mining methods and include a combination of planned and contact dilution.

(2) Mineral Reserves are based on prices of $2.90/lb Cu, $0.90/lb Pb, $1.10/lb Zn, $1250/oz Au and $18/oz Ag Fixed process recoveries of 90.0% Cu, 89.9% Pb, 91.7% Zn, 61.1% Au and 49.7% Ag

(3) Mining costs: $3.00/t incremented at $0.02/t/15 m and $0.015/t/15 m below and above 710 m elevation respectively.

(4) Processing costs: $36.55/t. Include process cost: $19.86/t, G&A: $8.92/t, sustaining capital: $4.11/t closure cost: $1.00/t, and road toll: $2.66/t.

(5) Treatment costs of $70/t Cu concentrate, $180/t Pb concentrate and $300/t Zn concentrate. Refining costs of $0.07/lb Cu, $10/oz Au, $0.60/oz Ag. Transport cost of $149.96/t concentrate.

(6) Fixed royalty percentage of 1%.

(7) The Qualified Person for the Mineral Reserves is Antonio Peralta Romero P.Eng., an Amec Foster Wheeler employee who visited the project site in July 25, 2017 as part of the data verification process.

(8) The effective date of mineral reserves estimate is October 10, 2017.

The scheduling constraints set the maximum mining capacity at 32 Mt/year and the maximum process capacity at 10 kt/day. The production schedule results in a life of mine (LOM) of 12 years. The mine will require two years of pre-production before the start of operations in the processing plant.

1.8 Recovery Methods

The 10,000 t/d process plant design is conventional for the industry, will operate two 12 hour 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 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 talc pre-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 a tailings management facility (TMF). Copper, lead, and zinc concentrates will be thickened and pressure-filtered before being transported by truck to a port and shipped to smelters.

The LOM average mill feed is expected to contain 2.32% Cu, 3.24% Zn, 0.57% Pb, 0.49 g/t Au, and 35.98 g/t Ag. Based on the mine plan developed for the PFS and metallurgical testwork results, the life—of-mine (LOM) average metal recoveries and concentrate grades will be:

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· Copper concentrate:

o recovery:              90.0% copper; 11.8% gold; 35.0% silver

o copper grade:       30.3%

· Lead concentrate:

o recovery:               80.0% lead; 61.1% gold; 49.7% silver

o lead grade:            55.0%

· Zinc concentrate:

o recovery:               91.7% zinc

o zinc grade:            59.2%.

The average annual dry concentrate production is estimated as follows:

· Copper concentrate:      246,723 t/a

· Lead concentrate:          29,493 t/a

· Zinc concentrate:           180,219 t/a.

1.9 Project Infrastructure

The Arctic project site is a remote, greenfields site that requires construction of its own infrastructure to support the mining operation.

The Project site will be accessed through a combination of State of Alaska owned highways (existing), an Alaska Industrial Development and Export Authority (AIDEA) owned private road (proposed) and Trilogy owned access roads (proposed). The Ambler Mining District Industrial Access Project (AMDIAP) road is proposed by AIDEA to connect the Ambler mining district to the Dalton Highway. The AMDIAP road is being permitted as a private road with restricted access for industrial use. To connect the Arctic Project site and the existing exploration camp to the proposed AMDIAP road a 30.7 km access road (the Arctic access road) will need to be built.

The State of Alaska owned public Dahl Creek Airport will require upgrades to support the planned regular transportation of crews to and from Fairbanks. Power generation will be by six Liquefied Natural Gas (LNG) generators, producing a supply voltage of 4.16 kV. The total connected load will be 17.5 MW with an average power draw of 12.6 MW. Liquid natural gas (LNG) will be supplied via existing fuel supply networks near Port Mackenzie, Alaska.

The Project will require three different self-contained camps, equipped with their own power and heat generation capabilities, water treatment plant, sewage treatment plant, and garbage incinerator. The existing exploration camp will be used to start the construction of the Arctic access road. A construction camp will be constructed at the intersection of the AMDIAP road and Arctic access road, and will be decommissioned once construction is complete. The permanent camp will be constructed along the Arctic access road, closer to the planned processing facility. The permanent camp will be constructed ahead of operations to support the peak accommodation requirements during construction.

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Infrastructure that will be required for the mining and processing operations will include:

· Open pit mine

· Stockpiles and waste rock facilities

· Gatehouse

· Truck workshop, truck wash, mine offices, mine dry facility and warehouse

· Administration building

· Mill dry facility

· Plant workshop and warehouse

· Primary crushing building

· Fine ore stockpile building

· Process plant and laboratory

· Concentrate loadout building

· Reagent storage and handling building

· Raw water supply building.

· Tailings management facility

· Diversion and collection channels, culverts, and containment structures

· Waste rock collection pond

· Water treatment plants.

On-site communications comprise of inter-connected mobile and fixed systems, including landline telephone network, radios and internet.

Compressed air will be supplied by screw compressors and a duty plant air receiver. Fire protection will be supported by a firewater distribution network and standpipe systems, water mist systems, sprinkler systems, and portable fire extinguishers. Gas detection will be provided to detect dangerous levels of LNG gas within the generator room.

A large waste rock dump (WRD) will be developed north of the Arctic pit in the upper part of the Arctic Valley. The waste rock storage facility will be designed to store both waste rock and tailings in adjacent footprints. . The total volume of waste rock is expected to be 145.6 Mm³ (296 Mt), however there is potential for expanded volume in the waste if placement density is less than 2.0 t/m³. The dump will have a final height of 245 m to an elevation of 890 masl and is planned to be constructed in 20 m lifts with intermediary bench widths at 23.5 m on average at the dump face, to achieve an overall slope of 2.7H:1V. Most of the waste rock is anticipated to be potentially acid-generating (PAG) and there will be no separation of waste based on acid generation potential. Rather, seepage from the WRD will be collected and treated.

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There will also be two small overburden stockpiles to store the stripped topsoil and overburden from the TMF footprint. The topsoil stockpile will be placed in between the haul roads and will store up to 225,000 m³ of material while the overburden stockpile will be located below the lower haul road between the pit and the mill site with storage capacity up to 650,000 m³.

The tailings management facility (TMF) will be located at the headwaters of the Sub-Arctic Creek, in the upper-most portion of the creek valley. The 58.6 ha footprint of the TMF will be fully lined with an impermeable liner (HDPE). Tailings containment will be provided by an engineered dam that will be buttressed by the WRD constructed immediately downstream of the TMF and the natural topography on the valley sides. A starter dam will be constructed to elevation 830 m. Three subsequent raises will bring the final dam crest elevation to 890 m, which is the same as the final elevation of the waste rock dump. The TMF is designed to store approximately 27.3 Mm³ (38.7 Mt) of tailings plus 3.0 Mm³ of water produced over the 12 year mine life as well as the PMF and still provide 2m of freeboard.

The tailings delivery system pipeline will transport slurried tailings from processing plant to the TMF. The delivery system will be sized initially on the basis of a 10 kt/d operation. This pipeline will transport 1,050 m³/h of tailings to the TMF. The return water delivery system for recycle water from the TMF has been sized on the basis of 770 m³/h of water being pumped from the TMF to the process water pond, for the 10 kt/d operation.

The proposed mine development is located in valley of Sub-Arctic Creek, a tributary to the Shungnak River. A surface water management system will be constructed to segregate contact and non-contact water. Non-contact water will be diverted around mine infrastructure to Sub-Arctic Creek. Contact water will be conveyed to treatment facilities prior to discharge to the receiving environment.

1.10 Market Studies

Trilogy Metals provided Ausenco with the metal price projections for use in the Pre-Feasibility Study on which the Technical Report is based. Trilogy Metals established the pricing using a combination of two year trailing actual metal prices, and market research and bank analyst forward price projections, prepared in early 2018 by Jim Vice of StoneHouse Consulting Inc.

The long-term consensus metal price assumptions to be used in the Pre-Feasibility Study are:

· Copper: $3.00/lb

· Lead: $1.00/lb

· Zinc: $1.10/lb

· Gold: $1,300/oz

· Silver: $18.00/oz

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Smelter terms were applied for the delivery of copper, zinc and lead concentrate. It was assumed that delivery of all concentrates would be to an East Asian smelter at currently available freight rates. These terms are considered to be in line with current market conditions. Total transport costs for the concentrate are estimated at $270.37/dmt.

1.11 Environmental, Permitting, Social and Closure Considerations

1.11.1 Environmental Considerations

The Arctic Project area includes the Ambler lowlands and Subarctic Creek within the Shungnak River drainage. To date, a moderate amount of baseline environmental data collection has occurred in the area including surface water quality sampling, surface hydrology monitoring, wetlands mapping, stream flow monitoring, aquatic life surveys, avian and mammal habitat surveys, cultural resource surveys, hydrogeology studies, meteorological monitoring, and acid base accounting studies.

1.11.2 Permitting Considerations

Trilogy performs mineral exploration at the Arctic Deposit under State of Alaska and Northwest Arctic Borough (NWAB) permits. Trilogy is presently operating under a State of Alaska Miscellaneous Land Use Permit (APMA permit) that expires at 2017 year-end, and an application to renew was submitted by Trilogy to the Alaska Department of Natural Resources (ADNR) in February 2018.

Mine development permitting will be largely driven by the underlying land ownership; regulatory authorities vary depending on land ownership. The Arctic Project area includes patented mining claims (private land under separate ownership by Trilogy and NANA), State of Alaska land, and NANA land (private land). The open pit would situate mostly on patented land while the mill, tailings and waste rock facilities would be largely on State land. Other facilities, such as the camps, would be on NANA land. Federal land would likely be part of any access road between the Dalton Highway and the Arctic Project area. Permits associated with such an access road are being investigated in a separate action by the State of Alaska.

Because the Arctic Project is situated to a large extent on State land, it will likely be necessary to obtain a Plan of Operation Approval (which includes the Reclamation Plan) from the ADNR. The Project will also require certificates to construct and then operate a dam(s) (tailings and water storage) from the ADNR (Dam Safety Unit) as well as water use authorizations, an upland mining lease and a mill site lease, as well as several minor permits including those that authorize access to construction material sites from ADNR.

The Alaska Department of Environmental Conservation (ADEC) would authorize waste management under an integrated waste management permit, air emissions during construction and then operations under an air permit, and require an 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 review the USACE Section 404 permit to certify that it complies with Section 401 of the CWA.

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The Alaska Department of Fish and Game (ADFG) would have to authorize any culverts or bridges that are required to cross fish-bearing streams or other impacts to fish-bearing streams that result in the loss of fish habitat.

US Army Corps of Engineers (USACE) would require a CWA Section 404 permit for dredging and filling activities in Waters of the United States (WOTUS) including jurisdictional wetlands. The USACE Section 404 permitting action would require the USACE to comply with the Natural Environmental Policy Act (NEPA) and, for a project of this magnitude, the development of an Environmental Impact Statement (EIS). The USACE would likely be the lead federal agency for the NEPA process. As part of the Section 404 permitting process, the Arctic Project will have to meet USACE wetlands guidelines to avoid, minimize and mitigate impacts to wetlands.

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 to the US Army Corp of Engineers (ACOE). The timeline includes the development and publication of a draft and final EIS and ends with a Record of Decision, and 404-permit issuance. 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 require between 1.5 to three years to complete, and could potentially take longer.

1.11.3 Social Considerations

The Arctic Project is located approximately 40 km northeast of the native villages of Shungnak and Kobuk, and 64 km east-northeast of the native village of Ambler. The population in these villages range from 156 in Kobuk (2016 Census) to 262 in Shungnak (2016 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. Trilogy Metals is working directly with the villages to employ residents in the ongoing exploration program as geotechnicians, drill helpers, and environmental technicians. Trilogy Metals and NANA have established a Workforce Development Committee to assist with developing a local workforce. In addition, Trilogy Metals has existing contracts with native-affiliated companies (such as NANA Management Services and WHPacific Inc.) that are providing camp catering and environmental services for the Project, respectively.

Local community concerns will also be formally recognized during the development of the project EIS. Early in the EIS process, the lead federal permitting agency will hold scoping meetings in rural villages to hear and record the concerns of the local communities so that the more significant of these concerns can be addressed during the development of the EIS. In addition, the lead federal agency would have government-to-government consultations with the Tribal Councils in each of the villages, as part of the EIS process, to discuss the project and hear Council concerns.

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1.11.4 Closure Planning

Mine reclamation and closure are largely driven by State regulations 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 from acid rock drainage or metal leaching on the site. A detailed reclamation plan will be submitted to the State agencies for review and approval in the future, during the formal mine permitting process.

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 land, 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 Plan. In addition, private land owners must formally concur with the portion of the reclamation plan for their lands so that it is compatible with their intended post-mining land use.

The estimate cost of closure is based on unit rates used by SRK. Long-term water treatment and maintenance of certain water management facilities were calculated separately, and a net present value (NPV) is provided for the first 200 years, at a discount rate of 4.3%.

Reclamation costs have been estimated to be $65.3 million for this PFS, in 2017 undiscounted US dollars. Annual costs associated with long-term operations of the water treatment plant are estimated to be about $1.27 million for the first five years and $0.96 million thereafter.

1.12 Capital Costs

The capital cost estimate uses US dollars as the base currency. The total estimated initial capital cost for the design, construction, installation and commissioning of the Arctic Project is estimated to be $779.6 million. A summary of the estimated capital cost is shown in Table 1-4.

Table 1-4        Initial Capital Costs

Cost Type	Description	US$M
Direct	Mine	281.1
	Crushing	18.3
	Process	113.8
	Tailings	30.3
	On-Site Infrastructure	84.5

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Cost Type	Description	US$M
	Off-Site Infrastructure	15.6
	Direct Subtotal	543.8
Indirect	Indirects	121.9
	Contingency	92.0
	Owners Costs	21.9
	Indirect Total	235.8
Project Total		779.6

The total sustaining capital cost estimate is $65.9 million for the 12 year LOM which includes equipment, tailings and other items. Closure costs were estimated to be $65.3 million. These costs are summarized in Table 1-5.

Table 1-5        Sustaining Capital and Closure Costs

	Sustaining Capital (US$M)
G&A	0.9
Tailings	19.9
Mining	45.1
Total Sustaining Capital	65.9
	Closure Cost (US$M)
Closure Costs	65.3

1.13 Operating Costs

The operating cost estimates use US dollars as the base currency. An average operating cost was estimated for the Arctic Project based on the proposed mining schedule. These costs included, mining, processing, G&A, surface services, and road toll costs. The average LOM operating cost for the Arctic Project is estimated to be $46.81/ t milled. The breakdown of costs in Table 1-6 is estimated based on the average LOM mill feed rate.

Table 1-6        Sustaining Capital and Closure Costs

Description	LOM Average Unit Operating Cost ($/ t milled)	Percentage of Total Annual Operating Costs
Mining*	20.47	44%
Processing	15.09	32%
G&A	5.60	12%
Surface Services	0.95	2%
Road Toll	4.70	10%
Total Operating Cost	46.81	100%

*Excludes pre-production costs

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1.14 Economic Analysis

An economic analysis was undertaken to determine the internal rate of return (IRR), net present value (NPV) and payback on initial investment of the Arctic Project. The project consists of a three year pre-production construction period, followed by 12 years of production.

The results of this economic analysis, represents forward looking information. The results depend on the inputs that are subject to a number of known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented in this section. Information that is forward looking includes mineral reserve estimates, commodity prices, the proposed mine production plan, construction schedule, projected recovery rates, proposed capital and operating cost estimates, closure cost estimates, toll road cost estimates, and assumptions on geotechnical, environmental, permitting, royalties, and hydrogeological information.

Ausenco developed a pre-tax cash flow model for the Arctic Project and the NPV and IRR were calculated at the beginning of the construction period in Year -3.

The pre-tax financial model incorporated the production schedule and smelter term assumptions to produce annual recovered payable metal, or gross revenue, in each concentrate stream by year. Off-site costs, including the applicable refining and treatment costs, penalties, concentrate transportation charges, marketing and representation fees, and royalties were then deducted from gross revenue to determine the NSR. The operating cash flow was then produced by deducting annual mining, processing, G&A, surface services, and road toll charges from the NSR. Initial and sustaining capital was deducted from the operating cash flow in the years they occur, to determine the net cash flow before taxes. Initial capital cost includes all estimated expenditures in the construction period, from Year -3 to Year -1 inclusive. First production occurs at the beginning of Year 1. Sustaining capital expenditure includes all capital expenditures purchased after first production, including mine closure and rehabilitation. The model includes an allocation of a 1% NSR attributable to NANA.

The pre-tax financial results are:

· 38.0% IRR

· $1,935.2 million NPV at an 8% discount rate

· 1.9 year payback period, on the initial capital costs of $779.6 million

The following tax regimes were incorporated in the post-tax analysis: US Federal Income Tax, Alaska State Income 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, including the US Federal enactment of the Tax Cuts & Jobs Act (TCJA) on December 22, 2017. At the base case metal prices used for this study, the total estimated taxes payable on the Arctic Project profits are $1,162.2 million over the 12-year mine life.

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The post-tax financial results are:

· 33.4% IRR

· $1,412.7 million NPV at an 8% discount rate

· 2.0 year payback period, on the initial capital costs of $779.6 million

1.15 Sensitivity Analysis

Ausenco investigated the sensitivity of the Project’s pre-tax NPV, and IRR to several project variables, including metal prices (copper, lead, zinc, gold, silver), capital costs, and operating costs (onsite and offsite). The metal grade is not presented in these findings because the impacts of changes in the metal grade mirror the impact of changes in metal price.

The Project’s pre-tax NPV at an 8% discount rate is most sensitive to changes in copper price, followed by zinc price, off-site operating costs, on-site operating costs, capital costs, silver price, gold price, and lead price.

The Project’s pre-tax IRR is most sensitive to changes in copper price and capital cost, followed by zinc price and off site operating costs, and in then decreasing order, on-site operating costs, silver price, gold price, and lead price.

1.16 Interpretations and Conclusions

Under the assumptions presented in this Report, the Project shows positive economics.

The financial analysis excludes consideration of the NANA Agreement, whereby NANA has the right, following a construction decision, to elect to purchase a 16% to 25% direct interest in the Arctic Project or, alternatively, to receive a 15% Net Proceeds Royalty.

The financial analysis excludes consideration of the South32 Option Agreement, whereby South32 has the right to form a 50/50 Joint Venture with Trilogy Metals over Trilogy Metal’s Alaskan interests, including the Arctic Project.

The cost assumptions for the AMDIAP road are estimates provided by Trilogy Metals. There is a risk to the capital and operating cost estimates, the financial analysis, and the Mineral Reserves if the toll road is not built in the time frame required for the Arctic Project, or if the toll charges are significantly different from what was assumed.

1.17 Recommendations

A single-phase work program is recommended, which will include: geotechnical investigations and studies; optimization of the plant and related service facilities and evaluation of the power supply; examination of water management, water treatment, WRD and TSF designs; baseline studies and environmental permitting activities; and additional metallurgical testwork. The budget for this work is estimated at about $3.3 million.

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2.0 Introduction

2.1 Terms of Reference

Trilogy Metals Inc. (Trilogy Metals or Trilogy) commissioned Ausenco Engineering Canada Inc. (Ausenco) to compile a Technical Report (the Report) on the results of a Pre-Feasibility Study on the Arctic Deposit, part of the Arctic Project (the Project) in the Ambler District of Northwest Alaska (Figure 2-1).

Figure 2-1        Property Location Map (Tetra Tech, 2013)

The Report supports disclosure by Trilogy Metals in the news release dated 20 February 2018, entitled “Trilogy Metals Announces Pre-Feasibility Study Results and Reserves for the Arctic Project, Alaska”.

The firms and consultants who are providing Qualified Persons (QPs) responsible for the content of this Report, which is based on the Pre-Feasibility Study completed in 2018 (the 2018 PFS) and supporting documents prepared for the 2018 PFS, are, in alphabetical order, Amec Foster Wheeler Americas Ltd. (Amec Foster Wheeler); BD Resource Consulting, Inc., (BDRC); SRK Consulting (Canada) Inc. (SRK), and SIM Geological Inc. (SIM).

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The Report presents Mineral Resource and Mineral Reserve estimates for the Project, and an economic assessment based on open pit mining operations and a conventional processing circuit that would produce copper, zinc and lead concentrates.

2.2 Units of Measurement

All units of measurement in this Report are metric, unless otherwise stated.

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

2.3 Qualified Persons

The following serve as the qualified persons for this Technical Report as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, and in compliance with Form 43-101F1:

· Mr. Paul Staples, P.Eng., Vice President and Global Practice Lead, Minerals and Metals, Ausenco Engineering Canada Inc.

· Mr. Justin Hannon, P.Eng., Senior Mining Engineer and Financial Analyst, Ausenco Engineering Canada Inc.

· Mr. Antonio Peralta Romero, PhD, P.Eng, Principal Mining Engineer, Amec Foster Wheeler Americas Ltd.

· John J. DiMarchi, CPG, Principal Core Geoscience Inc.

· Mr. Bruce Davis, FAusIMM, BD Resource Consulting, Inc.

· Mr. Jeffrey B. Austin, P. Eng, International Metallurgical & Environmental Inc.

· Mr. Robert Sim, P.Geo, SIM Geological Inc.

· Mr. Calvin Boese, P.Eng, M.Sc., Senior Consultant, Geotechnical Engineering, SRK Consulting (Canada) Inc.

· Bruce Murphy, P.Eng., M.Sc., Principal Consultant, Rock Mechanics, SRK Consulting (Canada) Inc.

· Tom Sharp, PhD, P.Eng, Principal Consultant, Water Management and Treatment Engineering, SRK Consulting (Canada) Inc.

2.4 Site Visit

Antonio Peralta Romero visited the Arctic Project site July 25, 2017. The following activities occurred during this visit:

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· Inspection of access to property and surface topography where open pit and waste dumps are to be located, as well as available space for other mine facilities.

· Inspection of drill core relevant to the rock formations that would support the pit wall.

· Observed structural features in bedrock exposure that could affect pit slope stability

Paul Staples visited the Arctic Project site July 25, 2017. The following activities occurred during this visit:

· Inspection of access to property and surface topography where open pit and waste dumps and other mine facilities are to be located.

Calvin Boese visited the Arctic Project site July 24 – 25, 2017. The following activities occurred during this site visit:

· Inspection of property and surface topography where waste rock dump, tailings management facility, are to be located, as well as available space for other mine facilities.

Bruce Davis conducted a site visit to the Project on July 26 - 27, 2011, on September 25, 2012, and again on August 10-12, 2015.

· The site visit included a review of: drilling procedures, site facilities, historic and recent drill core, logging procedures, data capture, and sample handling. During the 2015 Arctic site visit, Mr. Davis undertook a helicopter traverse along proposed access corridors and potential site layouts.

2.5 Effective Dates

The Report has a number of effective dates as follows:

· The effective date of the Mineral Resource Estimate is April 25, 2017

· The effective date of the Mineral Reserve estimate is October 10, 2017

· The effective date of the financial analysis is February 20, 2018

· Release date of the Report is April 6, 2018

The overall effective date of the Report is taken to be the date of the financial analysis and Mineral Reserve estimate and is February 20, 2018.

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2.6 Information Sources

Reports and documents listed in Section 2.7, Section 3.0, and Section 27.0 were used to support the preparation of the Report. Additional information was sought from Trilogy Metals personnel where required.

2.7 Previous Technical Reports

Technical reports filed by Trilogy Metals, and its predecessor companies, NovaCopper Inc. (NovaCopper), and NovaGold Resources Inc. (NovaGold) include:

· Davis, B., Sim, R., and Austin, J., 2017: NI 43-101 Technical Report on the Arctic Project, Northwest Alaska, USA: report prepared by BD Resource Consulting, Inc., SIM Geological Inc., and International Metallurgical & Environmental Inc. for Trilogy Metals Inc., effective date April 25, 2017.

· Wilkins, G., Stoyko, H.W., Ghaffari, H., DiMarchi, J., Huang, J., Silva, M., O’Brien, M.F., Chin, M., and Hafez, S.A., 2013: Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska: report prepared by Tetra Tech for Nova Copper, effective date September 12, 2013.

· Rigby, N., White, R., Volk, J., Braun, T., and Olin, E.J., 2012: NI 43-101 Preliminary Economic Assessment Ambler Project Kobuk, AK: report prepared by SRK Consulting (US) Inc. for Nova Copper, effective date February 1, 2012.

· Rigby, N., and White, R., 2011: NI 43-101 Preliminary Economic Assessment Ambler Project Kobuk, AK: report prepared by SRK Consulting (US) Inc. for Nova Copper, effective date May 9, 2011.

· Rigby, N., and White, R., 2008: NI 43-101 Technical Report on Resources Ambler Project Arctic Deposit, Alaska: report prepared by SRK Consulting (US) Inc. for Nova Copper, effective date January 31, 2008.

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3.0 Reliance on Other Experts

3.1 Introduction

The QPs have relied upon the following expert reports, which provided information regarding mineral rights, surface rights, property agreements, royalties, legal assumptions and taxation and this Report.

3.2 Mineral Tenure, Surface Rights, Royalties, Property Agreements

The QPs have not independently reviewed ownership of the Project area and any underlying property agreements, mineral tenure, surface rights, or royalties. The QPs have fully relied upon, and disclaim responsibility for, information derived from Trilogy Metals and legal experts retained by Trilogy Metals for this information through the following documents

Kennecott Exploration Company, Kennecott Arctic Company, Alaska Gold Company and NovaGold Resources Inc., 2010:  Net Smelter Returns Royalty Agreement dated effective January 7, 2010:  51 p.

NovaCopper US Inc. and NANA Regional Corporation Inc., 2011:  Exploration Agreement and Option to Lease, dated effective October 19, 2011:  144 p.

NovaCopper US Inc. and NANA Regional Corporation Inc., 2012: Amending Agreement, dated effective May 10, 2012:  7 p.

This information is used in Section 4.0 of the Report. The information is also used in support of the Mineral Resource estimate in Section 14.0, the Mineral Reserve estimate in Section 15.0, and the financial analysis in Section 22.0.

3.3 Legal Considerations

The QPs have not independently reviewed ownership of the Project area and any underlying property agreements, mineral tenure, surface rights, or royalties. The QPs have fully relied upon, and disclaim responsibility for, information derived from Trilogy Metals and legal experts retained by Trilogy Metals for this information through the following document:

Reeves, J.N., 2018: Arctic Project: legal opinion prepared by Holmes Weddell & Barcott for Trilogy Metals Inc., 4 April 2018, 58 p.

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This information is used in Section 4.0 of the Report. The information is also used in support of the Mineral Resource estimate in Section 14.0, the Mineral Reserve estimate in Section 15.0, and the financial analysis in Section 22.0.

3.4 Taxation

The QPs have fully relied upon, and disclaim responsibility for, information supplied by Trilogy Metals staff and experts retained by Trilogy Metals for information related to taxation as applied to the financial model as follows:

Ernst & Young LLP, 2018: Provisions of income tax and mineral tax portions of economic analysis for the Pre-feasibility technical study report on Trilogy’s Arctic Project, 19 February, 2018.

This information is used in the financial analysis in Section 22.0 of the Report.

3.5 Marketing and Contracts

The QPs have fully relied upon, and disclaim responsibility for, information supplied by Trilogy Metals staff and experts retained by Trilogy Metals for information related to market assumptions as applied to the financial model as follows:

StoneHouse Consulting Inc.: Market Assumptions for Trilogy Technical Report: letter prepared by Jim Vice, President of StoneHouse Consulting Inc. for Ausenco, dated March 22, 2018.

This information is used in support of the marketing assumptions in Section 19.0, financial analysis in Section 22.0, and the Mineral Reserve estimate in Section 15.0.

3.6 Metal Prices and Exchange Rates

The QPs have fully relied upon, and disclaim responsibility for, information supplied by Trilogy Metals staff and experts retained by Trilogy Metals for information related to metal price and exchange rate assumptions as applied to the financial model as follows:

StoneHouse Consulting Inc.: Metal Price Assumptions for Trilogy Technical Report: letter prepared by Jim Vice, President of StoneHouse Consulting Inc. for Ausenco, dated March 26, 2018.

This information is used in support of the metal price assumptions in Section 19.0, the financial analysis in Section 22.0, the Mineral Reserve estimate in Section 15.0 and the Mineral Resource estimate in Section 14.0; and the exchange rate assumptions for the capital cost estimates in Section 21.0.

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4.0 Property Description and Location

4.1 Location

The Project is situated in the Ambler mining district of the southern Brooks Range, in the Northwest Arctic Borough (NWAB) of Alaska. The Project 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 about 270 km east of the town of Kotzebue, 37 km northeast of the village of Kobuk, and 260 km west of the Dalton Highway, an all-weather State maintained public road, at geographic coordinates N67.17° latitude and W156.39° longitude (Universal Transverse Mercator (UTM) North American Datum (NAD) 83, Zone 4 coordinates 7453080N, 613110E).

4.2 Ownership

The Project is indirectly 100% owned by Trilogy Metals. The original ownership interest was held by NovaGold. In 2012, NovaGold transferred all copper projects to NovaCopper. NovaCopper subsequently underwent a name change to Trilogy Metals in 2016. Additional information on historical ownership is provided in Section 6.1 of this Report.

4.3 Mineral Tenure

The Project comprises approximately 46,336 ha of State of Alaska mining claims and US Federal patented mining claims in the Kotzebue Recording District. The land tenure consists of 1,388 contiguous claims totaling 114,954.9 acres, including 883 40-acre State claims, 503 160-acre State claims, and 18 Federal patented claims comprising 271.9 acres held in the name of NovaCopper US Inc. (NovaCopper), a wholly owned subsidiary of Trilogy Metals. Claim locations are shown in Figure 4-1 to Figure 4-3 and listed in Appendix A, List of Claims. The Arctic deposit is located near the southern edge of the centre of the claim block Figure 4-4, primarily within the Federal patented claims.

The Federal patented claim corners were located by the US Geological Survey (USGS). In Appendix A, the Federal patented claims are reported using the completed mineral surveys, USMS2245-1 and USMS2245-2. USMS2245-2 covers the Arctic 10 and Arctic 49 Federal patented claims and is included in Appendix A under 50-83-0174. USMS2245-1 covers the remaining 16 Federal patented claims (Arctic 1, 2, 4, 11, 13, 15, 17, 19, 29, 28, 27, 26, 25, 24, 23, and 9), and is included Appendix A under 50-81-0127. Figure 4-4 included the locations of the Federal patented claims. There is no expiration date or labour requirement on the Federal patented claims.

Rent for each State claim is paid annually to the Alaska Department of Natural Resources (ADNR). An Annual Labour Statement must be submitted annually to maintain the State claims in good standing. Legal opinion provided to Trilogy Metals supports that the State mining claims are in “active” status and in good standing with the Alaska Department of Natural Resources.

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Figure 4-1        Upper Kobuk Mineral Projects Lands (Trilogy Metals, 2017)

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Figure 4-2        Arctic Project Mineral Tenure Plan (Trilogy Metals, 2017)

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Figure 4-3        Mineral Tenure Layout Plan (Trilogy Metals, 2018)

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Figure 4-4        Arctic Deposit Location (Trilogy Metals, 2018)

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4.4 Royalties, Agreements and Encumbrances

4.4.1 Kennecott Agreements

A Net Smelter Returns Royalty Agreement (the Kennecott Royalty Agreement) dated effective January 7, 2010 was entered into by and among Kennecott Exploration Company, Kennecott Arctic Company (collectively Kennecott), Alaska Gold Company and NovaGold Resources Inc. A copy of the Kennecott Agreement was recorded in the Kotzebue Recording District on January 8, 2010, as document no. 2010-000013-0.

The Kennecott Royalty Agreement documents a net smelter returns royalty that was reserved to the Kennecott parties in a Purchase and Sale Agreement dated December 18, 2009 whereby Alaska Gold Company and NovaGold Resources Inc. acquired mining properties from Kennecott. The mining properties referenced in the Kennecott Royalty Agreement consist of the Federal patented mining claims and many, but not all, of the State mining claims that are the subject of this Report.

The Kennecott Royalty Agreement provides for the payment of a 1% royalty on net smelter returns from production from the properties to which it applies. The Kennecott Royalty Agreement gives the Grantors (Alaska Gold Company and NovaGold Resources Inc.) an option to purchase the royalty for a payment of $10,000,000. By operation of the enurement clause of the Kennecott Royalty Agreement, it appears that this option follows the ownership of the property and may be exercised by a successor owner, such as NovaCopper US Inc.

4.4.2 NANA Agreement

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 13 corporations are known as the Alaska Native Regional Corporations (ANCSA Corporations). One of these 13 regional corporations is the NANA Regional Corporation, Inc. (NANA). ANCSA Lands controlled by NANA bound the southern border of the Project claim block (refer to Figure 4-1).

On October 19, 2011, Trilogy Metals and NANA 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 Trilogy Metals’ 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 Trilogy Metals 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 use temporary access roads, camps, airstrips and other incidental works.

The NANA Agreement has a term of 20 years, with an option in favour of Trilogy Metals 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 Trilogy Metals does not meet certain expenditure requirements on NANA’s lands.

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If, following receipt of a feasibility study and the release for public comment of a related draft EIS, Trilogy Metals decides to proceed with construction of a mine on the lands subject to the NANA Agreement, Trilogy Metals 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 Trilogy Metals from such project. The cost to exercise such back-in-right is equal to the percentage interest in the Project multiplied by the difference between (i) all costs incurred by Trilogy Metals 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 Trilogy Metals 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 electing to participate between 16% to 25%, and Trilogy Metals 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 Trilogy Metals and be entitled to all the benefits of Trilogy Metals 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, Trilogy Metals and NANA will execute a mining lease (the Mining Lease) to allow Trilogy Metals or a joint venture vehicle to construct and operate a mine on the Bornite Lands or ANCSA Lands. The Mining Lease will provide NANA with a 2% net smelter royalty (NSR) as to production from the Bornite Lands and a 2.5% NSR as to production from the ANCSA Lands.

If Trilogy Metals 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 Trilogy Metals which will afford Trilogy Metals access to the Project along routes approved by NANA (the Surface Use Agreement). In consideration for the grant of such surface use rights, Trilogy Metals will grant NANA a 1% NSR 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.

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Figure 4-1 showed the locations of the Bornite and ANCSA Lands that are included in the NANA Agreement. The Bornite Lands are not considered to be part of the Arctic Project, because the mineralization styles identified to date in the Bornite Lands are distinctly different to the mineralization styles in the Ambler claims, and it is expected that any mining operation in the Bornite Lands would be developed as a stand-alone operation using different infrastructure.

4.4.3 South32 Agreement

South32 may become entitled to exercise an option (the Option) to require that Trilogy Metals form a new limited liability company, 50% of the Membership Interest of which would be held by South32, and Trilogy Metals would be obligated to transfer all of its assets to the new entity. This would include all of the property that is the subject of this Report. Hence, by virtue of this South32 Agreement, it could be said that South32 holds a contingent right to become the indirect beneficial owner of a 50% interest in the property. The South32 Agreement also provides that during the term South32 Agreement prior to exercise of the Option and for so long as it remains in effect the rights of Trilogy Metals and NovaCopper to convey, hypothecate or otherwise encumber the property is strictly limited. The South32 Agreement also addresses the coordination of the rights and obligations it creates with those arising under the NANA Agreement.

4.4.4 State Royalty

The owner of a State mining claim or lease will be obligated to pay a production royalty to the State of Alaska in the amount of 3% of net income received from minerals produced from the State mining claims.

This royalty does not apply to patented federal mining claims.

4.5 Surface Rights

Surface use of the private land held as Federal patented claims is limited only by reservations in the patents and by generally-applicable environmental laws.

Surface use of State claims allows the owner of the mining claim to make such use of the surface as is “necessary for prospecting for, extraction of, or basic processing of minerals.”

4.6 Environmental Considerations

Environmental considerations are discussed in Section 20.0.

There may be some environmental liabilities associated with sites explored during the 1950s and 1960s. The exploration camp would require rehabilitation if the Project is closed.

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4.7 Permits

Permitting considerations for the Project are discussed in Section 20.0.

4.8 Social Considerations

Social considerations for the Project are discussed in Section 20.0.

4.9 Comment on Section 4

To the extent known, Trilogy Metals has advised Ausenco that there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the property that have not been discussed in this Report.

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5.0 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Accessibility

5.1.1 Air

Primary access to the Project 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 Project, capable of accommodating charter fixed wing aircraft. These airstrips are located 64 km west at Ambler, 46 km southwest at Shungnak, 37 km southwest at Kobuk, and 34 km southwest at Dahl Creek. There is daily commercial air service from Kotzebue to the village of Kobuk, the closest community to the Project. During the summer months, the Dahl Creek Camp airstrip is suitable for larger aircraft, such as a 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. The airstrip at Bornite is suited to smaller aircraft, which support the Bornite Camp with personnel and supplies. There is also a 450 m airstrip (Arctic airstrip) located at the base of Arctic Ridge that is able to support smaller aircraft.

5.1.2 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 located at Bornite to the Dahl Creek airstrip southwest of the Arctic deposit. An unimproved gravel track connects the Arctic airstrip with the Arctic deposit.

5.1.3 Water

There is no direct water access to the Project. During spring runoff, river access is possible by barge from Kotzebue Sound to Ambler, Shungnak, and Kobuk via the Kobuk River.

5.2 Climate

The climate in the region is typical of a sub-arctic environment. . Weather conditions on the Project 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 Project. During winter months, the Project 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 500 mm for elevations lower than 600 meters above sea level (“MASL”) with the most rainfall occurring from June through September, and the most snowfall occurring from November through January.

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It is expected that any future mining activity will be conducted on a year-round basis. Exploration activities are generally confined to the period from late May to late September.

5.3 Local Resources and Infrastructure

The Project is currently isolated from major public infrastructure. Infrastructure assumptions and the proposed infrastructure layout for the Project are discussed in Section 18.0 of the Report.

The Project is approximately 270 km east of the town of Kotzebue, on the edge of Kotzebue Sound, 37 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 the location of one of the airstrips near the Project. Several other villages are also near the Project, including Shungnak located 46 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 Project. Fairbanks (population 31,036; 2010 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 Bornite Camp and Dahl Creek Camp. The 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. Power is supplied by two Caterpillar diesel generators – one 300kW and one 225 kW. Water was supplied by the permitted artesian well located 250 m from camp; however, a water well was drilled in camp during the 2017 field season that will be permitted by Spring 2018 to provide all potable water for the Bornite Camp.

5.4 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.

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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. Intermittent 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.

Wildlife in the Project 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. There are no anadromous fish species in the upper reaches of the Shungnak and Kogoluktuk Rivers due to natural fish barriers. Other fish species such as trout, sculpin, and grayling are common. The caribou seen on the Project belong to the Western Arctic herd that migrate once a year heading south in late August through October from their summer range north of the Brooks Range. The caribou migrate north in March from their winter range along the Buckland River to the north slope of the Brooks Range, but take a more westerly route and do not cross the Project during that migration.

5.5 Comment on Section 5

In the opinion of the QP:

· The planned infrastructure, availability of staff, power, water, and communications facilities, the design and budget for such facilities, and the methods whereby goods could be transported to the proposed mine, and any planned modifications or supporting studies are reasonably well-established, or the requirements to establish such, are reasonably well understood by Trilogy Metals, and can support the declaration of Mineral Resources and Mineral Reserves

· There is sufficient area within the Project to host an open pit mining operation, including mine and plant infrastructure, and waste rock and tailings management facilities

· It is a reasonable expectation that any additional surface rights that would be required to support Project development and operations can be obtained through appropriate negotiation

· It is expected that any future mining operations will be able to be conducted year-round.

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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).

Bear Creek Mining Company (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 1966 revealed a significant copper anomaly in Arctic Creek roughly 27 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, culminating 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, and the Smucker Deposit located 36 km west of the Arctic Deposit.

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.

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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 (MS2245 totalling 240 acres over the Arctic Deposit – later amended to include another 32 acres; and MS2233 and MS2234 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 Ambler District

6.1 Prior Ownership and Ownership Changes – Arctic Deposit and the Ambler 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 M2245 covering 16 mining claims totaling 240.018 acres. In 1983, US Mineral Survey patent M2245 was amended to include two additional claims totaling 31.91 acres.

With the passage of the Alaska National Interest Lands Conservation Act (ANILCA) in 1980, which expedited native land claims outlined in the ANSCA and State lands claims under the Alaska 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.

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.

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6.2 Previous Exploration and Development Results – Arctic Deposit

6.2.1 Introduction

Kennecott’s ownership of 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 re-evaluation 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 number 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

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	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
	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 Northern 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

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
	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	BCMC (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
Red Nora	Nora	BCMC/GCO-HOMEX	Disseminated chalcopyrite within chlorite altered volcanics in two zones (Sulphide Gulch and Northern 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

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	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
	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)

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Area	Prospects	Company	Mineralization	Mapping	Soil Geochem	Geophysics	Reports
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 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 Trilogy Metals. 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 one exploration hole 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:

· anomaly 98-3 required further drilling;

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· anomalies 98-7 and 98-22 were drilled ready; and

· anomalies 98-8, -9, -14, -35, and -38 required 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 mineralization;

· grade and thickness continuity; and

· 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

A complete and comprehensive discussion of SG determinations captured during both the Kennecott and Trilogy Metals/NovaGold work programs are provided in Section 11.0 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”; and Bonnie Broman completed a master’s thesis for University of Alaska, Fairbanks in 2014 entitled “Metamorphism and Element Redistribution: Investigations of Ag-bearing and associated minerals in the Arctic Volcanogenic Massive Sulfide Deposit, SW Brooks Range, NW 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.

Geotechnical Studies

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.

Hydrological Studies

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-Base Accounting Studies

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 potentially acid generating.

· 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

Kennecott 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 Development Studies

A number of mining and technical studies have been completed over the Project history, as summarized in Table 6-2.

Table 6-2       Mining and Technical Studies

Company	Year	Consultant	Study
Kennecott	1974	internal	Ambler District Evaluation
	1976	internal	Arctic Deposit Order of Magnitude Evaluation
	1978	internal	Arctic Prospect Summary File Report Arctic Deposit
	1981	internal	Evaluation of the Arctic and Ruby Creek Deposits
	1984	internal	Evaluation Update
	1985	internal	Pre-AFD Report
	1990	internal	Re-Evaluation
	1997	internal	Arctic Project Mining Potential
	1999	internal	Interim Report Conceptual Level Economic Evaluations of the Arctic Resource
	1998	SRK	Preliminary Arctic Scoping Study
NovaGold	2011	SRK	Preliminary Economic Assessment
NovaCopper	2012	SRK	Preliminary Economic Assessment
	2013	Tetra Tech	Preliminary Economic Assessment

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7.0 Geological setting and mineralization

7.1 Regional Geology – Southern Brooks Range

The 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 Late Jurassic 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 allochthons of variable metasedimentary and volcanogenic rocks of Paleozoic age.

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Figure 7-1      Geologic Terranes of the Southern Brooks Range (Trilogy Metals, 2017)

 

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 Jurassic 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.

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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).

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 allochthons 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/Ar³⁹cooling ages of hornblende and white mica near Mount Igikpak, and lasted only a few million years (Vogl et al., 2003). 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 (Trilogy Metals, 2017)

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 Trilogy Metals 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. Trilogy Metals 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.

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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 K-feldspar 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 Trilogy Metals 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.

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 Trilogy Metals to define a generally north-northwest-trending depocentre through the central Ambler District. Volcanic debris flow 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 Arctic 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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Figure 7-3 Ambler Sequence Stratigraphy in the Arctic Deposit Area (Trilogy Metals, 2017)

 

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Figure 7-4 Generalized Geology of the Central Ambler District (Trilogy Metals, 2017)

 

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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.

F1 Deformation

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

Figure 7-5 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 Trilogy Metals 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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Figure 7-6 Generalized Geologic Map of the Arctic Deposit (Trilogy Metals, 2017)

 

7.3.1 Lithologies and Lithologic Domain Descriptions

Historically, five lithologic groupings have been used 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 Trilogy Metals 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 30% 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 massive sulphide (MS) and semi-massive sulphide (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 K-feldspar 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 represents both altered and texturally distinct felsic tuffs and volcaniclastic lithologies.

Volcanic Debris FLOW (Dm)

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 locally-derived debris flows or slumps.

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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.

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 lithologies are logged during data capture, Trilogy Metals models the deposit area as two distinct structural plates, Upper Plate and Lower Plate separated by the Warm Springs Fault. The Upper and Lower plates contain similar lithologic domains which are primarily defined by lithogeochemical characteristics, but are also consistent with their respective acid-generating capacities and spatial distribution around the fold axes, and include the following units: the Button Schist (a meta-rhyolite porphyry - MRP), aphanitic meta-rhyolite (AMR), a series of felsic quartz mica schists (QMS), and carbonaceous schists of the Grey Schist unit (GS). An alteration model has been built to adequately characterize the chlorite and talc schists found within the deposit (ChS, ChTS, and TS). The mineralization is modelled as eight distinct zones (Zones 1–8) found both in the Upper and Lower plates and range from massive sulphide to semi-massive sulphide layers (MS and SMS).

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 Deposit 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 2007 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 Trilogy Metals 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 mineral species are associated with both alteration and mineralization and demonstrate local remobilization during metamorphism (Schmandt, 2009). Though little has been done to document their distribution, they do have a significant impact on bulk density measurements.

Additional discussion of the potential impacts of barite is discussed in the SG Section 13.0 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 metallurgical processing as well as for geotechnical pit slope stability. A great deal of effort has gone into modeling the distribution of talc and talc-chlorite units throughout the deposit area; even zones as small as 10cm have been logged and mapped. The majority of the talc zones occur between the upper, stratigraphically up-right zones and the lower, overturned zones. Significant metallurgical test work has demonstrated that a talc pre-float eliminates talc from interfering with subsequent extraction and concentration of the base and precious metals (see Section 13.0 for further details).

SRK has completed detailed studies on the impact of talc on the pit design; see discussions in Section 9.7.1.

7.4 Arctic Deposit Mineralization

Mineralization occurs as stratiform SMS to MS beds within primarily graphitic schists and fine-grained quartz mica schists. The sulphide beds average 4 m in thickness but vary from less than 1 m up to as much as 32 m in thickness. The sulphide mineralization occurs within eight 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. There are five zones of MS and SMS that occur at specific pseudo-stratigraphic levels which make up the bulk of the mineral resources. The other three zones also occur at specific pseudo-stratigraphic levels, but are too discontinuous to confidently model as resources.

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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.

Mineralization is predominately coarse-grained sulphides consisting mainly of chalcopyrite, sphalerite, galena, tetrahedrite-tennantite, pyrite, arsenopyrite, and pyrrhotite. Trace amounts of electrum are also present. Gangue minerals associated with the mineralized horizons include quartz, barite, white mica, chlorite, stilpnomelane, 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.

Figure 7-7 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-complexed 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 bimodal felsic dominant volcanism related to incipient rifting.

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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 volcanic rocks including pillow basalts and 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.

7.6 Deposits and Prospects

In addition to the Arctic Deposit, numerous other VMS-like occurrences are present on the Trilogy Metals land package. The most notable of these occurrences are the Dead Creek (also known as Shungnak), Sunshine, Cliff, Horse, and the Snow prospects to the west of the Arctic Deposit and the Red, Nora, Tom-Tom and BT prospects to the east. Figure 7-8 shows the Trilogy Metals land package and the prospect locations. Figure 7-8 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 Lead-FX., and 3) carbonate-hosted deposits and prospects in the Bornite Carbonate Sequence controlled by Trilogy Metals/NANA.

Figure 7-8 Major Prospects of the Ambler Mining District (Trilogy Metals, 2017)

 

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8.0 Deposit Types

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 sulphide 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 analogue 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 K-feldspar 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 the Arctic Deposit, sulphides occur as disseminated (<30%), 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-tennantite. 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.

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9.0 Exploration

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

Table 9-1 summarizes the exploration work conducted by NovaGold and Trilogy Metals 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.

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Table 9-1 Summary of Trilogy/NovaGold Exploration Activities Targeting VMS-style Mineralization in the Ambler Sequence Stratigraphy and the Arctic Deposit

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
-	2015, 2016	SRK	Geotechnical Structural Mapping
-	2016	-	Arctic Deposit surface geology
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
LiDAR Survey	2015, 2016	-	LiDAR over Arctic Deposit
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
Geotechnical and Hydrology	2015, 2016	SRK	Arctic PFS Slope Design
ML/ARD	2015, 2016, 2017	SRK	Static Kinetic Tests and ABA Update - ongoing
Metallurgy	2015, 2016, 2017	SGS, ALS	Cu-Pb Separation Test Work; Flotation and Variability Test Work
Project Evaluation
Resource Estimation	2008	SRK	Resource estimation
PEA	2011	SRK	PEA - Underground
	2012	Tetra Tech	PEA – Open Pit

Note: SWIR = short wave infrared; ML = metal leaching; BGC = BGC Engineering Inc.; SRK = SRK Consulting;
SGS = SGS Canada; ALS = ALS Metallurgy; PEA = preliminary economic assessment

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9.1 Grids and Surveys

Survey and data capture during Kennecott’s tenure on the Property was in the UTM coordinates system Zone 4, NAD27 datum. In 2010, NovaGold converted all historical geology and topographic data for the Arctic Deposit into the NAD83 datum for consistency. At that time NovaGold contracted WH Pacific, Inc. (WHPacific) to re-establish project-wide survey control and benchmarks for the Arctic Deposit. Current Mineral Resource and geologic models use topography completed in 2010 by PhotoSat Inc. The resolution of the satellite imagery used was at 0.5 m and a 1 m contour map and digital elevation model (DEM) were generated.

Trilogy Metals retained WHPacific (and sub-consultant Quantum Spatial, Inc.) to conduct an aerial LiDAR survey over the Upper Kobuk Mineral Projects area during 2015. Due to scheduling difficulties and poor weather conditions only 70% of the survey was completed in 2015. The remaining 30% of the aerial survey, as well as the final post-processing work, was completed between June and October 2016.

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 (Trilogy Metals, 2017)

 

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)

· Andy West (2011, 2012, and 2016).

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 (Trilogy Metals, 2017)

 

SRK was contracted in 2015 to create a structural geology model primarily based on brittle structures of the Arctic deposit for pit design and mine scheduling. The majority of the structural mapping took place along the north-south trending Arctic Ridge, and along the northwest trending ridge above the cirque to the south of the deposit, both of which provides the greatest exposure.

Geologic and structural mapping were completed by Trilogy Metals geologists in the Arctic area during the 2016 field season. The objectives of the mapping project were threefold; 1) to ground-truth the northeast and north-south trending fault structures identified by SRK in 2015 and to otherwise support SRK’s 2016 geotechnical mapping efforts, 2) field check the outcrops mapped in 2006 and 2008 recorded in the current GIS database, and 3) determine the nature of the Warm Springs Fault by mapping in the immediate hangingwall of this apparent structural feature. The first objective was successfully accomplished and the pending SRK geotechnical structural model is robust. The two other objectives were partly met during the short field season and the geologic knowledge of the Arctic area was advanced due to the work. All the surface work completed during the field season contributed to the updated Arctic surface geology map Figure 9-3.

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Figure 9-3 2016 Updated Arctic Surface Geology Map (Trilogy Metals, 2017)

 

9.3 Geochemistry

During NovaGold and Trilogy Metals collective tenure in the Ambler District, significant soil and silt geochemical sampling was used 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 Minerals in Fairbanks (a division of ALS Global, formerly ALS Chemex) or Alaska Assay Labs in Fairbanks for sample preparation. The samples were dried and sieved to 80 mesh and forwarded to ALS Minerals 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, copper, chromium, iron, potassium, magnesium, manganese, molybdenum, sodium, nickel, phosphorus, lead, sulphur, antimony, strontium, titanium, vanadium, tungsten, and zinc).

Figure 9-4 and Figure 9-5 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 9-4 Copper Distribution in Silt and Soil Samples in the Dead Creek Area

 

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Figure 9-5 Zinc Distribution in Silt and Soil Samples in the Dead Creek Deposit Area

 

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9.4 Geophysics

A number of different geophysical survey methods have been used 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 9-6 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 drill 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 9-6 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 using 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.

Trilogy Metals has 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 Sulphide 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 Trilogy Metals geologist, completed a Master of Science thesis for the University of Alaska-Fairbanks, 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

Trilogy Metals 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 and Hydrogeological Assessments

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.

In 2016, Trilogy Metals retained SRK Consulting to provide professional engineering services to support pre-feasibility level engineering studies at the Arctic project – this work was a continuation of the work package that was initiated in 2015. The work package was divided into four phases. Phase 1 included LiDAR processing and Quality Assurance/Quality Control (“QA/QC”) to ensure data is of suitable quality to support future engineering planning and design. Phase 2 included a structural desktop study and field investigation plan. Phase 3 included structural and hydrogeological field investigations as well as supervision of targeted drilling to improve understanding of the structural, geotechnical and hydrogeological regime across the proposed open pit terrain, hydrogeological installations, rock laboratory test work, and QA/QC of collected data. Phase 4 included geotechnical and hydrogeological analysis, modeling, pit slope design, final reporting and is summarized below (refer to SRK: Pre-feasibility Slope Geotechnical and Hydrogeological Report for the Arctic Deposit for further details).

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The following conclusions were issued for the Pre-feasibility Slope Geotechnical and Hydrogeological Report for the Arctic Deposit.

Data Sources

Following the compilation of previous work and information collected during 2015 and 2016 field season, a robust geotechnical, structural and hydrogeological dataset was developed for use in the ongoing Arctic slope design studies.

Five dedicated geotechnical-hydrogeological drill holes were completed at Arctic during the 2015 and 2016 field season. Geotechnical logging was completed on a further 15 resource drill holes. This work was complemented by structural mapping, acoustic televiewer surveys, and hydrogeological installations. Laboratory strength testing was completed on resource and geotechnical-hydrogeological drill holes.

With the completion of the 2016 work, SRK believes that the Arctic open pit program satisfies full pre-feasibility level study requirements.

Geology and Structure

As a framework for the Arctic Deposit rock mass assessment, SRK has considered the existing 2016 Trilogy lithology model in conjunction with interval data from the geological drilling database.

The talc alteration occurs as continuous to semi-continuous massive bands that may host economic mineralization many metres thick, to much thinner (2 to 20 cm) bands parallel to the dominant foliation within more competent quartz-mica and quartz-chlorite mica schists. Both occurrences represent potential weak or slip foliation surface. Trilogy Metals has developed talc wireframes delineating the spatial distribution and extent of the main talc-rich horizons at the Arctic Deposit. Nine stratabound layers of talc within four major stratigraphic packages have been defined.

Pervasive weathering is present in the upper levels of the Arctic Deposit. SRK has reviewed the geotechnical data and core photographs in order to define a base of weathering isosurface that represents the boundary of upper, more pervasive weathering.

SRK completed a detailed update of the 3-D structural model (Figure 9-7). The model includes refinement of existing major structures and generation of new structures. A “structural matrix” was created which provides information on the physical properties and confidence of major and minor structural features.

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Figure 9-7 SRK Structural Model used in the Slope Stability Analysis

The structural model is considered adequate for use in pre-feasibility level studies.

Six structural and geomechanical domains were identified (Figure 9-8). These domains, each containing discontinuity sets and major structures, formed the basis of the kinematic assessment.

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Figure 9-8 Six structural and geomechanical domains were identified.

 

Rock Mass Assessment

Based on similar geotechnical conditions, the majority of lithological units have been grouped together into broad domains represented by the Upper and Lower Plates (separated by the Warm Springs Fault). The exceptions to these groupings are the weaker talc units, shallow weathered material, and fault zones. The following rock mass domains were defined:

· Upper Plate

· Lower Plate

· Weathered

· Talc Zone

· Fault Zones.

Mean rock mass parameter values and ranges were defined for each rock mass domain (e.g., fracture frequency, rock mass rating). Particular attention was paid to the assessment of intact rock strength within the defined rock mass domains. Laboratory strength testing, supported by point load testing and empirical field estimates, suggests that strength within the various lithological groups of the Upper and Lower Plates is reasonably homogeneous (Table 9-3).

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Table 9-3     Summary of derived rock mass parameter values per rock mass domain.

Domain	RQD (%)	FF/m	UCS MPa	RMR89 Joint Condition (0-30)	RMR89	GSI	Ei (GPa)	mi
Upper Plate	90	1-4	50-60	20	60-70	55-65	18	18
Lower Plate	80	2-5	50-60	19	55-65	50-60	18	18
Weathered	50	15	40-50	16	40-50	35-45	18*	13**
Talc Zone	30	15	5-10	16	35-45	30-40	ND	ND
Faults	0	40	30-40	16	30-40	25-35	ND	ND

*estimated parameter

**after BGC (2012)

ND – no data available; estimated for numerical modelling

The Talc Zone domain consisting of talc schist (TS) and chlorite-talc schist (ChTS), represents the weakest rock type (outside of fault zones) observed at the Arctic Deposit. The domain is characterized by low intact rock strength, well developed S₁/S₀ fabric and low shear strength discontinuity surfaces. The extent and persistence of the unit is of concern for pit slope stability. Future updates to the talc model should be re-evaluated by the geotechnical team to assess how they could impact the pit walls.

Kinematic Stability Assessment

A complete assessment of bench and inter-ramp kinematic stability was undertaken. Full descriptions of toppling, planar, and wedge instability risks were provided per geomechanical domain and design sector.

The most significant discontinuity sets, in terms of limiting slope angles, are related to shallow to intermediate dipping S₁/S₀ fabric which impacts the northeast, east and southeast slopes.

Hydrogeology

The 2016 field program was successfully completed, and the hydrogeology database was updated. The hydrogeological database is sufficient for pre-feasibility study purposes.

· There are 22 hydraulic conductivity measurements available from 10 drill holes, including results from a 12-hour airlift test. Bulk hydraulic conductivity ranges between 3x10^-9 and 3x10^-6 m/s. The previous interpretation of increasing K with depth is still possible but is not as consistent as previously believed.

· Water level measurements are available from 18 locations, including eight locations with just less than one full year of continuous record each. Water level data show variable elevations and different responses to recharge events suggesting the presence of multiple water systems. Seasonal variations in water levels are up to 120 m.

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The hydrogeological conceptual model for the Arctic Deposit has been updated with results gained from the 2015/2016 field programs. Data indicate a system generally characterized by low hydraulic conductivity and average recharge, but also multiple water systems. Two alternative conceptual models should be considered going forward, either of which can explain the observed data:

· Multiple water system model: this model includes two distinct groundwater systems: 1) at least one (regional) shallow perched water table in the northeast areas of the pit, and 2) a deeper water table existing over most of the pit footprint

· Compartmentalized water system model: The observed spatial distribution in water levels is the results of compartmentalization by features such as faults and/or talc surfaces and/or permafrost. Compartments may be hydraulically isolated. This model is not as clearly defined as the multiple water system model, but is considered possible. At this stage, the talc surfaces are considered to be the most likely cause of compartmentalization. Groundwater pressures could be confined below the talc surfaces (Figure 9-9).

Figure 9-9     SRK Preliminary Hydrology Model – talc Confinement Model

Pore pressure conditions were estimated for use in slope stability modelling. Models were constructed in 2-D, using the sections selected for slope stability analyses, and calibrated to observed water level data. Predictive models were completed for conservative pore pressure conditions and sensitivities completed by varying hydraulic conductivity in different lithologic units and damage zones, and by varying recharge. Slope stability models were stable under all conditions suggesting final slopes are relatively insensitive to pore pressure conditions.

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The two sources of water to the pit will be surface runoff or precipitation, and groundwater inflow. Estimates of groundwater inflows to the pit were completed using analytical solutions. Ranges of hydraulic conductivity and recharge were assessed to determine sensitivity to inputs. Total inflow is estimated to be about 3,800 m³/d, with a range of ± 1,600 m³/d. Inflow rates of this magnitude can be managed by in-pit sumps.

Stability Modelling

The minimum acceptable factor of safety (FoS) for the Arctic open pit varies depending on the component of the pit slope. Based upon the current plans, there is no major infrastructure set to be constructed proximal to any pit walls. If this were to change, it would be necessary to examine the selected acceptance criteria. Table 9-4 summarizes the selected acceptance criteria for the Arctic pit slope design.

Table 9-4     Selected acceptance criteria

Slope Component	Acceptance Criteria
	FoS (min)	PoF (max) P[FoS ≤ 1]
Bench	1.1	50%
Inter-ramp	1.3	10%
Overall	1.3	5%

Phase2 modelling result validated the findings from the kinematic assessment and suggest that final pit wall slopes are relatively insensitive to pore pressure conditions. A seismic hazard assessment has not been completed on the pit, but seismic events could be a potential risk to ongoing stability.

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Table 9-5     Summary of slope modelling results

Section ID	Section Location	SRF	Instability Mechanisms	Design Considerations
A-A’	North wall (Domain 2L-E & 3)	1.5	Planar sliding on J2 + Flexural toppling on J3b	This area should be confirmed with 3-D modelling if the slope geometry remains similar (i.e. convex slope)
B-B’	East wall (Domain 3)	1.3	Planar sliding on J1 (S1/S0)	It is unlikely that benches will be mined on this slope. The slope is likely to be stripped along J1 (S1/S0) fabric with geotechnical berms spaced every 60m. Ramps should not be constructed along this slope.
C-C’	South wall (Domain 4U & 4L)	1.4	Toppling on J2	J6 was not included in this model as a basal plane for toppling. It will be important to insure that the set does not daylight in this section to allow for toppling otherwise a shallower angle will be required for the lower stack.
D-D’	South wall (Domain 4U & 4L)	1.3	Toppling over J2 on top of the talc intersection	Slopes that have partial sections of talc in the final wall can cause instability risk.

Slope Design

Slope design recommendations are based on the findings of the kinematic evaluation, with additional adjustments from the finite element analyses. Hydrogeology was considered based on the predictions of 2-D groundwater models that suggested stability of the final slopes were relatively insensitive to static pore pressure. Slope angles have been determined for each sector. (Figure 9-10).

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Figure 9-10    SRK Design Sectors and the Recommended Range of Inter-Ramp Angles to be used in the Slope Design

· East Wall: Slopes subparallel to the main foliation (J1, S₁/S₀) should be stripped along the dip of the feature. The majority of the east wall will be mined with stacks consisting of two benches (height of 60 m) carrying a geotechnical berm or ramp every 120 m vertical spacing.

· The slope is within the acceptance criteria at friction angles as low as 22°, however if there is more talc than currently modelled (i.e. friction angle ~ 17°) then the slope will not meet the requirements.

· Slopes subparallel to the main foliation (J1, S₁/S₀) should be stripped along the dip of the feature. However, where the dip of J1 increases beyond 35°, the slope should be converted to a standard bench/berm configuration.

· North Wall: The current pit design includes a segment of convex slope along the north wall. Although 2-D modelling and kinematic analyses suggest such a design is stable, potentially complex failure modes may impact this area (possibly including J1 and other minor/major structure).

A complete investigation of early pit phase slopes has not been completed as part of this phase of study, but initial findings show that there is a potential risk if the slopes are left with thick bands of talc in the toe.

9.7.2 Acid-Base Accounting Studies

Five sampling campaigns in 1998, 2010, 2013, 2015 and 2016 resulted in accumulation of a dataset of 1,557 samples tested using various methods. In January 2017, following a large (1004 sample) infill sampling campaign in 2016, Trilogy Metals consolidated the databases and SRK used mineralogical data to develop site-specific methods for acid potential (AP) and neutralization potential (NP).

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About 70% of samples were classified as potentially acid generating (PAG), which is defined as sulfide content greater than 0.1% and NP/AP<2 where NP and AP are calculated using site-specific methods. Acid rock drainage (ARD) potential occurs in all rock types to varying degrees. Ore and the associated Gray Schist have the strongest ARD potential, followed by felsic schists. The rock types with weaker ARD potential are chlorite schist and metarhyolite porphyry (MRP).

In 2017, tailings generated by metallurgical testing were tested by acid-base accounting (ABA) and found to have potential for acid generation.

9.7.3 Geochemical Kinetic Studies

Geochemical kinetic testing using six rock type composites was initiated in 2015. The composites were used to fill seven barrels (one duplicate) at the site. Leachates from the barrels were collected and tested once in 2015 and several times in 2016 and 2017. Parallel laboratory humidity cells are evaluating the weathering behavior of the composites. All of the samples have the potential to generate acid, but to date only the Gray Schist has generated acid due to its low carbonate content. The testwork has provided data on metal release rates under acidic and non-acidic conditions which have been used to estimate contact water composition.

After 40 weeks of testing in a humidity cell, the tailings did not generate acid as carbonate minerals continue to neutralize acidity generated by sulfide mineral oxidation.

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10.0 Drilling

10.1 Introduction

Drilling at the Arctic Deposit and within the Ambler District has been ongoing since the initial discovery of mineralization in 1967. Approximately 56,480 m of drilling has been completed within the Ambler District, including 39,232 m of drilling in 174 drill holes at the Arctic Deposit or on potential extensions in 27 campaigns spanning 50 years. Drilling outside the Arctic Deposit area is discussed in Section 10.8.

All of the drill campaigns at the Arctic Deposit have been run under the auspices of either: 1) Kennecott and its subsidiaries (BCMC), 2) Anaconda, or 3) Trilogy Metals and its predecessor companies, NovaGold and NovaCopper. Table 10-1 summarizes operators, campaigns, holes and metres drilled on the Arctic Deposit. All drill holes listed in Table 10-1, except 5 holes drilled in 2017 for metallurgical purposes, were considered for use in the estimate of mineral resources described in Section 14.0

Table 10-1     Companies, Campaigns, Drill Holes and Metres Drilled at the Arctic Deposit

Year	Company	No. of Holes	Metres
1967	BCMC	7	752
1968	BCMC	18	3,836
1969	BCMC	3	712
1970	BCMC	3	831
1971	BCMC	1	257
1972	BCMC	1	407
1973	BCMC	2	557
1974	BCMC	3	900
1975	BCMC	26	4,942
1976	BCMC, Anaconda	10	805
1977	BCMC, Anaconda	4	645
1979	BCMC, Anaconda	3	586
1980	Anaconda	1	183
1981	BCMC, Anaconda	2	632
1982	BCMC, Anaconda	5	677
1983	BCMC	1	153
1984	BCMC	2	253
1986	BCMC	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
2015	NovaCopper	14	3,055
2016	Trilogy Metals	13	3,058
2017	Trilogy Metals	5	790
Total	-	174	39,232

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Additional historical exploration drilling by operators other than Bear Creek/Kennecott occurred within in the VMS belt; however, a portion is unavailable or has been lost over the years. Figure 10-1 shows the locations of drill holes in the vicinity of the Arctic Deposit.

Trilogy Metals and its predecessor company NovaGold drilled 22,144 m in 79 drill holes targeting the Arctic Deposit and several other prospects within the Ambler Schist belt. Table 10-2 summarizes all of the Trilogy Metals/NovaGold tenure drilling on the Project.

Table 10-2     Summary of Trilogy/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
2015	3,055	14	AR15-132 to 145	Geotechnical-hydrogeological studies, resource infill
2016	3,058	13	AR16-146 to 158	Geotechnical-hydrogeological studies, resource infill
2017**	790	5	AR17-159 to 163	Ore sorting studies

Notes:

^**Holes drilled in 2017 are not included in the current resource estimate as they were completed for metallurgical purposes.

^***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 in the Vicinity of the Arctic Deposit (Trilogy Metals, 2017)

10.2 Drill Companies

Over the Arctic Project’s history, a relatively limited number of drill companies have been used by both Kennecott and Trilogy/NovaGold at the Arctic Deposit. During Kennecott’s work programs, 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 used Boart Longyear as their only drill contractor. Table 10-3 summarizes drill companies and core sizes used.

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Table 10-3     Drill Contractors, Drill Holes, Meterage 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
2015	Trilogy Metals	14	3,055	NQ and HQ	Boart Longyear
2016	Trilogy Metals	13	3,058	NQ and HQ	Boart Longyear
2017	Trilogy Metals	5	790	PQ	Major Drilling/Tuuq Drilling

Sprague and Henwood used company-manufactured drill rigs during their work programs on the Project. Many of their rigs remain at the Bornite Deposit and constitute a historical inventory of 1950s and 1960s exploration artifacts. The 2004 to 2011 Trilogy Metals/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. The 2015 and 2016 NovaCopper/Trilogy Metals drill programs used two helicopter-portable LF-70 core rigs, drilling HQ or NQ core. The drill was transported by helicopter to various drill pads. The 2017 Trilogy Metals metallurgical drill program used a helicopter-portable LF-90 core rig, drilling PQ core to be used in future metallurgical test work. The drill was transported by helicopter to various drill pads.

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10.3 Drill Core Procedures

10.3.1 Kennecott

There is only partial knowledge of specific drill core handling procedures used by Kennecott during their drill programs at the Arctic Deposit. All of the drill data collected during the Kennecott drilling programs (1965 to 1998) were 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 Trilogy Metals. Drill core was cut with half core submitted to various assay laboratories and the remainder stored in Kennecott’s core storage facility at the Bornite Camp. 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 Minerals 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.

10.3.2 NovaGold/Trilogy Metals

Throughout Trilogy Metals’ work programs, 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 2017) camp core-logging facilities. 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 rock quality designation (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.

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 (see Section 11.0). Occasionally, if warranted by the need for better resolution of geology or mineralization, smaller sample intervals were 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.

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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 an absolute 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 labelled for sample processing and analysis. Select specific gravity measurements are also taken and are further discussed in Section 11.0 of this Report. The remaining half core is stored on site or at Trilogy Metals; Fairbanks warehouse.

10.4 Geotechnical Drill Hole Procedures

Five HQ3 diameter diamond drill holes were completed during NovaCopper’s 2011 geotechnical site investigation program at the Arctic Deposit. The 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 an 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 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.

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Five dedicated geotechnical-hydrogeological drill holes were completed at Arctic during the 2015 and 2016 field season. Geotechnical logging was completed on a further 15 resource drill holes. This work was complemented by structural mapping, acoustic televiewer surveys and hydrogeological installations. Laboratory strength testing has been completed on resource and geotechnical-hydrogeological drill holes.

10.5 Collar Surveys

10.5.1 Kennecott

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

10.5.2 NovaGold/Trilogy Metals

Collar location coordinates have been determined in all NovaGold/Trilogy Metals drill campaigns with two Ashtech ProMark 2 GPS units using the Riley Vertical Angle Bench Mark (VABM; 611120.442E, 7453467.486N) as the base station for all surveys. Data collection times varied from 30 minutes to two 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 were 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 Tetra Tech’s 2013 site visit, 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.

10.6 Downhole Surveys

BCMC did not perform downhole surveys prior to 1971 (drill hole 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. BCMC was 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.

Downhole surveys from 2004 to 2017 were collected using either a Reflex EZ-shot camera or a Ranger single-shot tool with individual survey readings collected at the drill rig on roughly 30 to 60 m intervals. The downhole survey data show a pronounced deviation of the drill holes toward an orientation more normal to the foliation.

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10.7 Recovery

10.7.1 Kennecott

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.

10.7.2 NovaGold/Trilogy Metals

Core recovery during NovaGold/Trilogy Metals drill programs 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 10-4 shows recoveries and RQD for each of the NovaGold/Trilogy Metals campaigns exclusive of the geotechnical drill holes in 2011.

Table 10-4     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
2011	1,193	96.0	68.8
2015	3,055	91.3	69.0
2016	3,058	91.5	69.7
2017	790	95.5	75.0

10.8 Drill Intercepts

All drill holes at the Arctic Deposit are collared on surface and are generally vertically oriented, or steeply inclined in a northeast direction. The majority of holes are spaced at 75 m to 100 m intervals, but there are rare instances where holes are located within 10 m of one another. Drill holes typically intersect at approximately right angles to the generally shallow-dipping mineralized horizon. The distribution of drilling on the Arctic deposit is shown in plan in Figure 10-1 and in several isometric views in Section 14.0.

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10.9 Prospect Drilling

Significant exploration drilling has been carried out elsewhere on the Project targeting numerous occurrences along the Ambler Schist belt. Table 10-5 summarizes the drilling on the Project outside that completed on the Arctic Deposit.

Figure 10-2 shows the locations of known major prospects and drill collar locations for the Ambler District including Trilogy Metals-controlled Ambler and Bornite sequence targets. Note that some of the drill holes are located outside the current land package held by Trilogy Metals.

Table 10-5     Drill, Meterage and Average Drill Depth for Trilogy Ambler Sequence VMS Targets

Area	Drill Holes (number)	Metres	Average Depth (m)
Dead Creek/West Dead Creek	21	3,470	165
Sunshine/Bud	36	7,111	198
Snow/Ambler	11	1,527	139
Horse/Cliff/DH	22	2,277	104
Red/Nora/BT	18	2,399	133
Total	108	16,784	155

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Figure 10-2     Collar Locations and Principal Target Areas – Ambler District (Trilogy Metals, 2017)

There have only been two drill campaigns (2006 and 2012) as shown in Table 10-2 by Trilogy Metals during targeting additional prospects beyond the Arctic Deposit in the Ambler Schist belt.

Exploration in 2006 targeted a series of geophysical anomalies in the central portion of the Ambler Schist belt near the Arctic Deposit. Twelve holes totalling 3,100 m were drilled. In 2012, Trilogy Metals drilled an additional four holes totalling 1,752 m to explore the down dip extension of the Sunshine prospect. Both programs are summarized in Table 10-6 and Figure 10-3 shows the Sunshine prospect drill hole collar locations.

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Table 10-6    Trilogy Metals 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

Figure 10-3     Sunshine Prospect and Drill Hole Locations (Trilogy Metals, 2017)

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11.0 Sample Preparation, Analyses, and Security

11.1 Sample Preparation

11.2 Core

The data for the Arctic Deposit were generated over three primary drilling campaigns: 1966 to 1986 when BCMC, a subsidiary of Kennecott was the primary operator, 1998 when Kennecott resumed work after a long hiatus, and 2004 to present under NovaGold. NovaCopper, and Trilogy Metals.

11.2.1 Kennecott and BCMC

Sampling of drill core prior to 1998 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.

11.2.2 NovaGold/Trilogy Metals

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 inductively-coupled plasma (ICP) geochemistry dataset for the Arctic Deposit and ensure continuous sampling throughout the deposit.

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.

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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 Minerals Laboratories (formerly 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 Trilogy Metals 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 Minerals 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 standard fish totes 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.

Samples were logged into a tracking system on arrival at ALS Minerals, 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 49-element suite was assayed by inductively coupled plasma-mass spectroscopy (ICP-MS) methodology, following nitric acid aqua regia digestion. The copper, zinc, lead, and silver analyses were completed by atomic absorption (AA), following a triple acid digest, when overlimits.

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11.3 Acid-Base Accounting Sampling

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.

In 2015, Trilogy Metals retained SRK to provide metal leaching (ML) and ARD characterization services for the Arctic Deposit. Activities focused on three objectives: 1) construction of on-site barrel tests and parallel humidity cells, 2) expansion of the current ABA database to support future evaluation for ARD potential management for open pit mining, and 3) evaluation of the use of proxies for ABA parameters in the exploration database with the purpose of being able to use the exploration database for block modelling of ML/ARD potential, if needed. Barrel test samples were collected during July and August 2015 and eight on-site barrel tests were constructed and initiated in late August 2015. Following the set-up of the on-site barrel tests, representative composite samples were shipped to Maxxam Analytics of Burnaby, British Columbia and parallel humidity cells were initiated in late October 2015. Trilogy Metals and SRK selected 321 samples to be analyzed for a conventional static ABA package with a trace element scan using the same method as the exploration database.

In 2016, Trilogy Metals evaluated the distribution of the existing samples to select additional samples in preparation for block modelling of ML/ARD potential. A drilling program was designed, and infill samples were collected from holes in drilled in 2015 and 2016. This program was completed, and the resulting data combined with the previous datasets.

11.4 Density Determinations

Representative specific gravity (SG) determinations conducted before 1998 for the Arctic Deposit 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 drill core, 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 Minerals and Golder and Associates, and were based on short (6 to 12 cm) whole core samples. SG was 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.

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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 ALS Minerals for SG determination.

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 used 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-sample-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. A total of 266 sample pulps were also submitted to ALS Minerals for SG determination by pycnometer analysis. In total, 459 valid SG determinations were collected, ranging from 2.64 to 4.99.

Between 2015 and 2016, Trilogy Metal geologists collected SG data consistent with the 2011 campaign. The samples averaged 2.19 m in length. Samples were collected continuously within mineralized zones and within a 2 to 3 m buffer adjacent to mineralized zones.

11.5 Security

Security measures taken during historical Kennecott and BCMC programs are unknown to NovaGold or Trilogy Metals. Trilogy Metals is not aware of any reason to suspect that any of these samples have been tampered with. The 2004 to 2016 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.

11.6 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 Minerals has attained International Organization for Standardization (ISO) 9001:2000 registration. In addition, the ALS Minerals 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 (prep) and Vancouver, BC (analytical)	1998, 2004, 2005, 2006, 2012, 2013, 2015, 2016	In 2004, ALS Minerals 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 - 2016 Primary Assay Lab

11.7 Quality Assurance/Quality Control

11.7.1 Core Drilling Sampling QA/QC

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

During 2013, Trilogy Metals conducted a 26% audit of the NovaGold era assay database fields: sample interval, gold, silver, copper, zinc, and lead. This audit is documented in a series of memos (West 2013). Trilogy Metals staff did not identify and/or correct any transcription and/or coding errors in the database prior to resource estimation. Trilogy Metals 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 NovaGold era 2004 to 2008 and NovaCopper/Trilogy Metals era 2011, 2015, and 2016 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 Minerals 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, 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 Minerals, and check samples from 2004 on original samples re-analyzed at ALS Minerals.

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 (Trilogy Metals, 2017)

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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.

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 Minerals 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 Minerals. 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 Trilogy Metals 2004 to 2013 analytical results. These QA/QC reviews serve to infer the precision of the Trilogy Metals 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 Trilogy Metals 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 Minerals 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 Minerals 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 Minerals 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 Minerals for the 2004 Arctic Project.

Overall, the analytical results of analysis for gold reported by ALS Minerals 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 Minerals 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 Minerals 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 Minerals.

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 Minerals related to drill holes AR06-98 through AR06-109.

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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 Minerals 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.

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.

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The detailed review of secondary lab check sample results reported by ALS Minerals 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.

2015

Twenty nine analytical certificates from ALS in Fairbanks, Alaska were added to the NovaCopper Inc. database. Each of the certificates was reviewed for inferred precision and inferred accuracy through detailed review of field duplicate, blank, and standard assays reported within the sample batches. The analysis of the drill core sample copper, silver, lead, and zinc was performed using four acid digest ICPMS analytical methodology. Gold assays were performed using fire assay with an atomic absorption finish.

The field duplicate sample pairs were reviewed statistically and using an average relative difference comparison. The field duplicate pairs were also reviewed within scatter plots displaying the correlation within the sample pairs. The strength of the correlation is a measure of the inferred precision within the results. Any significant differences within the duplicate pairs resulted in detailed review of the sample assays and any issues were fixed where possible.

This review has found that the duplicate pairs are well correlated overall and it is inferred that there is strong precision within the reported copper, silver, gold, lead, and zinc assay results reported by ALS.

Standards and blanks are a measure of the analytical results accuracy and the blanks also serve to indicate any issues with sample contamination or instrument calibration deficiencies.

The field standard and blank instances were reviewed and defined as failing when results were in excess of plus and minus three standard deviations from the expected mean for the standard material. Failing blanks or standards were re-analyzed along with the nearby samples in order to clear up potential accuracy deficiencies and to maintain top quality assays in the database.

Detailed review of the 67 reported blank issues has inferred that with an overall passing rate of 92.5% there is overall strong accuracy within the reported low grade copper, silver, gold, lead, and zinc results. In addition the review has shown that there were no significant or unresolved issues with sample contamination or instrument calibration deficiencies.

The standards review found that overall with 95.65% of the results within the control limits, it is the author's opinion that strong accuracy can be inferred within the reported copper, silver, gold, lead, and zinc assays.

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Secondary lab check samples were analyzed at SGS Canada Inc. located in Burnaby, British Columbia, Canada. These secondary lab check samples were carefully selected to represent the data population using a random selection of five percent of the samples within percentile range groups. These check sample assays have been compared to the primary lab assays in order to review the results for bias.

Statistics of the check samples compared to the primary samples has shown strong correlation within the data pairs.

The average differences were also calculated for the check sample pairs and it was inferred that the copper grades reported by ALS were reported with a negative bias with results reported on average 0.02356% less than the SGS results; the detailed review of the difference plot shows that the inferred bias begins at the high copper grade of 4.42 Cu %. It appears that the SGS methodology is reporting minimally higher copper grades at the ore grade level and specifically above this 4.42 Cu % mark. However, the author can see that the statistics and scatter plot show strong repeatability even at the high grades and ultimately it is the author's opinion that there is no need for concern with the ALS copper result quality.

The zinc results reported at higher grade (1.07 pct and higher) were also inferred to have bias; the average difference shows a bias level of 0.068 % zinc at these grade levels. The scatter plot also shows this bias, but the correlation within the results is shown to be quite strong even at the high grades. The difference chart shows that the samples with zinc reported below 1.07 % have negligible bias. Ultimately it appears that the ore grade level methodology used by SGS produces slightly lower zinc grades compared to that of ALS, it is difficult to say which lab is correct without further testing. However, the author does not feel that further testing is necessary. Overall it is the author's opinion that the zinc results reported by ALS are not significantly biased and the ALS zinc results are of overall good quality; this opinion is strongly influenced by the strong correlation shown within the data statistics and the scatter plot.

The silver, gold, and lead check sample assays were found to show insignificant bias levels as per this review.

This QAQC review by GeoSpark has found overall good quality within the copper, silver, gold, lead, and zinc results reported by ALS for NovaCopper Inc.'s 2015 Arctic project exploration program.

2016

Thirty analytical certificates were added to the Arctic database, these were analyzed at ALS in Vancouver, BC following sample preparation by ALS in Fairbanks, Alaska. The analysis of the drill core samples for copper, silver, lead, and zinc was performed using four acid digest ICP-MS analytical methodology. Gold assays were performed using fire assay with an atomic absorption finish.

The certificates were reviewed for inferred precision and inferred accuracy through detailed review of field duplicate, blank, and standard assays reported within the sample batches.

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One of the assay certificates (VA16159436) was specific to whole Metallurgical samples (MET_WCORE). This certificate was reviewed using the internal lab QAQC data. The author found that the internal lab QAQC had all passing duplicates showing good precision within the assays and also the review found all blank and standard instances passed the labs control tests inferring that the assays were with strong accuracy.

The field duplicate sample pairs were reviewed statistically and using an average relative difference comparison. The field duplicate pairs were also reviewed within scatter plots displaying the correlation within the sample pairs. The strength of the correlation is a measure of the inferred precision within the results. Any significant differences within the duplicate pairs resulted in detailed review of the sample assays and any issues were fixed where possible.

This review has found that the duplicate pairs are well correlated overall and it is inferred that there is strong precision within the reported copper, silver, gold, lead, and zinc assay results reported by ALS.

Standards and blanks are a measure of the analytical results accuracy and the blanks also serve to indicate any issues with sample contamination or instrument calibration deficiencies.

The field standard and blank instances were reviewed and defined as failing when results were in excess of plus and minus three standard deviations from the expected mean for the standard material. Failing blanks or standards were re-analyzed along with the nearby samples in order to clear up potential accuracy deficiencies and to maintain top quality assays in the database. Initial review of the assay certificates as they were reported found a few cases of standard instances failing; for any failed instances the nearby samples were also rerun in order to potentially improve the local accuracy statement.

Detailed review of the 58 reported blank instances has inferred that with all instances passing control tests percent there is overall strong accuracy within the reported low grade copper, silver, gold, lead, and zinc results. In addition the review has shown that there were no significant or unresolved issues with sample contamination or instrument calibration deficiencies.

The standards review found that for all assay certificates where the internal lab standards were reviewed the indication is that the copper, silver, gold, lead, and zinc results are of strong accuracy. The internal standards also have shown strong accuracy overall within the copper, silver, gold, lead, and zinc primary sample assay results.

Secondary lab check samples were analyzed at SGS Canada Inc. located in Burnaby, British Columbia, Canada. These secondary lab check samples were carefully selected to represent the data population using a random selection of five percent of the samples within percentile range groups. These check sample assays have been compared to the primary lab assays in order to review the results for bias.

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Statistics of the check samples compared to the primary samples has shown strong correlation within the data pairs. The average differences were also calculated for the check sample pairs and these do not indicate significant bias.

It is the author's opinion that the copper assays on check sample pairs do not infer any bias.

Considering the complete review of check sample pair silver grades, it is the author's opinion that there is a small level of implied bias (approximately 3.38 ppm lower silver in ALS results) for higher grade (greater than 12.85 Ag ppm) silver results, but this is not to an extent where concern is merited.

The gold assays on check sample pairs show overall strong correlation and in the author's opinion there is no indication of bias in the results.

It is the author's opinion that the review of lead results for the check sample pairs shows no indication of bias in the results.

Considering the entire review of check sample zinc results, it is the author's opinion that the level of inferred bias (average of 0.12 Zn % greater in ALS results when the over limit analysis methodology was used) does not show any need for concern with the overall primary lab zinc assay result quality.

This QAQC review by GeoSpark has found overall very good quality within the copper, silver, gold, lead, and zinc results reported by ALS for Trilogy Metals Inc.'s 2016 Arctic project exploration program.

11.7.2 Acid-Base Accounting Sampling QA/QC

SRK conducted a QA/QC review of the 2010 ABA dataset for the Arctic Project in July 2011 and concluded data quality was acceptable.

11.7.3 Density Determinations QA/QC

A QA/QC review of the SG dataset for the Arctic Deposit 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.

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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 Minerals for wet-dry SG determinations after being sealed in wax. The mean difference between the NovaGold unwaxed and the ALS Minerals waxed SG determinations was 0.01.

In 2011, to check the accuracy of the wet-dry measurements, the SG for 266 pulps was determined by pycnometer by ALS Minerals (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 allowed NovaCopper to check the wet-dry measurements by estimating the SG for an ideal stoichiometric distribution of the elements into sulphide and sulphate species.

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Stoichiometric SG values were estimated for 279 sample intervals from 2008 drill core 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.

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 iron and barium (and also for the varying density within sphalerite due to the Zn:Fe ratio).

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The best fit to the data was achieved by using the multiple regression tool in Microsoft Excel on barium, iron, zinc, and copper 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:

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.

11.8 QP’s Opinion

BD Resource Consulting, Inc. (BDRC) believes the database meets or exceeds industry standards for data quality and integrity. BDRC further believes the sample preparation, security and analytical procedures are adequate to support resource estimation.

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12.0 Data Verification

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 by Tetra Tech using a Garmin Etrex 20 global positioning system (GPS) unit. The offset distances between the collar coordinates reflected in the drill hole database provided by Trilogy Metals 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 the collar positions are adequately located to form the basis of resource estimation work.

BDRC checked the locations of holes drilled to infill the preliminary economic assessment (PEA) drill pattern. Infill holes were correctly located relative to the prior drilling. All holes were compared to the LiDAR survey of the topographic surface and found to be in the correct locations. All holes are adequately located to support resource estimation.

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.

Agreement between surveyed drill hole collar elevations and the LiDAR topographic surface verifies the correctness of the digital topography.

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12.3 Core Logging Verification

Tetra Tech visited the Trilogy Metals core storage facility in Fairbanks in 2013 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 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.

BDRC made similar observations of the core logging and geology data collection. The core logging information is acceptable for resource estimation purposes.

12.4 Database Verification

The Trilogy Metals drill database has been reviewed, and no significant concerns were noted. Nine holes were randomly selected from the Arctic database representing six percent of the data. The assay grades from these holes were dumped from MineSight™ and compared to the values listed in certified assay certificates. No errors were found.

The results of previous data verifications by external QPs (SRK 2012, Tetra Tech 2013), completed for Trilogy Metals, were also reviewed. The previous data verification exercises included extensive reviews of all NovaGold drilling as well as drilling completed by previous operators. Based on the current review, BDRC believes that the data verification completed on the Trilogy Metals dataset is sufficiently robust to support resource estimation.

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/Trilogy Metals era campaigns. GeoSpark conducted QA/QC reviews of all 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.

BDRC 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 Trilogy Metals, and previously by NovaGold and NovaCopper, are reliable and suitable for the purpose of this study.

12.6 QP Opinion

It is BDRC’s opinion that the drill database and topographic surface for the Arctic Deposit are reliable and sufficient to support the current estimate of mineral resources.

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13.0 Mineral Processing and Metallurgical Testing

13.1 Metallurgical Test Work Review

13.1.1 Introduction

Metallurgical studies have spanned over 30 years with metallurgical test work campaigns undertaken at the Kennecott Research Center (KRC) in Salt Lake City, Utah, Lakefield Research Ltd., Lakefield Ontario, SGS, Vancouver, BC and ALS Metallurgy, Kamloops, B.C.

The test work conducted in 2012 and 2017 has been under the technical direction of International Metallurgical and Environmental Inc. The basis of test work was focused on a conventional process flowsheet employing crushing, grinding, bulk flotation of a copper and lead concentrate, flotation of a zinc concentrate and the subsequent separation of copper and lead values via flotation. A flowsheet for the proposed process is shown in Figure 13-1. A summary of the test work programs, dates of test work and test work objectives is shown in Table 13-1.

Table 13-1    Metallurgical Test Work Programs

Year	Laboratory	Mineralogy	Grindability	Flotation
2017	ALS Met.		Ö	Variability testing, grindability, Cu/Pb separation
2012	SGS	Ö	Ö	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	-	Ö	Cu, Pb, Zn and Ag batch rougher flotation (selective flotation procedure)
1975	KRC	Ö	-	-
1972	KRC	-	-	Cu/Pb and Zn batch rougher and cleaner, Cu/Pb separation
1970	KRC	Ö	-	-

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Figure 13-1    Proposed Copper-Lead-Zinc Flowsheet Showing Talc Pre-float

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13.1.2 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 13-2 presents a complete list of metallic minerals identified in the Arctic Deposit samples.

Table 13-2     Metallic Mineral Identified in Arctic Project Samples

Mineral	Mineral Abundance	Mineral	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

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.

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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.

· 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 (BWi) determination procedure.

Table 13-3 shows the results of the BWi tests. Mineralization from the Arctic Deposit is relatively soft, with a BWi in the range of 5.7 to 12.0 kWh/t.

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Table 13-3     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 BWi 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.

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.

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Using zinc sulphate and sodium bisulphate to suppress galena and sphalerite, 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 1970’s 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 13-4.

Table 13-4 Head Analyses

Composite	Talc	Cu (%)	Zn (%)	Pb (%)	Fe (%)	Au (g/t)	Ag (g/t)	ST (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: S^T = 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:

· Methyl isobutyl carbinol (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.

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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 13-5. 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.

Table 13-5 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 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.

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

13.1.3 Mineralogical and Metallurgical Test Work – 2012 to 2017

Introduction

Test work conducted prior to 2012 is considered relevant to the project, but predictive metallurgical results are considered to be best estimated from test work conducted on sample materials obtained from exploration work under the direction of Trilogy Metals, conducted in 2012 and 2017.

In 2012, SGS conducted a test program on the samples produced from mineralization zones 1, 2, 3, and 5 of the Arctic Deposit. 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.

SGS used QEMSCAN™, a quantitative mineralogical technique using 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 abrasion index (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. The test work conducted in 2012 at SGS forms the bases for predicting metallurgical performance of the mineralized zone in terms of recovery of copper and lead to a bulk concentrate as well as predicting zinc recovery to a zinc concentrate.

In 2017, test work at ALS Metallurgy was focused on predicting the expected performance of the proposed copper and lead separation process, which required the use of larger test samples. A pilot plant was operated to generate approximately 50 kilograms of copper and lead concentrate, which became test sample material in locked cycle testing of the copper and lead separation process. This test work allows for the accurate prediction of copper and lead deportment in the process as well as provided detailed analysis of the final copper and lead concentrates, expected from the process. Additional metallurgical test work in the form of variability samples being subject to grindability and baseline flotation tests was also completed.

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.

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The 2017 test program involved the collection of approximately 4,000 kg of drill core from five drill holes within the Arctic deposit. The core was shipped in its entirety to ALS Metallurgy for use in grinding and flotation test work. Fifteen separate composites samples were generated by crushing defined intercepts of mineralization. These samples were riffle split to generate 15 individual samples which were separately tested for grindability and flotation response, as well, a large portion of each sample was blended to make a single large composite sample for use in copper-lead separation test work. The copper-lead separation test work involved operating a pilot plant for the production of a single sample of copper/lead concentrate which was then used in bench-scale flotation testing, including open circuit flotation tests as well as locked cycle flotation tests.

The head grades of the composites from the 2012 test work are shown in Table 13-6.

Table 13-6     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

The feed grades of samples used in the 2017 test work program at ALS Metallurgy are shown in Table 13-7.

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Table 13-7     Head Grade 2017 Variability Samples and Pilot Plant Composite

Assay			Composite	Composite	Composite	Composite	Composite	Composite
Element	Symbol	Unit	1	2	3	4	5	6
Copper	Cu	%	5.05	2.06	1.67	2.25	3.68	1.02
Lead	Pb	%	1.53	0.25	0.80	0.24	1 .01	0.36
Zinc	Zn	%	7.40	1.05	2.93	3.15	5.55	1.61
Iron	Fe	%	15.0	4.6	6.6	13.1	10.6	8.0
Sulphur	S	%	24.4	3.68	4.93	16.2	13.9	9.45
Silver	Ag	g/t	64	34	43	18	69	24
Gold	Au	g/t	0.68	0.52	0.10	0.20	0.78	0.45
Arsenic	As	g/t	4350	218	525	265	3490	1755
Magnesium	Mg	%	2.69	11.2	7.51	6.26	7.16	1.92
Assay			Composite	Composite	Composite	Composite	Composite	Composite
Element	Symbol	Unit	7	8	9	10	11	12
Copper	Cu	%	1.75	3.00	5.46	4.16	2.78	1.53
Lead	Pb	%	0.58	0.68	1.37	1.24	0.40	0.07
Zinc	Zn	%	2.71	4.65	6.60	5.63	4.56	0.56
Iron	Fe	%	8.9	10.0	9.2	13.4	12.7	4.4
Sulphur	S	%	12.9	13.3	14.0	22.1	16.9	3.15
Silver	Ag	g/t	32	56	50	34	40	16
Gold	Au	g/t	0.21	0.75	0.15	0.06	0.64	0.59
Arsenic	As	g/t	497	1350	205	296	983	611
Magnesium	Mg	%	3.46	9.18	5.65	3.58	6.84	10.6

Assay			Composite	Composite	PP
Element	Symbol	Unit	13	14	Composite 1
Copper	Cu	%	1.98	2.37	2.92
Lead	Pb	%	0.30	2.43	0.86
Zinc	Zn	%	1.48	9.50	4.66
Iron	Fe	%	6.2	15.1	10.8
Sulphur	S	%	6.25	23.7	13.8
Silver	Ag	g/t	38	62	41
Gold	Au	g/t	0.26	0.84	0.56
Arsenic	As	g/t	143	1390	1315
Magnesium	Mg	%	9.61	0.81	5.93

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.

The mineral modal abundance for the composites is shown in Table 13-8.

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Table 13-8     Mineral Modal Abundance for Composite Samples – 2012

Mineral	Mass (%)
	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
Iron Oxides	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 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 13-9 shows the results of the grindability 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.

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Table 13-9     Bond Ball Mill Grindability and Abrasion Index Test Results

Sample	Mesh of Grind Size	P80 (µm)	BWi (kWh/t)	Ai (g)
2012 SGS samples
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
2017 ALS samples
Composite 1	106	106	9.0	-
Composite 2	300	228	8.6	-
Composite 3	300	232	8.1	-
Composite 4	300	226	6.6	-
Composite 5	300	233	7.1	-
Composite 6	300	233	6.1	-
Composite 7	300	223	6.2	-
Composite 8	300	234	9.0	-
Composite 9	300	236	6.4	-
Composite 10	300	237	5.3	-
Composite 11	300	225	7.2	-
Composite 12	300	234	10.3	-
Composite 13	300	229	10.1	-
Composite 14	300	231	6.4	-
PP Composite 1	300	231	7.2	-

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 and these test results for the basis for predicting copper and zinc recovery to a bulk concentrate as well as predicting zinc recovery to a zinc concentrate.

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.

Flotation test work conducted in 2017 conducted at ALS Metallurgy, was focused on a detailed evaluation of the performance of a copper and lead separation process including open circuit flotation tests and locked cycle flotation tests.

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Open Circuit Flotation Test Work

The initial flotation tests at SGS evaluated rougher flotation 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. 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 13-10.

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 lower copper recovery. This result is likely due to insufficient sample for developing optimized flotation conditions for this sample. Additional work would likely bring this result in line with other sample test results.

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. Once again, it is expected that the results of zone 1 & 2 will improve with additional test work, if sample were available.

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 13-10     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 Tailings		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
	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

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Northwest Alaska

 

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

Trilogy Metals Inc.	13-18
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Copper/Lead Separation Test Work

SGS performed preliminary open-circuit copper and lead separation tests on the bulk copper-lead concentrates produced from the locked cycle tests in open circuit flotation tests. Sodium cyanide was used to suppress copper minerals; 3418A was used as the lead collector and lime was added to adjust the pulp pH to 10. Table 13-11 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.

Trilogy Metals Inc.	13-19
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Table 13-11     SGS Open Circuit 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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

Trilogy Metals Inc.	13-20
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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 2nd Cleaner 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 1st Cleaner 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 2nd Cleaner 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

Trilogy Metals Inc.	13-21
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

2017 ALS Metallurgy

ALS Metallurgy conducted detailed copper and lead separation flotation test work using a bulk sample of copper-lead concentrate produced from the operation of a pilot plant.

The summary results of the copper and lead separation work are shown in Table 13-12 and indicate the effectiveness of the copper and lead separation process. Two locked cycle tests are shown in 13-12 and copper recovery to a saleable copper concentrate is consistently shown to be 97.4% of the available copper in the bulk concentrate. As well, approximately 86% of the lead is recovered to a lead concentrate.

Table 13-12     ALS Metallurgy Locked Cycle Copper-Lead Separation Test Results

Product	Weight %	Assay-percent or g/tonne					Distribution - percent
		Cu	Pb	Zn	Ag	Au	Cu	Pb	Zn	Ag	Au
Test 33 (with respect to the Bulk Concentrate)
Flotation Feed	100	22.7	6.13	2.61	301	1.79	100	100	100	100	100
Lead Concentrate	22.1	2.68	24.2	1.35	960	6.90	2.6	87.3	11.4	70.5	85.1
Lead 1st Cleaner Tail	11.3	25.8	1.64	2.08	138	0.39	12.9	3.0	9.0	5.2	2.5
Lead Rougher Tail	66.6	28.8	0.89	3.11	110	0.33	84.5	9.7	79.5	24.3	12.5
Copper Concentrate	77.9	28.3	1.00	2.96	114	0.34	97.4	12.7	88.6	29.5	14.9
Test 34 (with respect to the Bulk Concentrate)
Flotation Feed	100	23.3	5.93	2.61	285	1.88	100	100	100	100	100
Lead Concentrate	21.9	2.75	23.3	1.29	906	7.33	2.6	86.0	10.8	69.5	85.3
Lead 1st Cleaner Tail	11.2	26.8	1.76	2.02	132	0.44	12.9	3.3	8.7	5.2	2.6
Lead Rougher Tail	66.9	29.4	0.95	3.14	108	0.34	84.5	10.7	80.5	25.3	12.1
Copper Concentrate	78.1	29.0	1.06	2.98	111	0.35	97.4	14.0	89.2	30.5	14.7

The lead concentrate produced from the locked cycle work at ALS Metallurgy contained only about 24% lead, due to contamination of the concentrate with talc minerals. This contamination is due to the high levels of talc in the sample provided for this specific test work. Lead concentrate grades produced during the 2012 test work ranged from 41 to 59% lead using samples that had substantially lower levels of talc in the process feed.

An overall metallurgical balance for the project is summarized in Table 13-13. This table of metal recoveries is based on an expected average recovery over the entire resource based on grades and detailed results of metallurgical test work conducted in 2012 and 2017.

Trilogy Metals Inc.	13-22
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Table 13-13     Summary of Overall Metal Recovery – Arctic Deposit

Process stream	Mass %	Concentrate Grade					Metal Recoveries
		Cu %	Pb %	Zn %	Au g/t	Ag g/t	Cu %	Pb %	Zn %	Au %	Ag %
Process Feed	100.0	2.31	0.59	3.22	0.49	38
Copper Conc	7.15	29.5	0.75	3.0	0.35	240	91.2	8.7	5.7	5.2	45.1
Lead Conc	1.02	1.7	50.0	0.9	28.0	1300	0.7	80.0	0.3	58.9	34.9
Zinc Conc	4.85	1.7	0.5	59.2	0.55	49.6	3.6	4.0	91.0	5.5	6.3
Process Tailings	86.98	0.12	0.05	0.15	0.17	6	4.5	7.3	3.0	30.5	13.7

Expected Concentrate Quality

ICP assays were conducted on the copper and lead concentrates produced from the locked cycle tests at ALS Metallurgy and the zinc concentrate from the locked cycle tests at SGS. The samples are thought to represent the expected concentrate quality. The main impurity elements are shown in Table 13-14.

The results indicated that key penalty elements, as well as precious metals are typically concentrated into a lead concentrate, leaving the copper concentrate of higher than expected quality given the levels of impurities seen in the test samples.

The lead concentrate may have penalties for the high arsenic and antimony concentrations seen in the results of this test work.

Precious metal deportment into a lead concentrate is very high and should benefit the payable levels of precious metals at a smelter.

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

Trilogy Metals Inc.	13-23
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Table 13-14     Multi-element Assay Results –Lead Concentrate and Copper Concentrate

Element		Unit	Content		Distribution - percent
			Lead	Copper	Lead Con	Copper Con
Antimony	Sb	g/tonne	3470	39	96.1	3.9
Arsenic	As	g/tonne	3750	769	57.8	42.2
Bismuth	Bi	g/tonne	1920	130	80.6	19.4
Cadmium	Cd	g/tonne	105	200	12.8	87.2
Chlorine	Cl	g/tonne	60	70	19.4	80.6
Copper	Cu	%	2.75	29.0	2.6	97.4
Fluorine	F	g/tonne	7780	750	74.4	25.6
Gold	Au	g/tonne	7.33	0.35	14.7	85.3
Lead	Pb	%	23.3	1.06	86.0	14.0
Magnesium	Mg	%	9.82	0.74	78.8	21.2
Mercury	Hg	g/tonne	2	1	35.9	64.1
Selenium	Se	g/tonne	2980	370	69.3	30.7
Silver	Ag	g/tonne	906	111	69.6	30.4
Sulphur	S	%	7.54	33.2	6.0	94.0
Tellurium	Te	g/tonne	54.7	6.4	70.6	29.4
Zinc	Zn	%	1.29	2.98	10.8	89.2

Within the zinc concentrates produced at SGS in 2012 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 shown in Table 13-15. 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. Silica analysis of a zinc concentrate produced from a bulk sample test showed low silica levels of 0.89% Si.

Table 13-15     Multi-element Assay Results – Zinc Concentrate

Test Composite	LCT 1	LCT 2	LCT 3	LCT 4	LCT 5	LCT 6
	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(ppm)	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

Trilogy Metals Inc.	13-24
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

13.2 Recommended Test Work

In general, the flowsheet developed in the 2012 test program and further tested in the 2017 test work program at ALS Metallurgy, is feasible for the Arctic mineralization (see Figure 13-1). Further metallurgical test work is recommended on representative samples to optimize the flowsheet and better understand the impact of talc levels in the process feed. Lead concentrate quality may be impacted by the level of talc in the process feed and a better understanding of the level of talc in an expected process feed is critical in maximizing the value of a lead concentrate. Test work to investigate the impact of talc is currently underway at ALS Metallurgy. There are no outstanding metallurgical issues related to the production of a copper or zinc concentrate from all of the materials tested.

On-going grinding test work is recommended prior to the commencement of a feasibility study, including SAG mill characterization test work, concentrate thickening testing and filtration testing.

Trilogy Metals Inc.	13-25
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

14.0 Mineral Resource Estimate

14.1 Introduction

This section describes the generation of the mineral resource estimate for the Arctic Deposit. The estimate has been prepared by Robert Sim, P.Geo. and Bruce M. Davis, FAusIMM. During the summers of 2015 and 2016, Trilogy Metals conducted drilling programs designed to upgrade previous in-pit Inferred Mineral Resources to the Indicated category. During the fall of 2016, following the completion of the final drilling program, Trilogy Metals geologists reinterpreted the geologic units present in the vicinity of the Arctic Deposit. This section incorporates the new geologic model and all available sample data as of April 25, 2017.

This section describes the resource estimation methodology and summarizes the key assumptions considered by the QPs. In the opinion of the QPs, the resource evaluation reported herein is a sound representation of the mineral resources for the Arctic Deposit at the current level of sampling. The mineral resources have been estimated in conformity with the generally accepted 2003 CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines and are reported in accordance with NI 43-101. Mineral resources are not mineral reserves and do not have demonstrated economic viability.

The database used to estimate the Arctic Deposit mineral resource was audited by the QPs. The QPs are of the opinion that the current drilling information is sufficiently reliable to confidently interpret the boundaries of the mineralization and the assay data are sufficiently reliable to support mineral resource estimation.

The resource estimate was generated using MineSight® v11.60-2. Some non-commercial software, including the Geostatistical Library (GSLib) family of software, was used for geostatistical analyses.

14.2 Sample Database and Other Available Data

Trilogy Metals provided the Arctic database in Microsoft™ Excel format, exported from the master database (GeoSpark Core Database System). The files contain collar, survey, assay, lithology, ABA and specific gravity data, and other geological, and geotechnical information.

The Project drilling database comprises 322 diamond drill (core) holes totalling 64,260 m, this includes exploration holes that test for satellite deposits for distances up to 40 km from the Arctic Deposit and includes information from 40 drill holes located outside of the Trilogy Metals property. There are 152 drill holes (32,699 m) in the immediate vicinity of the Arctic Deposit that have been used to develop the estimate of mineral resources described in this report.

Trilogy Metals Inc.	14-1
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

The database contains a total of 12,594 samples, most of which have been analyzed for a variety of elements through a combination of ICP and XRF multi-element packages. Sample data for copper, lead, zinc, gold and silver have been extracted from this database for use in the generation of this resource estimate. Note: The number and total length of drill holes described here represents the database used to generate the estimate of mineral resources. These numbers may differ from those described in previous Trilogy reports.

Individual sample intervals range from 5 cm to 35.5 m in length and average 2.14 m. The few very long sample intervals represent samples taken in talus and overburden. Sample selection in the majority of drill holes has been guided by the visual presence of appreciable amounts of sulphide mineralization. As a result, most core intervals where samples have not been taken are assigned default zero grade values. There are exceptions where samples were purposely not taken, such as wedge holes or holes that were drilled to provide metallurgical test material. In these cases, the un-sampled intervals remain as “missing”.

All drill holes at the Arctic Deposit are collared on surface and are generally vertically oriented, or steeply inclined in a northeast direction. The majority of holes are spaced at 75 m to 100 m intervals, but there are rare instances where holes are located within 10 m of one another.

SG measurements have been conducted on 3,024 samples in the database and range from a minimum of 2.43 to a maximum of 4.99 and average 3.08. The distribution of SG data is considered sufficient to support block model estimation.

Drill core recovery data is available for 107 holes with an overall average value of 94%. Samples in the interpreted mineralized domains average >95% recovery. There are no apparent relationships between drill core recovery and sample grade. There are no adjustments to the sample database to account for core recovery.

The database also contains lithology information derived during core logging. There are 33 different rock types in this dataset.

Trilogy Metals provided a topographic digitalterrain surface, produced from LiDAR data in 2016, measuring approximately 2 km east-west by 2 km north-south that is centred over the Arctic Deposit. Drill hole collar locations, surveyed using a differential GPS, correlate very well with the local digital terrain (topographic) surface.

The distribution of copper grades in drill holes proximal to the Arctic Deposit is shown from two isometric viewpoints in Figure 14-1 and Figure 14-2.

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NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Figure 14-1     Isometric View of Copper Grades in Drill Holes (Sim, 2017)

Trilogy Metals Inc.	14-3
NI 43-101 Technical Report on the Arctic Project,
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Figure 14-2     Isometric View of Copper Grades in Drill Holes (Sim, 2017)

14.2.1 ABA Data

Also included in the resource model are items used to evaluate the acid generating and neutralizing potential of the rocks in the vicinity of the deposit. This includes 1,557 samples that have been analyzed for AP and NP. The distribution of AP and NP data, shown in Figure 14-3, is somewhat limited due to a lack of available drill core. The majority of the available AP and NP samples are located around the perimeter of the deposit and from rocks in the hanging-wall to the mineralized zones. Although the distribution of these data is not ideal, it is felt there is sufficient information available to provide reasonable estimates of the acid and neutralizing potential of the waste rocks at the Arctic Deposit. There have been no adjustments to account for missing AP and NP data.

In addition to estimates of AP and NP in the model, estimates of total sulphur content are also generated. There are a total of 9,316 samples that have been analyzed for total sulphur content. Approximately one half of drill holes have sulphur analysis throughout the entire length of the hole and the remainder of the holes have sulphur analyses taken on 10 m intervals down the hole. This provides a consistent and extensive distribution of samples that is sufficient to provide reasonable estimates of sulphur content in the block model.

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NI 43-101 Technical Report on the Arctic Project,
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Figure 14-3     Isometric Views of Available AP and NP Data (Sim, 2017)

Figure 14-4     Isometric View of Available Sulphur Data (Sim,2017)

Table 14-1 contains a summary of the sample data used in the development of the Arctic Deposit resource block model. Note that the primary and adjusted values for copper, lead, zinc, gold and silver are included in the table (value #1 is initial data and #2 includes zero grade values assigned to select un-sampled intervals).

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Table 14-1     Summary of Sample Data Used to Develop the Resource Block Model

Element	Number	Total Length (m)	Minimum	Maximum	Mean	Std. Dev.	Co. Of Variation
Copper1 (%)	12,252	17,551	0.00	31.00	0.50	1.67	3.3
Copper2 (%)	15,662	31,392	0.00	31.00	0.28	1.28	4.5
Lead1 (%)	12,041	17,361	0.00	8.15	0.12	0.50	4.0
Lead2 (%)	15,451	31,202	0.00	8.15	0.07	0.38	5.4
Zinc1 (%)	12,151	17,458	0.00	27.60	0.72	2.56	3.6
Zinc2 (%)	15,561	31,299	0.00	27.60	0.40	1.95	4.8
Gold1 (g/t)	10,986	14,604	0.00	32.800	0.138	0.783	5.7
Gold2 (g/t)	14,396	28,446	0.00	32.800	0.071	0.565	8.0
Silver1 (g/t)	12,154	17,459	0.00	1,155.00	8.20	30.58	3.7
Silver2 (g/t)	15,564	31,300	0.00	1,155.00	4.57	23.20	5.1
Sulphur (%)	8,937	15,450	0.01	10.00	1.37	2.18	1.6
AP	2,261	5,018	0.31	1,307.50	68.19	148.50	2.2
NP	2,261	5,018	0.08	972.75	18.34	50.54	2.8
SG	3,100	na	2.43	4.99	3.09	0.53	0.2

Notes: Value#1 is initial sample data. Value#2 includes zero grades assigned to select unsampled intervals.

The total core length of drilling is 32,699m.

14.3 Geologic Model

Trilogy Metals geologists have interpreted three dimensional domains representing the distributions of various lithologic units, mineral domains, alteration facies, geotechnical domains, talc-rich zones and an area of near surface weathering. All of these domains were evaluated to determine if they should be used to control the estimation of the various elements included in the resource block model.

In order to replicate the stratiform nature of the mineralization in the resource model, a dynamic anisotropy approach relative to the overall trends of sulphide mineralization has been applied. Three-dimensional planes are interpreted that represent the trends of the sulphide mineralization, with separate planes interpreted for each of the eight main mineralized domains. These “trend planes” generally represent the centre of each interpreted mineralized domain. These trend planes are used to control search orientations during subsequent interpolations in the model. Variograms are generated using distances relative to the trend planes rather than the true sample elevations. This approach essentially flattens out the zone during interpolation relative to the defined trend plane.

The interpretation of most of the geology domains are derived from a combination of information recorded during surface geologic mapping and the visual logging of drill core as well as properties exhibited by various elements in the ICP database. A series of mineralized zones (MinZone domains) have been interpreted by Trilogy Metals that represent zones that exceed a grade of 0.75% copper equivalent (CuEq). Of these, there are four or five primary domains and 12 sub-domains. The sub-domains are much smaller and often are interpreted about only one or two drill holes. Essentially all of the mineral resource is located within the larger, primary, MinZone domains. Several examples of the interpreted lithologic model are shown in Figure 14-5 and Figure 14-6.

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Figure 14-5     Cross Section 613250E Showing Lithology Domains at Arctic (Sim, 2017)

Figure 14-6     Cross Section 7453000N Showing Lithology Domains at Arctic (Sim, 2017)

Six separate geotechnical domains have been interpreted by SRK, based on a review of the local geology, alteration, weathering, overburden, major structures, minor structures (discontinuity sets), a rock mass assessment, a kinematic stability assessment and a hydrogeological assessment. These domains define differing slope sectors used in the generation of open pit designs. The distribution of these domains is shown in Figure 14-7.

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Figure 14-7     Isometric View of Geotechnical Domains (Sim, 2017)

A series of nine separate domains have been interpreted that encompass zones where the presence of talc has been of observed. The shape and distribution of the talc domains are shown in Figure 14-8. These domains tend to mimic the trends of mineralization in the deposit. Some talc domains consist of small and discontinuous patches.

Figure 14-8     Isometric Views of Talc Domains (Sim, 2017)

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Four alteration domains have been interpreted as shown in Figure 14-9. These tend to mimic the general trends of mineralization in the deposit. These domains are locally patchy and discontinuous, reflecting a lack of continuity of these alteration assemblages.

Figure 14-9     Isometric Views of Alteration Domains (Sim, 2017)

14.3.1 Summary of Geologic Domains

The interpreted lithology domains are summarized in Table 14-2. The lithology domains have been segregated into five general groups as follows:

LTHDM 1-8: Somewhat generalized representation of the mineralized domains (MinZone domains) that host the majority of the mineralization at the Arctic Deposit.

LTHDM 100-series: Meta-Rhyolite Porphyry (MRP)

LTHDM 200-series: Grey Schist (GS)

LTHDM 300-series: Quartz Mica Schist (QMS)

LTHDM 400-series: Aphanitic Meta-Rhyolite (AMR)

Table 14-2     Summary of Lithology Domains

Lithology Unit	LTHDM	Lithology Unit	LTHDM	Lithology Unit	LTHDM	MinZone Domain	LTHDM
AMR 2A	401	GS WS	209	MRP WS	104	1	1
QMS 2WS	309	GS 2C	208	MRP WSsub	103	2, 2.5	2
QMS WS	308	GS 2B	207	MRP 2A	102	3, 3sub	3
QMS 2B	307	GS 2A	206	MRP 1A	101	4	4
QMS 2A	306	GS 2	205			5	5
QMS 1A	305	GS WX	204			7a, b, bHW, c	7
QMS 1CX	304	GS X	203			8a, b, c, cHW, d	8
QMS 1BY	303	GS Y	202
QMS 1CZ, 1C	302	GS Z	201

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In order to retain the detail between the various MinZone domains, distinct codes have been assigned using the individual interpreted MinZone domains as listed in Table 14-3.

Table 14-3     Summary of Mineralized Zone (MinZone) Domains

MinZone Domain	MNZNE code	MinZone Domain	MNZNE code
1	10	7b	73
2	20	7bHW	74
2.5	25	7c	75
3	30	7cHW	76
3sub	35	8a	81
4	40	8b	82
5	50	8c	83
7a	71	8cHW	84
7aHW	72	8d	85

The remaining interpreted geotechnical, alteration, talc and weathered domains are summarized in Table 14-4.

Table 14-4     Summary of Geotech, Alteration, Talc and Weathering Domains

Geotech Domain	GTECH Code	Alteration Domain	ALTDM Code	Talc Domain	TALC Code	Weathered Domain	Weathered Code
2L-E	1	FW Chlorite	701	Talc	1	Weathered	1
2L-W	2	FW Chlorite-Sericite	702	No Talc	2	Fresh	2
2U	3	Intense Magnesium	703
3	4	Sodium Enrichment	704
4L	5	Sodium Enrichment HW	705
4U	6	Other	706

14.4 Compositing

Compositing drill hole samples standardizes the database for further statistical evaluation. This step eliminates any effect the sample length may have on the data. To retain the original characteristics of the underlying data, a composite length that reflects the average, original sample length is selected: a composite that is too long can sometimes result in a degree of smoothing that can mask certain features of the data.

At the Arctic Deposit, the average sample length of all samples is 1.45 m but inside the MinZone domains samples tend to be much shorter, with an average of 0.68 m. A composite length of 1 m has been selected for use in the estimate of mineral resources.

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Drill hole composites are length-weighted and are generated down-the-hole, meaning composites begin at the top of each drill hole and are generated at constant intervals down the length of the hole. The drill hole composites honour the MinZone domain boundaries, meaning individual composites are broken at the boundary between the MinZone domain and the surrounding rocks.

14.5 Exploratory Data Analysis

Composited samples were captured in the various interpreted domains including the lithology domains (including the Minzones), alteration domains, talc domains and the near-surface weathered domain.

14.5.1 As-Logged Geology and Domain Statistics

This section examines the relationship between metal content and the as-logged lithology units in the database. The drill core was examined and logged for lithology type and geo-technical characteristics. The geotechnical groups were not related to grade and are, therefore, not included in this discussion.

Twenty-seven lithology designations with associated grades occur in the database. The frequency distributions for the grades of each metal by as-logged lithology are compared by a series of boxplots. An example for copper appears in Figure 14-10. The boxplots show that significantly high grades occur, as expected, in massive and semi-massive sulphides but it also shows that high grades may occur in almost any lithology. These results suggest that individual lithology type is not a strong controlling factor in the distribution of metal in the deposit.

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Figure 14-10     Boxplots of Copper by Logged Lithology Type (Sim, 2017)

A matrix was constructed listing the individual logged lithology types within each of the interpreted MinZone domains. This matrix indicated that between 32% and 44 % of the mineral domains consisted of massive or semi-massive sulphides. The remainder consisted of as many as 15 other rock types. After all, the Minzone domains encompass rocks which, in general, contain greater than 0.75% CuEq and, as shown in the boxplot in Figure 14-10, mineralization of this tenor occurs in the majority of rock types.

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Similarly, Trilogy Metals grouped the 27 individual lithology types into four groups (QMS, GS, MRP and AMR). Each of these four groups contains a mix of logged rock types. The matrix of individual lithology by group showed the interpreted MRP group contained 54.7% individually-logged MRP rocks with the remainder from 24 other logged rock types; the interpreted GS group had 76.2% logged as GS and 16 other logged rock types; the interpreted QMS group had 79.5% logged as QMS and 26 other rock types; and the interpreted AMR group had 86.4% logged as AMR and five other rock types. This type of simplification of rock types resulting from the interpretation of lithology domains is not uncommon.

14.5.2 Interpreted Lithology and MinZone Domain Statistics

The composited sample data were assigned distinct lithology domain codes, as listed in Table 14-2, using the domains interpreted by Trilogy Metals. Boxplots describing the distributions of each element by lithology domain were generated. The distributions for copper, zinc, lead, gold, and silver are similar relative to the Minzone domains; the interpreted MinZone domains (lithology domain codes 1to 8) host the majority of the mineralization where the other lithology domains (100, 200, 300 and 400 series codes) only exhibit a few rare significant grade values of which there is no apparent continuity. Domains 7 and 8 show elevated metal grades compared to the other lithology groups, but the important grade distributions occur in domains 1 to 5. The distributions of copper and gold by lithology domain are shown in Figure 14-11 and Figure 14-12.

Figure 14-11     Boxplots of Copper by Lithology Domain (Sim, 2017)

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Figure 14-12     Boxplots of Gold by Lithology Domain (Sim, 2017)

The MinZone domains interpreted by Trilogy Metals were used to assign MinZone domain codes, as listed in Table 14-3, to the composited drill hole samples. A series of boxplots were generated for each of the five metals included in the resource model. There are similar relative distributions exhibited by each of these five metals among or across the Minzone domains. An example showing the distribution of copper by MinZone domain is shown in Figure 14-13. The primary domains are those enclosing appreciable volumes of sample data (domains 1, 2, 2.5, 3, 4 and 5). There are limited numbers of data, and much lower average grades, in the sub-domains (3sub, 7’s and 8’s). For statistical and estimation purposes, the data in MinZone domain 2.5 is combined with domain 2, and the data in domain 3sub has been combined with domain 3. Since the frequency distributions are fairly similar in the 7-series, and there are relatively few samples in each sub-domain in the 8-series, the smaller domain samples were grouped into two domains labeled 7 and 8 for statistical and estimation purposes.

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Figure 14-13     Boxplots of Copper by MinZone Domain (Sim, 2017)

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Evaluations were also made comparing the five main metals relative to the geotechnical, alteration, talc and weathering domains. There are no indications that these domains control the distribution of copper, lead, zinc, gold or silver in the deposit.

A series of boxplots was produced comparing the AP, NP and sulphur sample data in relation to the interpreted MinZone, lithology, alteration, talc and weathered domains. As expected, higher AP and S% values occur in the MinZone domains. Higher NP values in the vicinity of the MinZone domains are likely the result of the talc alteration from a carbonate-rich protolith typically seen in these areas. Outside of the mineralized domains, most lithology groups tend to have similar distributions of AP, NP, and S%. The grey schist has elevated values of AP and S% compared to other domains. The boxplots in Figure 14-14 show the distributions of AP, NP and S% by lithology type.

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Figure 14-14     Boxplots of AP, NP and Sulphur by Lithology Domain (Sim, 2017)

The alteration domains show only minor differences between domains. The sodium depletion domain has lower sulphur and AP values, and there are higher NP values in the FW chlorite and magnesium enrichment domains, but there is significant overlap in the boxplot results between domains suggesting these are not distinct distributions.

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There are differences evident in AP, NP, and S% between both talc and the weathered domains as shown in Figure 14-15 and Figure 14-16.

Figure 14-15     Boxplots of AP, NP and Sulphur by Talc Domain (Sim, 2017)

Figure 14-16     Boxplots of AP, NP and Sulphur by Weathered Domain (Sim, 2017)

SG samples were evaluated between the various interpreted domains. Only the MinZone domains contain samples that significantly differ from SG samples in the surrounding rocks. The boxplot in Figure 14-17 shows the distribution of SG data between the various MinZone domains and the lithology groups. There is weak correlation evident between SG and copper and zinc grade in some MinZone domains but there is scatter due to the variable presence of chalcopyrite, sphalerite, barite, and galena as well as arsenopyrite, pyrite and pyrrhotite.

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Figure 14-17     Boxplots of SG by MinZone and Lithology Group Domains (Sim, 2017)

14.5.3 Contact Profiles

Contact profiles evaluate the nature of grade trends between two domains: they graphically display the average grades at increasing distances from the contact boundary. Those contact profiles that show a marked difference in grade across a domain boundary indicate that the two datasets should be isolated during interpolation. Conversely, if a more gradual change in grade occurs across a contact, the introduction of a hard boundary (e.g., segregation during interpolation) may result in a much different trend in the grade model; in this case, the change in grade between domains in the model is often more abrupt than the trends seen in the raw data. Finally, a flat contact profile indicates no grade changes across the boundary; in this case, hard or soft domain boundaries will produce similar results in the model.

Contact profiles were generated to evaluate the change in grades across prominent lithologic group and mineralized (MinZone) domain boundaries. The results for all metals are similar; a marked change in grade between the MinZone domains and the surrounding host rocks. An example showing the change in copper grade between the (combined) MinZone domains and the three main lithology groups is presented in Figure 14-18.

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Figure 14-18     Contact Profiles of Copper Between MinZone and other Lithology Domain Groups (Sim, 2017)

Contact profiles were generated to evaluate the change in AP, NP, and S% across prominent domain boundaries.

Even though the talc and weathering surface show the frequency distributions of AP, NP, and S% are different inside and outside of the domains, contact profiles show these variables tend to be similar or transition at the boundary. The contact profiles for AP, NP and sulphur for the weathering surface are shown in Figure 14-19 and similar profiles at the talc boundary appear in Figure 14-20.

Figure 14-19     Contact Profile of AP, NP and Sulphur Between Weathered and Fresh Rocks (Sim, 2017)

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Figure 14-20     Contact Profile of AP, NP and Sulphur Inside / Outside of the Talc Domains (Sim, 2017)

14.5.4 Modeling Implications

The results of the EDA indicate that all metal, ABA and density samples located inside the MinZone domains distinctly differ from samples outside and these data should not be mixed during estimations in the block model.

The most consistently important metal grades occur in MinZone domains 1 to 5. MinZones 1 to 5 tend to host higher-grades and are continuous over relatively large areas (several hundred metres). MinZones 7 and 8 contain lower grades and there tends to be far less continuity of mineralization.

Although the nature of mineralization may be similar between most of the MinZone domains, they each represent distinct stratigraphic mineralized horizons and, as a result, the contained sample data in each mineralized horizon should remain segregated during the interpolation of block grades in the model. Therefore, “hard” boundary conditions are applied to all MinZone domains for grade estimation purposes (even the individual small domains that comprise MinZones 7 and 8).

The rocks surrounding the MinZone domains are essentially void of appreciable mineralization and, as a result, grade estimates in the model for copper, lead, zinc, gold and silver are restricted only to the MinZone domains. It is assumed that all areas outside of the MinZone domains have zero grade values for these five metals.

The results of the EDA indicate that Grey Schist (GS) contains AP and sulphur data that differs from samples in the surrounding rock types and, as a result, this lithology type has been segregated during the estimation of these items in the block model. NP does not differ across the GS domain and, as a result, it is not honoured during the estimation of NP in the model.

There are no indications that the talc domains or the weathered zone contain any distinct properties in the distribution of metals, ABA samples or density. These domains are ignored during the development of the block model.

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There are no distinct differences in the density of rocks, other than the MinZone domains, between lithologies, alteration types, talc domains or in the weathered zone. Model blocks in the overburden domain are assigned a default SG value of 2.1.

Table 14-5 lists the domains used to estimate the various items in the in the resource block model.

Table 14-5     Summary of Estimation Domains

Item	MinZone Domain	Lithology Domain
Copper	Hard	na
Lead	Hard	na
Zinc	Hard	na
Gold	Hard	na
Silver	Hard	na
AP	Hard	Hard GS Only
NP	Hard	None
Sulphur	Hard	Hard GS Only
SG	Hard	None

Note: There are no estimates of Cu, Pb, Zn, Au or Ag outside of the MinZone domains.

In order to retain the banded nature of the distributions of items outside of the MinZone domains, the estimations of AP, NP, sulphur and SG are made using the dynamic search orientations relative to the more prominent zones of mineralization. The areas outside of the MinZone domains have been combined into four separate trend groups; a lower group parallel MinZone 1, a middle group parallel to MinZone 3, an upper group parallel to MinZone 5 and a fourth group located above the Warm Springs fault.

14.6 Treatment of Outlier Grades

Measures have been taken to control the effects of potential outlier sample data for copper, lead, zinc, gold and silver. There is no need for changes in sulphur data, as several maximum values of 10% S in the database are a reflection of the upper detection limit of the ICP technique. There are no modifications to the AP, NP or SG data prior to estimation in the block model.

Histograms and probability plots were generated from 1 m composited sample data to show the distribution of metal in each estimation domain. These were used to identify the existence of anomalous outlier grades in the composite database. The physical locations of these potential outlier samples were reviewed in relation to the surrounding data and it was decided that their effects could be controlled through the use of outlier limitations. An outlier limitation approach limits samples above a defined threshold to a maximum distance of influence during grade estimates. With the majority of the drill holes piercing the mineralization at 75 m spacing, samples above the outlier thresholds are limited to a maximum distance of influence of 40 m during block grade interpolation (approximately ½ the distance between drill holes). During the estimation of SG in areas outside of the MinZone domains, samples greater than 3.80 t/m3 are limited to a maximum distance of influence of 40 m.

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Table 14-6 summarizes the treatment of outlier sample data.

Table 14-6     Summary of Treatment of Outlier Sample Data

	Copper %		Lead %		Zinc %		Gold g/t		Silver g/t
MinZone Domain	Max.	Outlier Limit	Max.	Outlier Limit	Max.	Outlier Limit	Max.	Outlier Limit	Max.	Outlier Limit
1	17.07	10	7.84	5	25.60	20	8.080	3	501.0	300
2	9.10	6	3.27	2	15.10	10	10.850	5	141.0	100
2.5	12.40	8	5.65	4	20.30	14	6.960	3.5	285.0	200
3	18.00	12	5.84	4	20.80	17	6.857	3	542.1	200
4	17.17	8	4.56	3	17.89	16	2.307	2	467.9	150
5	17.67	15	5.00	3.8	20.90	17	32.800	15	967.5	350
7	4.29	1.5	6.65	5	25.84	10	0.832	0.7	159.0	100
8	3.66	2.5	2.47	1.5	15.00	7	5.140	0.8	341.8	100

Samples above the outlier limit were restricted to maximum range of 40 m during block grade interpolation.

The proportion of metal lost, calculated in model blocks in the combined Indicated and Inferred categories, is 3% copper, 5% lead, 4% zinc, 9% gold and 6% silver. The proportion of lost metal is a function of drill hole spacing and the nature of the underlying sample data—the more skewed distributions show higher losses, as seen in the gold model. The proportions of metal lost due to the treatment of outlier sample data are considered appropriate for a project with this level of delineation drilling.

14.7 Specific Gravity Data

Approximately 45% of the available SG data occurs inside the interpreted MinZone domains ranging from a minimum of 2.55 to a maximum of 4.99 and average 3.46. Outside of the MinZone domains, SG values range from a minimum of 2.43 to a maximum of 4.56 and average 2.78.

The base metal content and SG are moderately correlated. There is little variation in the SG values in the MinZone domains with coefficient-of-variation values that are typically less than 0.2. Outside of the MinZone domains, the coefficient of variation is 0.05.

SG data are available in approximately two-thirds of the drill holes in the vicinity of the Arctic Deposit. The distribution of SG samples varies between drill holes; about one-third of the holes have SG measurements for either every sample interval or on 10 m spaced intervals down the hole. The other third of the holes have SG measurements that are primarily restricted to the mineralized intervals.

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The distribution of SG data is considered sufficient to support estimation in the resource model. The relatively low variability in the sample data indicates that SG values can be estimated into model blocks using inverse distance-squared moving averages. The MinZone domains are used as hard boundaries during the estimation of densities in the model and the trends planes are used to control the dynamic anisotropy during the estimation of SG values in the block model.

14.8 Variography

The spatial evaluation of the data was conducted using a correlogram instead of the traditional variogram. The correlogram is normalized to the variance of the data and is less sensitive to outlier values; this generally gives cleaner results.

Many of the individual estimation domains do not contain sufficient sample data from which to generate reasonable correlograms. As a result, separate correlograms have been generated for samples inside MinZone domains 1, 3 and 5. The remaining MinZone domains (2, 2.5, 4, 7 and 8) use correlograms that have been generated using combined data from those five zones. Correlograms have been generated using 1 m composited drill hole data that has been top-cut to reduce the effects of rare anomalous high-grade composites.

Correlograms were generated using the commercial software package SAGE2001 developed by Isaaks & Co. Correlograms were generated using elevations relative to the trend planes described in Section 14.3 of this report. This ensures that the local undulations of the typically banded mineralization are replicated in the block model. The correlograms are summarized in Tables 14-7 through 14-14.

Table 14-7     Copper Correlogram Parameters

MinZone Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
1	0.121	0.686	0.193	17	62	0	361	133	0
	Spherical			4	332	0	220	43	0
				4	90	90	8	90	90
3	0.300	0.504	0.196	20	22	0	3,316	272	0
	Spherical			8	112	0	135	2	0
				6	90	90	6	90	90
5	0.140	0.352	0.509	272	50	0	97	67	0
	Spherical			16	320	0	6	90	90
				3	90	90	3	157	0
2, 2.5, 4, 7 & 8	0.033	0.800	0.167	30	67	0	449	85	0
	Spherical			5	157	0	180	355	0
				5	90	90	5	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

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Table 14-8     Lead Correlogram Parameters

MinZone Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
1	0.141	0.737	0.121	96	26	0	2,589	356	0
	Spherical			16	116	0	138	86	0
				5	90	90	6	90	90
3	0.275	0.393	0.332	10	90	90	405	66	0
	Spherical			10	43	0	112	336	0
				7	133	0	10	90	90
5	0.300	0.551	0.149	6	60	0	4,159	44	0
	Spherical			5	90	90	136	314	0
				5	330	0	8	90	90
2, 2.5, 4, 7 & 8	0.107	0.597	0.296	11	67	0	803	54	0
	Spherical			10	90	90	153	324	0
				5	157	0	4	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

Table 14-9     Zinc Correlogram Parameters

MinZone Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
1	0.102	0.737	0.162	40	346	0	461	339	0
	Spherical			16	76	0	185	69	0
				5	90	90	5	90	90
3	0.108	0.583	0.309	53	37	0	330	91	0
	Spherical			8	127	0	195	1	0
				5	90	90	10	90	90
5	0.020	0.869	0.111	14	62	0	5,151	173	0
	Spherical			8	332	0	246	83	0
				3	90	90	8	90	90
2, 2.5, 4, 7 & 8	0.203	0.530	0.267	11	71	0	313	55	0
	Spherical			11	90	90	225	145	0
				5	341	0	3	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

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Table 14-10     Gold Correlogram Parameters

MinZone Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
1	0.065	0.804	0.131	31	122	0	754	17	0
	Spherical			7	32	0	116	107	0
				5	90	90	8	90	90
3	0.072	0.502	0.426	58	47	0	348	26	0
	Spherical			6	90	90	268	296	0
				5	137	0	6	90	90
5	0.275	0.602	0.123	117	49	0	279	103	0
	Spherical			5	90	90	58	13	0
				3	319	0	5	90	90
2, 2.5, 4, 7 & 8	0.016	0.764	0.220	23	78	0	392	14	0
	Spherical			4	90	90	279	104	0
				3	168	0	5	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

Table 14-11     Silver Correlogram Parameters

MinZone Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
1	0.194	0.647	0.159	65	358	0	364	122	0
	Spherical			4	88	0	150	32	0
				4	90	90	5	90	90
3	0.228	0.400	0.372	29	58	0	373	87	0
	Spherical			12	90	90	183	357	0
				5	148	0	10	90	90
5	0.176	0.468	0.356	155	46	0	120	79	0
	Spherical			4	316	0	9	90	90
				3	90	90	4	169	0
2, 2.5, 4, 7 & 8	0.011	0.774	0.214	31	76	0	338	67	0
	Spherical			4	90	90	204	337	0
				3	166	0	5	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

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Table 14-12     Sulphur Correlogram Parameters

Domain	Nugget		S1		S2	1st Structure			2nd Structure
						Range (m)	AZ	Dip	Range (m)	AZ	Dip
MinZones	0.200		0.689		0.111	177	13	0	6,956	48	0
	Spherical					19	103	0	808	318	0
						15	90	90	15	90	90
Grey Schist	0.050		0.690		0.260	50	96	0	1360	62	0
	Spherical					12	90	90	607	152	0
						11	6	0	13	90	90
LithGroup1	0.170		0.468		0.363	273	34	0	1,060	41	0
	Spherical					63	124	0	204	311	0
						25	90	90	24	90	90
LithGroup2	0.078		0.390		0.531	169	61	0	469	8	0
	Spherical					60	151	0	347	98	0
						10	90	90	12	90	90
LithGroup3	0.082	0.627		0.291		68	58	0	7,136	73	0
	Spherical					22	90	90	694	343	0
						17	328	0	22	90	90
LithGroup4	0.154	0.539		0.308		135	38	0	561	115	0
	Spherical					41	308	0	209	25	0
						28	90	90	30	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

Table 14-13     AP Correlogram Parameters

Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
MinZones	0.083	0.210	0.706	168	81	0	182	78	0
	Spherical			75	171	0	72	348	0
				12	90	90	12	90	90
Grey Schist	0.045	0.591	0.363	66	320	0	1,3704	106	0
	Spherical			21	90	90	640	16	0
				19	50	0	20	90	90
LithGroup1	0.079	0.387	0.535	57	66	0	3,322	56	0
	Spherical			14	156	0	169	326	0
				9	90	90	11	90	90
LithGroup2	0.027	0.592	0.381	61	325	0	546	52	0
	Spherical			13	90	90	333	322	0
				6	55	0	14	90	90
LithGroup3	0.109	0.462	0.429	175	85	0	11,311	95	0
	Spherical			26	355	0	674	5	0
				20	90	90	20	90	90
LithGroup4	0.034	0.188	0.778	26	74	0	203	31	0
	Spherical			26	90	90	53	121	0
				6	344	0	26	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

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Table 14-14 NP Correlogram Parameters

Domain	Nugget	S1	S2	1st Structure			2nd Structure
				Range (m)	AZ	Dip	Range (m)	AZ	Dip
MinZones	0.123	0.074	0.802	44	340	0	231	46	0
	Spherical			16	70	0	10	90	90
				10	90	90	7	136	0
LithGroup1	0.079	0.072	0.848	157	80	0	164	13	0
	Spherical			26	350	0	31	103	0
				12	90	90	12	90	90
LithGroup2	0.036	0.562	0.402	136	86	0	93	354	0
	Spherical			18	356	0	12	84	0
				3	90	90	6	90	90
LithGroup3	0.071	0.799	0.131	143	339	0	3,630	43	0
	Spherical			51	69	0	347	133	0
				6	90	90	7	90	90
LithGroup4	0.153	0.716	0.131	263	116	0	105	323	0
	Spherical			37	26	0	30	53	0
				14	90	90	15	90	90

Note: Correlograms generated from 1 m composited sample data using elevations relative to trend plane of mineralization.

14.9 Model Setup and Limits

A block model was initialized with the dimensions shown in Table 14-15. A nominal block size of 10 x 10 x 5 m is considered appropriate based on current drill hole spacing and relative to the planned scale of open pit extraction of this deposit. The limits of the block model are represented by the purple rectangles shown in the previous isometric views in Figure 14-1 and Figure 14-2.

Table 14-15     Block Model Limits

Direction	Minimum (m)	Maximum (m)	Block size (m)	Number of Blocks
X-axis (W-E)	612,190	614,100	10	191
Y-axis (N-S)	7,452,095	7,454,045	10	195
Elevation	345	1250	5	181

Using the domain wireframes, blocks in the model are assigned MinZone domain code values and the percentage of the block inside the MinZone domain is also stored—this is used to determine the proportion of in-situ resources. Blocks are defined as “overburden” if a majority (>50%) of the block occurs within the overburden domain. Similarly, blocks are defined in the Grey Schist domain on a majority basis.

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14.10 Interpolation Parameters

Grade estimates are made in model blocks using ordinary kriging (OK). The OK models were evaluated using a series of validation approaches as described in Section 14.11 of this Report. The interpolation parameters have been adjusted until the appropriate results were achieved. In general, the OK models have been generated using a relatively limited number of composited sample data. This approach reduces the amount of smoothing (also known as averaging) in the model and, while there may be some uncertainty on a localized scale, this approach produces reliable estimates of the potentially recoverable grade and tonnage for the overall deposit.

Interpolation parameters for the various items included in the resource block model are summarized in Table 14-16 through Table 14-20. Estimates for copper, lead, zinc, gold and silver are made only inside the MinZone domains as there is essentially no metals present (zero grade) in the surrounding rocks. All estimates are made using length weighted composites and model blocks are discretized into 4 x 4 x 2 points (L x W x H). Estimations for all items in the model use a dynamic search strategy where search orientations are designed to follow mineralization trend surfaces.

Table 14-16     Interpolation Parameters for Copper

MinZone Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
1, 2, 2.5, 3, 4, 5	200	200	4	3	21	7	1DH per Octant
7 & 8	200	200	10	2	21	7	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Table 14-17     Interpolation Parameters for Lead

MinZone Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
1	200	200	4	3	18	6	1DH per Octant
2, 2.5, 4	200	200	4	3	21	7	1DH per Octant
3	200	200	4	3	28	7	1DH per Octant
5	200	200	4	3	24	8	1DH per Octant
7 & 8	200	200	10	2	21	7	1DH per Octant

(1) Vertical range relative to distances from trend plane of mineralization.

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Table 14-18     Interpolation Parameters for Zinc

MinZone Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
1, 2, 2.5, 3, 4, 5	200	200	4	3	21	7	1DH per Octant
7 & 8	200	200	10	2	21	7	1DH per Octant

(1) Vertical range relative to distances from trend plane of mineralization.

Table 14-19     Interpolation Parameters for Gold

MinZone Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
1, 2, 2.5, 4, 5	200	200	4	3	21	7	1DH per Octant
3	200	200	4	3	28	7	1DH per Octant
7 & 8	200	200	10	2	21	7	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Table 14-20     Interpolation Parameters for Silver

MinZone Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
1, 2.5, 5	200	200	4	3	21	7	1DH per Octant
2, 3, 4	200	200	4	3	24	7	1DH per Octant
7 & 8	200	200	10	2	21	7	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Separate estimates for sulphur, AP, NP and SG are made for model blocks that are wholly or partially inside the MinZone domains and for blocks that are outside of the MinZone domains. Following estimation, final “whole block” values are calculated using the two estimated values and the proportion of the block inside and outside of the MinZone domains.

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Table 14-21     Interpolation Parameters for Sulphur

Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
MinZones	300	300	4	1	15	5	1DH per Octant
Grey Schist	500	500	7	1	15	5	1DH per Octant
LithGroup1-4	500	500	7	1	15	5	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Table 14-22     Interpolation Parameters for AP

Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
MinZones	500	500	5	1	15	5	1DH per Octant
Grey Schist	500	500	7	1	15	5	1DH per Octant
LithGroup1-4	500	500	7	1	15	5	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Table 14-23     Interpolation Parameters for NP

Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
MinZones	500	500	5	1	15	5	1DH per Octant
LithGroup1-4	500	500	7	1	15	5	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

Block estimates of SG were undertaken using an inverse distance (ID2) interpolation method. The parameters are listed in Table 14-24. During interpolation outside of the MinZone domains, anomalous high SG values exceeding 3.80 are restricted to a maximum distance of influence of 40 m. As stated previously, separate SG estimates are made representing areas inside the MinZone domains and for the surrounding unmineralized rocks. The final “whole block” densities are calculated using the two SG estimates and the proportion of blocks inside vs. outside of the MinZone domains.

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Table 14-24     Interpolation Parameters for Specific Gravity

Domain	Search Ellipse Range (m)			Number of Composites (1 m)			Other
	X	Y	Z(1)	Min/block	Max/block	Max/hole
MinZones	300	300	5	2	15	5
LithGroup1-4	500	500	5	2	15	5
LithGroup1-4	500	500	7	1	15	5	1DH per Octant

⁽¹⁾ Vertical range relative to distances from trend plane of mineralization.

14.11 Block Model Validation

The block models were validated using several methods: a thorough visual review of the model grades in relation to the underlying drill hole sample grades; comparisons with the change of support model; comparisons with other estimation methods; and, grade distribution comparisons using swath plots.

14.11.1 Visual Inspection

A detailed visual inspection of the block model was conducted in both section and plan to compare estimated grades against underlying sample data. This included confirmation of the proper coding of blocks within the respective domains. Examples of the distribution of copper grades in the block model are shown in cross section in Figure 14-21 and Figure 14-22.

Figure 14-21 North-South Vertical Section of Copper Estimates in the Block Model (Section 613250E) (Sim, 2017)

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Figure 14-22 West-East Vertical Section of Copper Estimates in the Block Model (Section 7453000N) (Sim, 2017)

14.11.2 Model Checks for Change of Support

The relative degree of smoothing in the block estimates was evaluated using the Hermitian Polynomial Change of Support (Herco) method, also known as the Discrete Gaussian Correction (Journel and Huijbregts, 1978). With this method, the distribution of the hypothetical block grades can be directly compared to the estimated ordinary kriging model through the use of pseudo-grade/tonnage curves. Adjustments are made to the block model interpolation parameters until an acceptable match is made with the Herco distribution. In general, the estimated model should be slightly higher in tonnage and slightly lower in grade when compared to the Herco distribution at the projected cut-off grade. These differences account for selectivity and other potential ore-handling issues which commonly occur during mining.

The Herco distribution is derived from the declustered composite grades which have been adjusted to account for the change in support moving from smaller drill hole composite samples to the larger blocks in the model. The transformation results in a less skewed distribution, but with the same mean as the original declustered samples.

Examples of Herco plots calculated for the distributions of metal in the three main MinZone domains, 1, 3 and 5, are shown in Figures 14-23 through 14-27. Note that these Change of Support calculations have been made for individual metals. Ore-waste selection will likely be made based on a NSR using all five metals. Therefore, the change of support calculations for the individual metals only serve as approximations for the distribution of NSR values above cut-off values.

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Figure 14-23 Herco and Model Grade / Tonnage Plots for Copper in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-24 Herco and Model Grade / Tonnage Plots for Lead in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-25 Herco and Model Grade / Tonnage Plots for Zinc in MinZone Domains 1, 3 and 5 (Sim, 2017)

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Figure 14-26 Herco and Model Grade / Tonnage Plots for Gold in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-27 Herco and Model Grade / Tonnage Plots for Silver in MinZone Domains 1, 3 and 5 (Sim, 2017)

Overall, the desired degree of correspondence between estimation models and change of support models has been achieved. It should be noted that the change of support model is a theoretical tool intended to direct model estimation. There is uncertainty associated with the change of support model, and its results should not be viewed as a final or correct value.

14.11.3 Comparison of Interpolation Methods

For comparison purposes, additional grade models were generated using the inverse distance weighted (ID) and nearest neighbour (NN) interpolation methods. The NN model was created using data composited to 5 m lengths to ensure all sample data are used in the model. The results of these models are compared to the OK models at various cut-off grades using a grade/tonnage graph. Figures 14-28 through 14-32 show comparison of models in the three main MinZone domains (combined 1, 3 and 5).

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There is good correlation between model types. The correspondence among the grade tonnage curves is typical for the interpolation methods being compared. The NN interpolation always has the higher grade and lower tonnage. It is an estimate that should produce a value close to the correct global mean at a zero cut-off grade. The NN grades and tonnages above cut-off are correct under the assumption that perfect selection of material above and below the cut-off can be executed at the scale of the composite samples. It is included to show the results of the averaging that takes place in the other two methods. The OK curves show the lowest grades and highest tonnages. The correct amount of averaging for the chosen block size is ensured for the OK estimate by the change of support calculation described in the preceding section.

Figure 14-28 Comparison of Copper Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-29 Comparison of Lead Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)

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Figure 14-30 Comparison of Zinc Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-31 Comparison of Gold Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)

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Figure 14-32    Comparison of Silver Model Types in MinZone Domains 1, 3 and 5 (Sim, 2017)

14.11.4 Swath Plots (Drift Analysis)

For validation of the five metals in the model, swath plots have been made for each individual MinZone domain and also a series of swaths from the three main domains (combined 1+3+5), as these contain the vast majority of the resources at the Arctic Deposit. Examples for the five metals in the deposit are shown in Figure 14-33 through Figure 14-37.

There is good correspondence between the models in most areas. The degree of smoothing in the OK model is evident in the peaks and valleys shown in the swath plots. Areas where there are large differences between the models tend to be the result of “edge” effects, where there is less available data to support a comparison. Note that the majority of the resource occurs between 7452750N and 7453450N. The validation results indicate that the OK model is a reasonable reflection of the underlying sample data.

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Figure 14-33    Swath Plot of Copper in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-34    Swath Plot of Lead in MinZone Domains 1, 3 and 5 (Sim, 2017)

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Figure 14-35    Swath Plot of Zinc in MinZone Domains 1, 3 and 5 (Sim, 2017)

Figure 14-36    Swath Plot of Gold in MinZone Domains 1, 3 and 5 (Sim, 2017)

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Figure 14-37    Swath Plot of Silver in MinZone Domains 1, 3 and 5 (Sim, 2017)

The swaths plots presented in Figure 14-38 through Figure 14-40 show the ABA items in the rocks that surround the MinZone domains.

Figure 14-38    Swath Plot of AP in Rocks Outside of the MinZone Domains (Sim, 2017)

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Figure 14-39    Swath Plot of NP in Rocks Outside of the MinZone Domains (Sim, 2017)

Figure 14-40    Swath Plot of Sulphur Rocks Outside of the MinZone Domains (Sim, 2017)

14.12 Resource Classification

The mineral resources were classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014). The classification parameters are defined relative to the distance between sample data and are intended to encompass zones of reasonably continuous mineralization that exhibit the desired degree of confidence in the estimate.

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Classification parameters are generally linked to the scale of a deposit: a large and relatively low-grade porphyry-type deposit would likely be mined at a much higher daily rate than a narrow, high-grade deposit. The scale of selectivity of these two examples differs significantly and this is reflected in the drill-hole spacing required to achieve the desired level of confidence to define a volume of material that represents, for example, a year of production. Based on engineering studies completed to date, the Arctic Deposit would likely be amenable to open pit extraction methods at a production rate of approximately 10,000 t/d. A drill hole spacing study, which tests the reliability of estimates for a given volume of material at varying drill hole spacing, suggests that drilling on a nominal 100 m grid pattern would provide annual estimates of volume (tonnage) and grade within ±15% accuracy, 90% of the time. These results were combined with grade and indicator variograms and other visual observations of the nature of the deposit in defining the criteria for mineral resource classification as described below. At this stage of exploration, there is insufficient density of drilling information to support the definition of mineral resources in the Measured category.

The following classification criteria are defined for the Arctic Deposit:

· Indicated Mineral Resources include blocks in the model with grades estimated by three or more drill holes spaced at a maximum distance of 100 m, and exhibit a relatively high degree of confidence in the grade and continuity of mineralization.

· Inferred Mineral Resources require a minimum of one drill hole within a maximum distance of 150 m and exhibit reasonable confidence in the grade and continuity of mineralization.

Some manual “smoothing” of the criteria for Indicated resources was conducted that includes areas where the drill hole spacing locally exceeds the desired grid spacing, but still retains continuity of mineralization or, conversely, excludes areas where the mineralization does not exhibit the required degree of confidence.

14.13 Mineral Resource Estimate

CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014) defines a mineral resource as:

“A mineral resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling”.

The “reasonable prospects for eventual economic extraction” requirement generally implies that quantity and grade estimates meet certain economic thresholds and that mineral resources are reported at an appropriate cut-off grade which takes into account the extraction scenarios and the processing recovery.

The Arctic Deposit comprises several zones of relatively continuous moderate- to high-grade polymetallic mineralization that extends from surface to depths of over 250 m below surface. The deposit is potentially amenable to open pit extraction methods. The “reasonable prospects for eventual economic extraction” was tested using a floating cone pit shell derived based on a series of technical and economic assumptions considered appropriate for a deposit of this type, scale and location. These parameters are summarized in Table 14-25.

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Table 14-25    Parameters Used to Generate a Resource-Limiting Pit Shell

Optimization Parameters
Open Pit Mining Cost	US$3/t
Milling Cost + G&A	US$35/t
Pit Slope	43 degrees
Copper Price	US$3.00/lb
Lead Price	US$0.90/lb
Zinc Price	US$1.00/lb
Gold Price	US$1300/oz
Silver Price	US$18/oz
Metallurgical Recovery: Copper	92%
Lead	77%
Zinc	88%
Gold	63%
Silver	56%

Note: No adjustments for mining recovery or dilution.

The pit shell has been generated about copper equivalent grades that incorporate contributions of the five different metals present in the deposit. The formula used to calculate copper equivalent grades is listed as follows:

CuEq%= (Cu% x 0.92) +(Zn% x 0.290)+(Pb% x 0.231)+(Augpt x 0.398)+(Aggpt x 0.005)

It is important to recognize that discussions regarding these surface mining parameters are used solely for the purpose of testing the “reasonable prospects for eventual economic extraction,” and do not represent an attempt to estimate mineral reserves. These preliminary evaluations are used to assist with the preparation of a Mineral Resource Statement and to select appropriate reporting assumptions.

Using the parameters defined above, a pit shell was generated about the Arctic Deposit that extends to depths approaching 300 m below surface. Table 14-26 lists the estimate of mineral resources contained within the pit shell. Based on the technical and economic factors listed in Table 14-25, a base case cut-off grade of 0.50% CuEq is considered appropriate for this deposit. The distribution of mineral resources is shown in Figure 14-41 in a series of isometric views.

There are no known factors related to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues which could materially affect the mineral resource. It is expected that a majority of Inferred resources will be converted to Indicated or Measured resources with additional exploration.

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Table 14-26    Mineral Resource Estimate for the Arctic Deposit

		Average Grade:					Contained metal:
Class	M tonnes	Cu %	Pb%	Zn%	Au g/t	Ag g/t	Cu Mlbs	Pb Mlbs	Zn Mlbs	Au koz	Ag Moz
Indicated	36.0	3.07	0.73	4.23	0.63	47.6	2,441	581	3,356	728	55
Inferred	3.5	1.71	0.60	2.72	0.36	28.7	131	47	210	40	3

Notes:

(1) Resources stated as contained within a pit shell developed using metal prices of US$3.00/lb Cu, $0.90/lb Pb, $1.00/lb Zn, $1300/oz Au and $18/oz Ag and metallurgical recoveries of 92% Cu, 77% Pb, 88% Zn, 63% Au and 56% Ag and operating costs of $3/t mining and $35/t process and G&A. The average pit slope is 43 degrees.

(2) The base case cut-off grade is 0.5% copper equivalent. CuEq = (Cu%x0.92)+(Zn%x0.290)+(Pb%x0.231)+(Augptx0.398)+(Aggptx0.005).

(3) The Mineral Resource Estimate is reported on a 100% basis without adjustments for metallurgical recoveries.

(4) The Mineral Resource Estimate is inclusive of Mineral Reserves. 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 into Mineral Reserves.

(5) Inferred resources have a great amount of uncertainty as to whether they can be mined legally or economically. It is reasonably expected that a majority of Inferred resources will be converted to Indicated resources with additional exploration.

(6) Effective date of the Mineral Resource Estimate is April 25, 2017.

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Figure 14-41    Isometric Views of Arctic Mineral Resource (Sim, 2017)

14.14 Grade Sensitivity Analysis

The sensitivity of mineral resources, contained within the resource limiting pit shell, is demonstrated by listing resources at a series of cut-off thresholds as shown in Table 14-27. The base case cut-off grade of 0.5% CuEq is highlighted in the table.

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Table 14-27    Sensitivity of Mineral Resource to Cut-off Grade

		Average Grade:					Contained metal:
Cut-off CuEq%	M tonnes	Cu %	Pb%	Zn%	Au g/t	Ag g/t	Cu Mlbs	Pb Mlbs	Zn Mlbs	Au koz	Ag Moz
Indicated
0.25	36.0	3.07	0.73	4.22	0.63	47.61	2,441	582	3,356	729	55
0.5	36.0	3.07	0.73	4.23	0.63	47.62	2,441	581	3,356	728	55
0.75	35.9	3.08	0.73	4.23	0.63	47.72	2,440	582	3,355	728	55
1	35.7	3.09	0.74	4.26	0.63	47.97	2,436	581	3,353	728	55
1.5	35.5	3.11	0.74	4.28	0.64	48.22	2,432	580	3,349	727	55
Inferred
0.25	3.8	1.58	0.56	2.52	0.34	26.76	133	47	212	42	3
0.5	3.5	1.71	0.60	2.72	0.36	28.69	131	47	210	40	3
0.75	3.0	1.93	0.65	3.04	0.36	31.99	129	44	203	35	3
1	2.5	2.29	0.73	3.52	0.37	37.04	124	39	192	29	3
1.5	2.3	2.46	0.76	3.71	0.39	39.32	122	38	184	28	3

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15.0 Mineral Reserve Estimates

15.1 Overview

Mineral Reserves were classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 10, 2014). Only Mineral Resources that were classified as Indicated were given economic attributes in the mine design and when demonstrating economic viability. Mineral Reserves for the Arctic Deposit incorporate appropriate mining dilution and mining recovery estimations for the planned open pit mining method.

The Mineral Reserve estimate for the Arctic Deposit is based on the resource block model that has an effective date of April 25, 2017, as well as information provided by Trilogy Metals, co-authors of this Report, and information generated by Amec Foster Wheeler based on the preceding PEA Study.

The following section outlines the procedures used to estimate the Mineral Reserves. The mine plan is based on the detailed mine design derived from the optimal pit shell produced by applying the Lerchs-Grossman (LG) algorithm.

15.2 Pit Optimization

The pit shells that define the ultimate pit limit, as well as the internal phases, were derived using the LG pit optimization algorithm. This process takes into account the information stored in the geological block model, the pit slope angles by geotechnical sector, the commodity prices, the mining and processing costs, the process recovery and the sales cost for the metals produced. Table 15-1 provides a summary of the primary optimization inputs.

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Table 15-1    Optimization Inputs

Parameter	Unit	Value
Metal Prices
Copper	$/lb	2.90
Lead	$/lb	0.90
Zinc	$/lb	1.10
Gold	$/oz	1,250.00
Silver	$/oz	18.00
Discount Rate	%	8
Inter-ramp Slope Angles
Sector 1 (2L-E)	degrees	45
Sector 2 (2L-W)	degrees	40
Sector 3 (2U)	degrees	45
Sector 4 (3)	degrees	30
Sector 5 (4L)	degrees	40
Sector 6 (4U)	degrees	45
Dilution	%	Estimated in a block-by-block basis
Mine Losses	%	Taken into account by block
Mining Cost
Base Elevation	m	710
Base Cost	$/t	3.00
Incremental Mining Cost
Uphill	$/t/15m	0.020
Downhill	$/t/15m	0.015
Process Costs
Operating Cost	$/t milled	19.86
G&A	$/t milled	8.92
Process Sustaining Capital	$/t milled	4.11
Road Toll Cost	$/t milled	2.66
Closure	$/t milled	1.00
Processing Rate	Kt/d	10
Process Recovery
Copper	%	90.0
Lead	%	89.9
Zinc	%	91.7
Gold	%	61.1
Silver	%	49.7
Treatment & Refining Cost	-	Variable by concentrate type/metal
Royalties
NANA Surface Use	%NSR	1.00
NANA1	%NP	0.00

1 NANA may elect to either (a) exercise a non-transferrable back-in-right to acquire between 16% and 25% (as specified by NANA) of the 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 . Upon the direction of Trilogy Metals, the PFS was evaluated based on 100% ownership by Trilogy Metals and does not include the impact on Trilogy Metals of the NANA options, either purchasing an interest in the Project or receiving a royalty payment.

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Amec Foster Wheeler imported the resource model, containing grades, block percentages, material density, slope sectors and rock types, and NSR, into the optimization software. The optimization run was carried out only using Indicated Mineral Resources to define the optimal mining limits.

The optimization run included 41 pit shells defined according to different revenue factors, where a revenue factor of 1 is the base case. To select the optimal pit shell that defines the ultimate pit limit, Amec Foster Wheeler conducted a pit-by-pit analysis to evaluate the contribution of each incremental shell to NPV, assuming a processing plant capacity of 10 kt/d and a discount rate of 8% (Figure 15-1). Following this analysis, the selected pit shell is usually smaller than the base case pit shell. The selected pit shell for the Arctic Deposit is shown in Figure 15-2. Although the NPV is slightly lower in comparison to the base case pit shell, the selected pit shell saves 72.6 Mt of waste while only losing 4.3 Mt of mineralized material.

Figure 15-1    Pit-by-Pit Analysis

Source: Amec Foster Wheeler, February 2018

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Figure 15-2    Selected Pit Shell

Source: Amec Foster Wheeler, February 2018

15.3 Dilution and Ore Losses

The mineral resources for the Arctic Deposit were reported undiluted. To determine mineral reserves, dilution was applied to the resource model in two steps: planned dilution and contact dilution. These procedures include ore losses.

Planned dilution was estimated as follows:

· The resource model was re-blocked from a block size of 10x10x5 m to 5x5x2.5 m to better approximate mining selectivity.

· The diluted grades of each block were calculated using the formula:

Contact dilution was estimated as follows:

· The grade of a given block will be diluted by blending 20% of the tonnage from each of the four adjacent blocks.

· If an adjacent block is classified as Inferred Mineral Resource, its grade is considered to be zero. If the adjacent block is classified as Indicated, but below cut-off, dilution is taken at the grade of the adjacent block.

· The procedure is illustrated in Figure 15-3.

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Figure 15-3 Contact Dilution Estimation Procedure

15.4 Mineral Reserve Statement

As the mining cost varies with depth individual blocks captured within the final pit design were tagged as either ore or waste by applying the parameters shown in Table 15-1. Using the partial block percentages within the final pit design the ore tonnage and average grades were calculated. The Mineral Reserves statement is shown in Table 15-2.

Table 15-2    Mineral Reserve Statement

	Tonnage	Grades
Class	t x 1000	Cu (%)	Zn (%)	Pb (%)	Au (g/t)	Ag (g/t)
Proven Mineral Reserves	-	-	-	-	-	-
Probable Mineral Reserves	43,038	2.32	3.24	0.57	0.49	36.0
Proven & Probable Mineral Reserves	43,038	2.32	3.24	0.57	0.49	36.0
Waste within Designed Pit	296,444
Total Tonnage within Designed Pit	339,482

Notes:

(1) Mineral Reserves are estimated assuming open pit mining methods and include a combination of planned and contact dilution.

(2) Mineral Reserves are based on prices of $2.90/lb Cu, $0.90/lb Pb, $1.10/lb Zn, $1250/oz Au and $18/oz Ag

Fixed process recoveries of 90.0% Cu, 89.9% Pb, 91.7% Zn, 61.1% Au and 49.7% Ag

(3) Mining costs: $3.00/t incremented at $0.02/t/15 m and $0.015/t/15 m below and above 710 m elevation respectively.

(4) Processing costs: $36.55/t. Include process cost: $19.86/t, G&A: $8.92/t, sustaining capital: $4.11/t closure cost: $1.00/t, and road toll: $2.66/t.

(5) Treatment costs of $70/t Cu concentrate, $180/t Pb concentrate and $300/t Zn concentrate. Refining costs of $0.07/lb Cu, $10/oz Au, $0.60/oz Ag. Transport cost of $149.96/t concentrate.

(6) Fixed royalty percentage of 1%.

(7) The Qualified Person for the Mineral Reserves is Antonio Peralta Romero P.Eng., an Amec Foster Wheeler employee who visited the project site in July 2017 as part of the data verification process.

(8) The effective date of mineral reserves estimate is October 10, 2017.

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15.5 Factors Affecting Mineral Reserves

The Arctic Mineral Reserves are subject to the types of risks common to open pit polymetallic mining operations that exist in Alaska. Those risks are generally well understood at the Pre-Feasibility level of study, and should be manageable.

There is currently no developed surface access to the Arctic Project area and beyond. Access to the Arctic Project is proposed to be via the Ambler Mining District Industrial Access Project (AMDIAP), a road approximately 340 km (211 miles) long, extending west from the Dalton Highway where it would connect with the proposed Arctic Project area. The final terminal for the road has not yet been determined. The capital costs of the road are not yet final; however, an estimate of approximately $300 million has been used in this PFS. Although Trilogy Metals has been in discussions with the Alaska Industrial Development & Export Authority (AIDEA) who had been investigating alternatives to reduce the cost to construct the AMDIAP, the final cost of the road could be higher than $300 million. The working assumption for this PFS 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 the Arctic Project) in order to pay back the costs of financing the AMDIAP. 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 and other users of the road which could also include commercial transportation of materials and consumer items that would use the AMDIAP to ship concentrates to the Port of Anchorage in Alaska and possibly provide goods and commercial materials to villages in the region.

For the purposes of this PFS, AIDEA and Trilogy Metals reviewed the current bonding ability of AIDEA based on a $300 million 30-year bond rate of 6% compounded semi-annually and a $300 million 15-year bond rate of 5.50% compounded semi-annually. Although the final toll payments will be negotiated with AIDEA and the Public-Private Partnership owners of the access road sometime in the future, it has been assumed that a toll would be paid based on the Arctic Project paying approximately $9.7 million each year for its 12-year mine life. Based on this, for Mineral Reserves determination a road cost of $2.66/t processed was assumed.

There is a risk to the Mineral Reserves if the toll road is not built in the time frame required for the Arctic Project, or if the toll charges are significantly different from what was assumed.

Proper management of groundwater will be important to maintaining pit slope stability, as well as additional pit slope design verification work to advance the project to the feasibility level. The geotechnical report authored by SRK (2017) mentions that the stability of the east wall is highly sensitive to a number of design parameters. The geotechnical assumptions used in the pit design may vary in future assessments and could materially affect the strip ratio, or mine access design.

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Additionally, the presence of talc layers in the rock has not been included in the geological model and their presence could affect recoveries in the process plant and therefore could be a risk to the mineral reserves. Trilogy Metals is aware of this risk and has considered the inclusion of a talc recovery circuit to mitigate this risk.

The assumed metallurgical recoveries may be different to the actual recoveries achieved during the production. 

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16.0 Mining Methods

16.1 Overview

The following section outlines the parameters and procedures used to design the mine as a conventional open pit, estimate the Mineral Reserves within the open pit mine plan, and establish a practical mining schedule for the Arctic Pre-feasibility Study (PFS). The mine plan is based on the Probable Mineral Reserves discussed in Section 15.0 of this Report.

16.2 Mine Design

The mine is designed as a conventional truck-shovel operation assuming 131 t trucks for waste and 91 t trucks for ore, as well as 17 m³ and 12 m³ shovels for waste and ore respectively. The pit design includes three nested phases to balance stripping requirements while satisfying the concentrator requirements.

The design parameters include a ramp width of 28.5 m, in-pit road grades of 8% and out-pit road grades of 10%, bench height of 5 m, targeted mining width between 70 and 100 m, berm interval of 15 m, variable slope angles by sector and a minimum mining width of 30 m. Table 16-1 shows the mine design parameters.

Table 16-1    Mine Design Parameters

				Geotechnical Sector
Parameter	Units	2L-E	2L-W	2U	3	4L	4U
Inter-Ramp Angle	degrees	45	40	45	30	40	45
Bench Face Angle	degrees	60	65	65	60	65	65
Bench Height	m	5	5	5	5	5	5
Catch Bench Spacing	bench	3	3	3	3	3	3
Target Road Gradient	%	8	8	8	8	8	8
Maximum Road Gradient	%	10	10	10	10	10	10
Road Width - Two Lanes	m	28.5	28.5	28.5	28.5	28.5	28.5
Road Width - One Lane	m	18.5	18.5	18.5	18.5	18.5	18.5

The smoothed final pit design contains approximately 43 Mt of ore and 296 Mt of waste for a resulting stripping ratio of 6.9:1. Within the 43 Mt of ore, the average grades are 2.32% Cu, 3.24% Zn, 0.57 % Pb, 0.49 g/t Au and 36.0 g/t Ag. Figure 16-1 shows the ultimate pit design. Figure 16-2 and Figure 16-3 show pit sections comparing the mine design to the selected pit shell.

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Figure 16-1    Ultimate Pit Design

Source: Amec Foster Wheeler, February 2018

Figure 16-2    Section 1 Showing Mine Design and Selected Pit Shell (looking West)

Source: Amec Foster Wheeler, February 2018

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Figure 16-3    Section 2 Showing Mine Design and Selected Pit Shell (looking North-West)

Source: Amec Foster Wheeler, February 2018

16.3 Waste Rock Facilities and Stockpile Designs

The design and construction of the waste rock dump (WRD) should ensure physical and chemical stability during and after mining activities. To achieve this, the waste areas and stockpiles are designed to account for benching, drainage, geotechnical stability, and concurrent reclamation.

The WRD design criteria include 23.5m benches every two lifts, 2.5H:1V overall slopes, 10 m lifts, and a 33% swell factor for estimating volumes. The overburden mined represent approximately 6% of the total waste and it is encapsulated within the waste rock. Figure 16-4 shows the WRD design.

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Figure 16-4    Waste Rock Dump

Source: Amec Foster Wheeler, February 2018

A small stockpile is required to store the ore mined during the pre-production period. Because this stockpile will be depleted at the beginning of the operation, it is located within the WRD footprint and have a total storage capacity of 260,000 m³. This volume is enough to satisfy the maximum stockpiling capacity of approximately 627 kt. Figure 16-5 shows the stockpile design with respect to the WRD at the end of the pre-production period.

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Figure 16-5    Ore Stockpile

Source: Amec Foster Wheeler, February 2018

16.4 Production Schedule

The production schedule includes the processing ramp-up. The processing plant ramp-up takes into account the inefficiencies related to the start of operations, and includes the tonnage processed as well as the associated recoveries, which steadily increase to reach the design capacity during the third quarter of operation. The mine will require two years of pre-production before the start of operations in the processing plant.

The deposit will be mined in three nested phases, including the ultimate pit limit. The schedule was developed in quarters for the pre-production period and for the first two years of production, then in yearly periods for the remainder of the life of mine (LOM). The scheduling constraints set the maximum mining capacity at 32 Mt per year and the maximum number of benches mined per year at eight in each phase.

The schedule produced based on the Probable Mineral Reserves shows a LOM of 12 years. The amount of rehandled mill feed is 627 kt, which is the ore mined during the pre-production period. The average grades to the mil over the LOM are 2.32% Cu, 3.24% Zn, 0.57 % Pb, 0.49 g/t Au and 36.0 g/t Ag. The yearly LOM schedule is shown in Table 16-2 and Figure 16-6. Figure 16-7 shows the scheduled copper feed grade and Figure 16-8 shows the stockpile balance.

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Table 16-2    Production Schedule

	Tonnage (kt)					Feed Grades
	To Mill			Mine to	Total	Cu	Zn	Pb	Au	Ag
Period	Direct	Stockpile	Total	Stockpile	Waste	%	%	%	g/t	g/t
Year -2	-	-	-	188	22,902	-	-	-	-	-
Year -1	-	-	-	439	31,564	-	-	-	-	-
Year 1	2,469	627	3,096	-	29,588	2.61	3.19	0.53	0.43	37.3
Year 2	3,649	-	3,649	-	28,327	2.30	3.32	0.62	0.53	37.0
Year 3	3,651	-	3,651	-	28,447	2.18	3.47	0.63	0.49	33.6
Year 4	3,650	-	3,650	-	28,833	2.19	3.24	0.62	0.50	32.3
Year 5	3,650	-	3,650	-	26,230	2.05	2.57	0.40	0.31	25.4
Year 6	3,651	-	3,651	-	24,646	2.21	2.65	0.44	0.37	29.1
Year 7	3,651	-	3,651	-	20,948	2.35	2.81	0.47	0.48	34.2
Year 8	3,650	-	3,650	-	16,928	2.32	3.58	0.60	0.52	39.0
Year 9	3,651	-	3,651	-	13,812	2.66	3.96	0.65	0.57	45.3
Year 10	3,650	-	3,650	-	12,226	2.45	3.29	0.54	0.56	39.6
Year 11	3,650	-	3,650	-	8,088	2.34	3.49	0.58	0.53	38.7
Year 12	3,438	-	3,438	-	3,906	2.17	3.37	0.70	0.55	40.9
Total	42,411	627	43,038	627	296,444	2.32	3.24	0.57	0.49	36.0

Figure 16-6    Production Schedule

Source: Amec Foster Wheeler, February 2018

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Figure 16-7    Scheduled Cu Feed Grade

Source: Amec Foster Wheeler, February 2018

Figure 16-8    Stockpile Balance

Source: Amec Foster Wheeler, February 2018

16.5 Waste Material Handling

Waste will be hauled to the WRD using 131 t trucks. The construction sequence starts at the bottom of the dump by dumping the material in 10-m lifts, leaving a 23.5 m berm every two lifts. The resulting overall slope angle of the dump face will be 2.5H:1V.

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16.6 Operating Schedule

The Arctic mine is scheduled to operate 24 hours a day, seven days a week using three rotating crews working 12-hour shifts. During the day, there are two 12 hour shifts scheduled, consisting of a day shift and a night shift. The crews “hot change” or overlap between shifts to allow for continuous mine operations. A number of duties only require work during the daylight hours. For these duties, two crews rotate on a two-week-on-one-week-off schedule to provide coverage.

For the rotating mine operations crews, approximately 3.5 hours are lost per day to standby time, inclusive of two hours for breaks, 30 minutes for fueling, 20 minutes for shift change, 20 minutes for blast delay, and 20 minutes for meetings (Table 16-3).

Over a year, approximately 10 days or 240 hours are assumed lost to poor weather conditions, predominantly in the winter time. It is assumed that the equipment is manned but delayed during these weather events.

Table 16-3    Gross Operating Hours per Year

Calendar Time
	Days	365
	Shifts per day	2
	Shift length (h)	12
	Calendar Time (h/year)	8,760
Available Time
	Availability	85%
	Down time (h/year)	1,314
	Available Time (h/year)	7,446
Gross Operating Time
	Operating Standby
	Internal (minutes/day)
	Lunch & breaks	120
	Blast delay	20
	Fueling	30
	Shift change	20
	Meetings	20
	External – Weather (h/year)	240
	Operating Standby (h/year)	1,290
	Gross Operating Hours (h/year)	6,156

Accounting for standby time and weather delays, equipment accumulates approximately 6,156 gross operating hours (GOH) per year in the example above. For productivity calculations, it is assumed that following preproduction, the trucks and shovels are in a productive cycle approximately 50 minutes each hour, or 83% of the time. For drills, the productive utilization is lower and in the range of 65% based on benchmarking with an area mine. During the preproduction period, the equipment’s productive utilization is de-rated to account for initial site conditions and operator skill level (Table 16-4).

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Table 16-4    Productive Utilization Ramp-up

Period	Productive Utilization
PP Q8	67%
PP Q7	67%
PP Q6	75%
PP Q5	75%
PP Q4	75%
PP Q3	75%
PP Q2	75%
PP Q1	75%
Yr1 Q1 plus	83%

As with mine operations, mine maintenance is scheduled to work a 24/7 schedule to allow for continuous maintenance coverage. However, the majority of planned maintenance work is done during the day shift with a skeleton crew scheduled for the night shift.

Blasting is only scheduled during the daylight hours. Two blasting crews rotate on a 12 hour day shift, for seven day-a-week coverage.

16.7 Mining Equipment

The Arctic Deposit will be mined using a conventional owner-operated truck fleet loaded by a combination of hydraulic shovels/excavators and front-end loaders (FELs). The truck fleet will consist of larger trucks for waste stripping and smaller trucks for mining the ore zones. Both types of truck will be diesel powered with a combined capacity to mine approximately 32.0 Mt per year operating on a combination of 5 m and 10 m benches. The loading fleet will also be diesel powered. Blasting will be contract performed with an emulsion product supplied from Fairbanks, Alaska.

Equipment requirements are estimated quarterly during preproduction and the first two years of mining, and annually thereafter. Equipment sizing and numbers are based on the mine plan, the operational factors shown in Table 16-5, and a 24 hour per day, seven day a week work schedule.

Table 16-5    Availability and Productive Utilization Post Ramp-up

Equipment	Availability	Productive Utilization
178 mm/45klb Production Drill	82%	65%
5 inch Top head hammer track drill	85%	65%
300 t/17 m3 Hydraulic Face Shovel	86%	83%
230 t/12 m3 Hydraulic Excavator	86%	83%
125 t/12 m3 Front End Loader	86%	83%
131 t Haul Truck	85%	83%
91 t Haul Truck	85%	83%
40 t Articulated Truck	85%	83%
68 t/4 m3 Hydraulic Excavator	82%	83%
35 t m3 Hydraulic Excavator	85%	83%
70 t/430 kW Track Dozer	85%	83%
50 t/370 kW Rubber Tired Dozer	85%	83%
40 t Articulated Sand Truck	85%	83%
27 t/221 kW Motor Grader	85%	83%
34,000 L water truck	85%	83%
40 t Articulated Fuel/Lube Truck	85%	83%

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16.7.1 Blasting

Blasting operations will be contracted to a blasting explosives provider who is responsible for shot design, loading, stemming, and initiation. The explosive provider will transport explosive product from Fairbanks, and store on site in two explosive silos each with 10 days of product capacity. One silo will be for storing Ammonium Nitrate (AN) and the other will be for storing emulsion.

In performing the explosive services, the blasting contractor is proposing to provide:

· Two three person blasting crews providing seven day a week coverage

· 2 MMU bulk explosive trucks

· 1 one ton blast truck

· 1 supervisor truck

The MMU explosive trucks will deliver a bulk emulsion product down the borehole that has a specific gravity of 1.2 g/cm³. Blasting quantities were estimated based on the 1.2 g/ cm³ explosive density and the blast design criteria provided in Table 16-6.

Table 16-6    Blasting Design Input

Description	Drill Type	Material Type	Rock UCS (MPa)	Rock Density (t/m3)	Bench Height (m)	Powder Factor (kg/t)
10m Bench - Waste	MD6250	Waste	65	2.8	10	0.25
10m Bench - Ore	MD6250	Ore	88	3.2	10	0.37

Based on benchmarking, a powder factor of 0.37 kg/t will be used for ore, and a powder factor of 0.25 kg/t will be used for waste. Although material will be mined on a combination of 5 and 10 m benches, all material will be drilled and blasted on 10 m benches. Summary blast designs are shown in Table 16-7.

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Table 16-7    Blast Designs

Material	Bench Height (m)	Powder Factor (kg/t)	Bit Size (mm)	Burden (m)	Spacing (m)	Stemming (m)	Sub Drill (m)
Waste	10	0.25	178	5.2	6.0	4.2	1.6
Ore	10	0.37	178	4.1	4.8	3.3	1.2

16.7.2 Drilling

Throughout the mine life, drilling is required for both ore control and blasting. Rock fragmentation achieved through blasting is the overriding design criteria for the drill hole pattern design. The drill pattern outlined in Table 16-7 along with the drill penetration rates described below are used to estimate drilling requirements.

Drill penetration is a function of bit size, bit load, drilling method, and rock strength properties. Amec Foster Wheeler relied on SRK’s 2017 Pre-Feasibility Slope Geotechnical and Hydrogeological report for rock strength properties. SRK completed unconfined compressive testing and point load testing on the primary rock types. The results of the uniaxial compressive tests for the primary ore hosting rocks (lithological units) are shown in Table 16-8. Amec Foster Wheeler calculated a weighted average uniaxial compressive strength (UCS) value for the ore hosting rocks of 88 Mpa. Table 16-8 also shows the representative lithologies used to calculate a weighted average UCS value of 65 Mpa for the waste areas.

Table 16-8    Rock Type Weight and UCS

Lithology	UCS (Mpa)	% of total	% of Ore Hosting
Ore Hosting
GS	97	11%	75%
MS	56	2%	13%
SMS	68	2%	12%
Average Ore	88	15%	100%
Waste
CHS	20	4%	6%
DM	71	1%	2%
MRP	73	16%	24%
QFMS	60	19%	27%
GS	97	11%	16%
MS	56	2%	3%
SMS	68	2%	3%
QMS	50	8%	12%
QMSp	62	3%	5%
TS	17	1%	2%
Average Waste	65	68%	100%

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According to the Workman Calder Rock classification, the rock types are rated as Moderate with a Rock Penetration Factor (RFI) of 100 for waste and 95 for ore respectively. Table 16-9 shows the calculated penetration rates using the Workman and Calder equation.

Table 16-9    PV271 Drill Penetration Rates

Material Type	Inst. Pen. rate	Tram Time	Setup Time	Sampling Time	Total Cycle	Average
	m/h	min	min	min	min	Pen Rate1 (m/h)
Ore	42.8	2.25	1.00	0.75	34.8	22.3
Waste	37.7	2.25	1.00	0.75	30.8	20.0

¹Assumes 65% efficiency

Table 16-10 shows the drill requirements, the metres drilled, the hours operated, and the average penetration rate by period. By the second year of preproduction, mining requires four production drills. Following Year 6, drill requirements drop to three along with a drop in total tonnes mined. Metres drilled include a 5% allowance for additional trim drilling and re-drills. Penetration rates average 21.7 over the LOM.

Table 16-10    Drill Requirements and Performance

Period	Drills Required	Meters Drilled	Operating Hours	Avg. Pen Rate
	#	(m)	(h)	(m/h)
PP -2	3	324,665	14,931	21.7
PP -1	4	450,776	20,719	21.8
Yr1	4	460,555	21,118	21.8
Yr2	4	464,653	21,271	21.8
Yr3	4	466,368	21,470	21.7
Yr4	4	471,763	21,529	21.9
Yr5	4	435,258	19,966	21.8
Yr6	4	413,065	18,948	21.8
Yr7	3	361,183	16,568	21.8
Yr8	3	304,794	13,981	21.8
Yr9	3	261,111	11,978	21.8
Yr10	2	238,861	11,000	21.7
Yr11	2	180,820	8,544	21.2
Yr12	1	118,258	5,758	20.5
Total		4,952,130	227,781	21.7

In addition to the production drills, one top head hammer (THH) drill with a 4 ½ inch (114 mm) bit is used for pre-split drilling. The THH can also be used for pioneer mining/bench development and road construction.

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16.7.3 Loading

From the total cost of ownership (TCO) analysis, the primary loading unit selected for waste stripping is a CAT 6030 hydraulic shovel (6030). The primary loading unit selected for ore mining is a CAT 6020 hydraulic excavator (6020). To assist the hydraulic shovels, two CAT 993K high lift FELs are scheduled throughout the mine life. The CAT 993K FEL will also be used for stockpile rehandle, all of which is schedule in Year 1.

The CAT 6030 shovel will four-pass load the CAT 785 truck in approximately two minutes. The LOM production rate will average 1,880 tdry/GOH. The peak productivity scheduled for the CAT 6030 occurs in Years 1, when it is scheduled at 2,099 tdry/GOH.

The CAT 6020 shovel will four-pass load the CAT 777 truck in approximately two minutes. The LOM production rate will average 1,145 tdry/GOH. The peak productivity scheduled for the CAT 6020 occurs in Years 9, when it is scheduled at 1,399 tdry/GOH.

The CAT 993K high lift FEL will four-pass load the CAT 777 truck in approximately two minutes and 34 seconds, and six-pass load the CAT 785 truck in approximately three minutes and 50 seconds. The LOM production rate will average 844 tdry/GOH. The peak productivity scheduled for the FELs occurs in Year 7, when the CAT 993K loaders are scheduled at 1,286 tdry/GOH.

In addition to the primary loading units, a CAT 374 excavator will be paired with two CAT 745 trucks throughout the mine life for bench development/pioneering work and winter snow removal. It will also be used to maintain haul roads and scale the pit walls as needed.

16.7.4 Hauling

From the TCO analysis, the primary hauling unit selected for waste stripping is the CAT 785 mechanical drive truck; and for ore stripping the primary hauling unit selected is the CAT 777 mechanical drive truck. The CAT 785 and CAT 777 have a payload capacity of 131 t wet and 91 t wet respectively, assuming a standard body with a full set of liners. The dry capacity is estimated at 127 and 88 t for the CAT 785 and CAT 777 trucks respectively, assuming 3% moisture and carry back.

Amec Foster Wheeler estimated truck requirements on a period-by-period basis using travel distances from a road network developed within Minesight®. Haul segment distances were reported for each material type from their location on a mining bench to their final destination. Assuming 2% rolling resistance for haul roads, travel speeds were estimated from the manufacture’s performance curves, and applied to each haul segment to estimate travel time.

Truck requirements by period are shown in Table 16-11 for the CAT 785 and CAT 777 trucks, together with the average one-way haul distance, average fuel consumption, and average truck productivity. Due to the Phase I initial stripping campaign, the CAT 785 waste mining trucks will be brought on early. Ten trucks will be commissioned during pre-production. CAT 785 requirements remain at 10 until Year 6 when they are reduced to five before being retired the following year.

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The CAT 777 truck requirements will start at four during pre-production. These trucks will be focused on mining the Phase II ore layback. Over the next two years, the CAT 777 fleet will be ramped up to 11. The CAT 777 fleet reaches its peak in Year 4 at 13, before dropping to 12 in Year 8. Following Year 8, the CAT 777 truck requirements will decline with mining production.

Table 16-11    Truck Requirements & Productivity Statistics

	CAT 785				CAT 777
Period	Trucks	Avg one- way	Avg	Avg Truck	Trucks	Avg one- way	Avg	Avg Truck
	Required	Haul	Fuel Burn	Production	Required	Haul	Fuel Burn	Production
	#	(m)	l/GOH	t/GOH	#	(m)	l/GOH	t/GOH
PP -2	10	2,939	70	301	4	2,939	45	166
PP -1	10	2,424	72	347	7	2,424	59	242
Yr1	10	3,417	98	294	11	3,342	70	214
Yr2	10	3,654	98	293	11	3,788	69	204
Yr3	10	2,223	90	283	12	3,905	77	197
Yr4	10	3,882	114	275	13	4,429	82	192
Yr5	10	3,177	102	254	13	3,177	80	177
Yr6	5	2,296	96	326	13	2,296	76	227
Yr7	-	-	-	-	13	2,001	76	306
Yr8	-	-	-	-	12	2,224	80	277
Yr9	-	-	-	-	11	2,470	83	257
Yr10	-	-	-	-	11	2,746	86	234
Yr11	-	-	-	-	7	2,234	82	271
Yr12	-	-	-	-	5	2,407	83	237
Total		3,049	92	295		2,940	76	230

16.7.5 Support

Support equipment will include excavators, track dozers, rubber-tired dozers (RTDs), sand trucks, graders, water trucks, fuel/lube trucks, and water trucks. The major tasks for the support equipment include:

· Bench and road maintenance

· Shovel support/clean-up

· Blasting support/clean-up

· WRD maintenance

· Stockpile construction/maintenance

· Road building/maintenance

· Pioneering and clearing work

· Field equipment servicing.

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A description of each support equipment fleet follows:

· 35 t Excavator (CAT 330) – One CAT 330 excavator is scheduled throughout the mine life. Its primary functions will be to support dewatering, maintain pit drainage, break rocks using the rock breaker attachment, and backup the larger CAT 374 excavator.

· 443 kW track dozers (CAT D10) are estimated at 0.5 dozers per production blast hole drill and production loading unit. The CAT D10 dozer fleet will peak at four machines in pre-production. Their primary functions will be to maintain pit floors, maintain dumps and stockpiles, build pit roads, and clean final pit walls. Due to limited mobility, the 71 t dozers will be transported between working areas using a 90-t capacity transport trailer. The transport will also be used to transport the 65 t drills and the 35 t CAT 374 excavator.

· 419 kW rubber-tired dozer (CAT 834K) requirements are estimated at approximately one RTD per hydraulic shovel. Their primary function will be to maintain shovel floors, provide drill pattern clean-up, clear rock spillage, and provide backup dump and stockpile maintenance. At peak, two CAT 834Ks will support one CAT 6030 shovel and associated mining areas.

· CAT 745 Sand Trucks – One CAT 745 truck will be fitted with a sander and used to support winter operations throughout the LOM.

· 221 kW motor graders (CAT 16M) are estimated at approximately one grader per eight trucks. Their primary function will be to maintain roads, dump areas, and pit areas. The peak 16M grader fleet of three graders will support a peak truck fleet of 10 CAT 785 trucks and 13 CAT 777 trucks.

· CAT 777 34,000 L water truck requirements are estimated at a ratio of one water truck per 12 haul trucks. During the winter season, from October to April, water trucks will be lightly scheduled. They will be primarily used for watering the drills and for fire patrol; nonetheless, even during the winter season roads become dusty. During May to September, when dust suppression requirements are at their highest, the water trucks will be fully scheduled. At peak, the mine will operate two water trucks.

· CAT 745 fuel/lube truck requirements are estimated at a ratio of one lube truck per 10 pieces of tracked field equipment. They will be used to fuel and service shovels and other tracked field equipment.

Support equipment requirements are shown in Table 16-12.

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Table 16-12    Support Equipment

Period	CAT 330 Excavator	CAT D10 Dozer	CAT 834 RTD	CAT 745 Sand Truck	CAT 16 Grader	CAT 777 Water Truck	CAT 745 Fuel/Lube Truck
PP -2	1	1	2	1	2	2	2
PP -1	1	4	2	1	2	2	2
Yr1	1	4	2	1	3	2	2
Yr2	1	4	2	1	3	2	2
Yr3	1	4	2	1	3	2	2
Yr4	1	4	2	1	3	2	2
Yr5	1	4	2	1	3	2	2
Yr6	1	4	2	1	2	2	2
Yr7	1	3	1	1	2	2	1
Yr8	1	3	1	1	1	1	1
Yr9	1	3	1	1	1	1	1
Yr10	1	3	1	1	1	1	1
Yr11	1	2	1	1	1	1	1
Yr12	1	2	1	1	1	1	1

16.7.6 Auxiliary

To support mine maintenance and mine operation activities, a fleet of auxiliary equipment is required. The types and numbers of auxiliary equipment are listed in Table 16-13 in five-year increments.

Table 16-13    Auxiliary Equipment

Equipment	Year 1	Year 5	Year 10
Mine Maintenance
Truck Mounted 40 t Crane	1	1	1
80 t Rough Terrain Crane	1	1	1
5 t Rough Terrain Forklift	2	1	2
10t Forklift	1	1	1
Mechanic Service Truck	3	3	2
Small Fuel/Lube truck	1	1	1
CAT 262 Skid Steer	1	1	1
Flatbed Truck	1	1	1
CAT TL1055 Telehandler	1	1	1
Mine Operations
CAT 450E backhoe/loader	1	1	1
CAT H180DS hydraulic hammer/impactor	1	1	1
90t Lowboy	1	1	1
Compactor	1	1	1
Light Plant	12	12	8
Transport Tractor	1	1	1
Tire Handler Truck	1	1	1
3/4 ton Pickup	13	8	8
1 ton Pickup	7	7	4
Crew Bus	4	4	3
MineStar	1	1	1
Mine & Geology Software	1	1	1
Pumps	2	2	2

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16.8 Open Pit Water Management

Pit water requiring management will be a combination of groundwater inflow, direct precipitation and runoff from areas surrounding the pit. Total average inflow is estimated at 3,800 m³/d. Pit water will be managed via in-pit sumps.

Measures for slope pore pressure control are not considered necessary at this time, but monitoring should be conducted to verify assumptions.

Further detail on pit hydrogeology was provided in Section 9.7.1.

16.9 Geotechnical Review

Amec Foster Wheeler carried out a high-level review of the geotechnical report authored by SRK (2017), which was provided as the basis to support the present PFS mine design. This report includes a hydrogeological assessment and the assessment of the overall pit slope stability.

Amec Foster Wheeler concurs with the recommendations provided by SRK for the open pit mine design at the PFS level. However, additional design verification work will be needed to advance the project to the feasibility level. The east wall stability is highly sensitive to a number of design parameters. A second factor that will affect stability is expected presence of talc layers that are not included in the current geological model.

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17.0 Recovery methods

17.1 Mineral Processing

A 10,000 t/d process plant has been designed to process the massive and semi-massive sulphide mineralization at Arctic. 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.

Run-of-mine (ROM) material will be hauled from the open pit to a primary crushing facility where it will be crushed by a jaw crusher. The crusher discharge will have a particle size of 80% passing 125 mm and will be fed to the grinding and flotation circuit. The crushed material will be ground in two stages of grinding, including a semi-autogenous grind (SAG) mill and a ball mill. The ball mill is operated in closed circuit with classifying cyclones (SAB circuit). The cyclone overflow will have a grind size of approximately 80% passing 70 μm. The mined materials will contain significant levels of talc and will first undergo talc pre-flotation to remove this mineral prior to traditional base metal flotation of three payable concentrates, copper, zinc, and lead concentrates. The final tailings from the zinc flotation circuit will be pumped to the tailings management facility (TMF). Copper, zinc, and lead concentrates will be thickened and pressure-filtered before being transported off site and shipped to market. A process flowsheet is shown in Figure 17-1.

The average annual dry concentrate production is estimated as follows:

·	Copper concentrate:	246,723 t/a
·	Lead concentrate:	29,493 t/a
·	Zinc concentrate:	180,219 t/a

17.1.1 Flowsheet Development

The process plant will consist of the following unit operations:

· Crushing:

o primary crushing by a jaw crusher

o associated conveying and dust suppression systems

· Grinding and flotation:

o primary grinding by a SAG mill and secondary grinding by a ball mill

o talc pre-flotation

o bulk copper and lead rougher/cleaner flotation

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o bulk copper lead concentrate regrinding and cleaner flotation

o copper and lead separation flotation

o zinc regrinding and rougher/cleaner flotation

o concentrate dewatering and concentrate load out,

o tailings disposal to the TMF.

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Figure 17-1    Simplified Process Flowsheet

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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 design criteria developed for the processing facilities are outlined in Table 17.1.

Table 17-1    Processing Facility Design Criteria

Design Criteria	Unit	Value
Daily Processing Rate	t/d	10,000
Operating Days per Year	d/a	365
Operating Schedule		two shifts/day; 12 hours/shift
Crushing Availability	%	65
Grinding/Flotation Availability	%	92
Abrasion Index	g	0.032
Bond Ball Mill Work Index	kWh/t	9.84
Crushing
Nominal Processing Rate	t/h	641
Crusher Feed Particle Size	mm	less than 1,000
Primary Crushing Product Particle Size, 80% Passing	mm	125
Grinding
Nominal Processing Rate	t/h	453
Secondary Grind Size, 80% Passing	µm	70
Grinding Recirculating Load	%	300

17.1.2 Process Plant Description

Crushing Plant

ROM material will be trucked to the crushing station where a 200 t receiving dump bin equipped with a stationary grizzly screen will feed the jaw crusher. The primary jaw crusher will be driven by a 200 kW motor. A rock breaker will be provided to break oversize rocks retained on the grizzly screen located above the crusher feed bin. The designed nominal crushing rate is 641 t/h at 65% availability.

The jaw crusher discharge product is expected to be 80% passing 125 mm. The crusher product will be conveyed to a 5,000 t live capacity stockpile, providing approximately 12 hours of live storage for the grinding/flotation plant.

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, and protect the downstream conveyor.

The main equipment in the crushing area will include:

· One 1,000 mm stationary grizzly

· One hydraulic rock breaker

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· One primary apron feeder, 1,600 mm wide by 6,100 mm long

· One jaw crusher, 1,070 mm by 1,400 mm; driven by a 200 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 5,000 t equivalent to approximately 12h of mill feed at the nominal mill feed rate, in addition to a dead capacity of approximately 20,000 t. The coarse material stockpile will be a covered facility to effectively control dust losses 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 151 t/h per feeder at 92% availability. Reclaimed material from the apron feeders will be discharged onto a 900 mm wide by 135m long SAG mill feed conveyor.

17.1.4 Grinding and Classification

Primary Grinding And Classification

Crushed material will be reclaimed from the coarse ore storage pile and conveyed from the stockpile to the concentrator at a rate of 453 t/h. The primary grinding circuit will consist of one 7.3 m diameter by 6.4 m long variable speed SAG mill equipped with a trommel screen. The installed power on the SAG mill will be 4,100 kW provided by a single motor and variable speed drive.

Discharge from the SAG mill will be screened and screen undersize material will flow by gravity to a pump box which feeds the classifying cyclones. Oversized material from the SAG mill screen will be conveyed back to SAG mill for additional grinding. The circulating load of oversize from the SAG mill discharge screen is estimated at 25% of SAG mill feed. As required, steel balls will be added to the SAG mill to maintain mill power.

Secondary Grinding And Classification

The secondary grinding circuit will include a single ball mill operated in closed circuit with a classifying cyclone cluster. The ball mill will be a 5.3 m diameter by 9.0 m effective grinding length (EGL) ball mill, powered by 4,100 kW motor. The SAG mill screen undersize will be combined with ball mill discharge to feed the classifying cyclone cluster. The cyclone underflow will gravity-flow to the ball mill, while the cyclone overflow, with a solid content of 33%, will be sent to the flotation plant. The designed circulation load for the ball mill is approximately 300%. The flotation feed slurry is estimated to have a particle size of 80% passing 70 µm.

As required, steel balls will be added into the ball mill to maintain the required mill power.

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17.1.5 Flotation

The flotation plant will produce a talc concentrate for disposal, as well as three payable base metal concentrates for shipment to market and sale. The talc concentrate will be produced prior to metal flotation to remove the talc and prevent possible contamination of the base metal concentrates. The base metal flotation process is industry standard and includes the production of a bulk copper and lead concentrate which is subsequently separated into copper and lead concentrates as well as a zinc concentrate.

The flotation tailings, which are a combined talc concentrate and zinc rougher tailings, will be pumped to the TMF for final disposal.

Talc Flotation

The cyclone overflow at 33% solids will feed five 50 m³ forced air tank flotation cells for removal of talc minerals. The talc rougher flotation concentrate will be further cleaned using flotation to reject any entrained payable sulphide minerals. The pre-flotation tailings will become feed to the copper-lead bulk flotation circuit, while the talc concentrate will be sent to the final tailings pump box for disposal.

Copper-Lead Bulk Flotation And Regrinding

The talc pre-flotation tailings will flow into the copper-lead bulk flotation conditioning tank where zinc depressants and copper-lead mineral collectors will be added. The conditioned slurry will be floated in five conventional 100 m³ tank flotation cells to produce a bulk copper-lead concentrate. The concentrate from the rougher bulk flotation circuits will be reground in a 750 kW vertical regrind mill. The regrind mill will be operated in closed circuit with cyclones 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 using column flotation.

The first stage of bulk copper-lead cleaner flotation will be conducted in one 5.0 m diameter by 10.0 m high flotation column. The first cleaner flotation tailings will be scavenged to further recover copper-lead minerals in five 10 m³ tank flotation cells. Concentrate from the first cleaner/scavenger flotation will flow to the regrinding pump box for further regrinding, while the tailings produced from the rougher/scavenger flotation stage will be sent to the copper-lead rougher/scavenger tailings pump box. Combined rougher/scavenger slurry from the tailings pump box will be pumped to the zinc flotation circuit.

Bulk copper-lead concentrate from the first cleaner flotation stage will be pumped to the second cleaner flotation stage, a 5.0 m diameter by 10.0 m high flotation column. Concentrate from the copper-lead second cleaner flotation will be pumped to the bulk copper-lead bulk concentrate thickener prior to being pumped to the copper and lead separation flotation circuit. Overflow of the thickener will be recycled to the bulk copper-lead flotation stage. Tailings from the copper-lead secondary cleaner flotation will return to the copper-lead first cleaner flotation.

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Bulk copper-lead rougher flotation and cleaner flotation will be carried out at a pH of 8.0 to 8.5. Reagents that will be used in the circuit include:

· ZnSO₄ as a zinc depressant

· SIPX and 3418A as collectors

· MIBC as a frother.

The major process equipment used in the copper-lead bulk flotation circuit will include:

· Two conditioning tanks

· Five 100 m³ rougher and scavenger flotation cells

· One vertical regrind mill with an installed power of 750 kW

· Classifying hydrocyclones

· One 5.0 m diameter by 10.0 m high flotation column for first cleaner flotation

· Five 10 m³ tank flotation cells for first cleaner/scavenger flotation

· One 5.0 m diameter by 10.0 m high flotation column for second cleaner flotation

· One 10.0 m diameter high rate thickener for the cleaned bulk concentrate.

Copper-Lead Separation Flotation Circuit

Underflow from the bulk copper-lead cleaner concentrate thickener will be pumped to the copper-lead separation flotation conditioning tanks, where sodium cyanide and lime will be added to depress copper minerals. Lead collector will be added in the second conditioning tank. The conditioned slurry will flow to the lead rougher flotation cells.

Lead concentrate produced in the lead rougher cells will be pumped to the first lead cleaner flotation circuit. Concentrate from the lead scavenger flotation cells will in turn be pumped to lead rougher flotation cells. Tailings from the rougher/scavenger flotation, effectively copper concentrate, will report to the copper concentrate thickener.

Lead concentrates from the lead rougher and scavenger flotation circuit will be further upgraded by three stages of cleaner flotation. Lead concentrates from the first stage of cleaner flotation will be fed to the second stage of lead cleaner flotation. The lead cleaner tailings, together with the lead 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 gravity flow to the proceeding flotation cell feed boxes.

Lead flotation will be operated at a pH range of 9.0 to 9.5. Reagents that will be used in the circuit include:

· Lime for adjusting slurry pH

· Sodium cyanide as a copper mineral depressant

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· 3418A as a collector

· MIBC as a frother.

The copper- lead flotation circuit will consist of:

· Two conditioning tanks

· Five 10 m³ lead rougher and rougher/scavenger flotation cells

· Five 5 m³ lead first cleaner and scavenger flotation cells

· Five 5 m³ lead second cleaner flotation cell

· Five 5 m³ lead third cleaner flotation cell.

Zinc Flotation and Regrinding Circuit

The bulk copper-lead rougher scavenger tailings along with the copper-lead first bulk cleaner scavenger tailings will be feed to the zinc flotation circuit. The zinc circuit will consist of feed slurry conditioning, rougher/scavenger flotation, zinc rougher concentrate regrinding, and two stages of cleaner flotation.

Bulk copper-lead flotation tails will be conditioned with lime to increase the pH to approximately 10.5, and copper sulphate will be added to the slurry to activate zinc minerals. Conditioning of the zinc flotation feed slurry will be completed in two tanks in series.

The flotation circuit will produce a zinc rougher/scavenger concentrate and zinc rougher/scavenger flotation tailings. The rougher/scavenger concentrates will report to the zinc concentrate regrinding circuit. Zinc rougher/scavenger tailings will flow to the final tailings pump box to be delivered to the TMF.

Zinc concentrate from the rougher/scavenger flotation stage will report to the regrinding cyclone feed pump box, and be fed in turn to the zinc regrind mill. The zinc regrind mill is a tower mill, with an installed power of 750 kW. Mill discharge is pumped to the regrinding cyclone cluster where it is classified. Overflow from the cyclones, having a particle size of approximately 80% passing 40 µm, will report to the first zinc cleaner flotation circuit; oversized material will flow back to the regrind mill.

There will be two stages of zinc cleaner flotation, both using standard column flotation equipment. The zinc cleaner circuit will operate at a pH of 11 or above to reject pyrite. 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 and from there tailings will be pumped to the TMF. 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.

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Reagents used in the zinc flotation circuit will include copper sulphate, lime, SIPX, and MIBC.

The main equipment used for the zinc flotation circuit will consist of:

· Two conditioning tanks

· Six 50 m³ zinc rougher and rougher scavenger flotation cells

· One 750 kW tower mill

· Classifying cyclones

· One 4.5 m diameter by 9.0 m high flotation column for the first zinc cleaner flotation

· Five 5 m³ tank cells for the first zinc cleaner scavenger flotation

· One 4.5 m diameter by 9.0 m high flotation column for the second zinc cleaner flotation.

Product Dewatering

Each of the flotation concentrates, including copper, lead, and zinc concentrates, will be thickened in individual thickeners. Concentrates will be further dewatered by pressure filters to a design moisture content of 9%; pressure filtration has been selected to achieve a low overall cake moisture.

Copper Concentrate Dewatering

Copper concentrate will be pumped to a 8.5 m-diameter high-rate thickener. The copper concentrate will be mixed with diluted flocculant solution at the thickener feed well. Prior to pressure filtration, thickener underflow slurry at approximately 65% solids will be pumped to an agitated concentrate stock tank with six hours storage capacity. Thickener overflow solution will be treated by a sulphur dioxide-air procedure to detoxify residual cyanide prior to being sent to the final flotation tailings pump box from where the slurry is delivered to the TMF.

The target filter cake design moisture is 9%. The copper concentrate will be discharged into a stockpile, from which FELs will load concentrate into containers for shipment. Filtrate from the filtration will return to the copper concentrate thickener for incorporation into the overflow solution. The in-plant storage will be capable of storing up to five days of copper concentrate production, to accommodate potential truck haulage interruptions.

The equipment required for concentrate thickening and filtration will include:

· One 8.5 m diameter high rate thickener

· One six hour concentrate stock tank

· 35 m² filtration area

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Lead Concentrate Dewatering

The lead concentrate will be blended with flocculant solution and discharged 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 provide six hour concentrate storage capacity. Thickener overflow will be directed to the lead rougher concentrate pumpbox in the lead separation circuit for recycle.

Filtered lead concentrate, at a design moisture content of 9%, will be loaded into a 20 ft concentrate container, and subsequently hauled to the ocean port for shipment. Filtrate will return to the lead concentrate thickener feed well for incorporation into the overflow solution stream.

The lead concentrate storage facility will be capable of storing at least five days of production at design rates

The equipment required for lead concentrate thickening and filtration will include:

· 3.5 m diameter high rate thickener

· Agitated concentrate stock tank

· 8 m² filtration area

Zinc Concentrate Dewatering

Zinc concentrate slurry will be directed to an 8.0 m diameter high-rate thickener. Flocculant will be added to improve settling of the concentrate. Thickener underflow slurry at 65% solids will be stored in an agitated concentrate stock tank prior to pressure filtration. The concentrate stock tank is designed to provide six hours of zinc concentrate storage at design rates. Thickener overflow will be delivered to the final flotation tailings pump box.

Filtered zinc concentrate at a design moisture content of is 9% will be conveyed and discharged to a concentrate stockpile, and a FEL would load the concentrate into containers for shipment off site to market. Filtrate will return to the zinc concentrate thickener for recycle within the plant

Zinc concentrate storage and dispatch facility will be capable of storing at least five days of production at design rate to provide for potential interruptions in shipping service.

Equipment required for concentrate thickening and filtration will include:

· One 8.0 mm diameter high-rate thickener

· One agitated concentrate stock tank

· 30 m² filtration area

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Tailings Disposal

The final flotation tailings will be pumped in four stages, approximately 275 m in elevation from the processing plant to the TMF for disposal. The tailings stream is a combination of the following:

· The talc flotation concentrate

· The zinc rougher scavenger tailings

· The zinc first cleaner scavenger flotation tailings

· Four stages of pumping.

The tailings management facility 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 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 recovery of valuable minerals to the concentrate products. 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 from the fresh water tank. Preparation of the various reagents will require:

· Bulk handling systems

· 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 drums. 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 lime 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; the bulk of the lime slurry will be required in the zinc flotation circuits. Lime feed to the slaker will be controlled automatically, based on the lime slurry levels in the holding tank.

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Flocculant

Stock flocculant solutions will be made up from powdered flocculants in a packaged preparation system, including a screw feeder, a flocculant eductor, 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.

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 antidotes will be provided in various areas as required.

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 allow for preparation of individual reagents not named here which may occasionally be tested during operations to determine the effect on metal recovery or 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. The WTP will treat effluents generated from mining operations.

Fresh Water Supply System

Process fresh water will be supplied from the water treatment plant or local creeks. A 1000 m³ 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

· Process make up water

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Potable water will be supplied from local well and chlorinated prior to delivery to the potable water distribution loop.

Process Water

Process water will be made up of fresh water, and water reclaimed from the TMF. Reclaimed water and fresh water will be directed to a 10.6 m diameter by 12.7 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:

o Low-pressure air will be provided for flotation tank cells by air blowers. High pressure wet air will be provided for the flotation column by air compressors.

· Filtration circuit:

o Wet high-pressure air will be provided for filtration and drying by air compressors.

· Plant air service:

o Wet high-pressure air will be provided for various services by dedicated air compressors.

· Instrumentation air:

o Dry high-pressure 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 fluorescence analyzers. The sampling and analysis system will also collect assay samples for further chemical analysis. The on-line assay data will be fed back to the central control room and used to manage and 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 process control, 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

· AAS

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· ICP for the routine assay laboratory

· ICP-MS for the environmental laboratory

· a LECO furnace

· Other determination instruments such as pH and redox potential meters and experimental balances.

The metallurgical laboratory will perform metallurgical tests for quality control and process flowsheet optimization as required. The laboratory will be equipped with laboratory crushers, ball mills, particle size analysis devices, laboratory flotation cells, balances, and pH meters.

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 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 at all times.

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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18.0 Project InfrastrUcture

18.1 Introduction

The proposed Arctic mine is a greenfields site, remote from existing infrastructure Figure 18-1 shows the proposed locations of the processing plant, mine infrastructure buildings and administration buildings. Figure 18-2 shows the proposed site layout.

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Figure 18-1    Proposed Location of the Processing Plant and Other Buildings (Ausenco, 2018)

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Figure 18-2     Proposed Site Layout (Ausenco, 2018)

  

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18.2 Access Roads

The Project site will be accessed through a combination of State of Alaska owned highways, proposed AIDEA owned private road, and proposed Trilogy Metals owned access roads.

18.2.1 Ambler Mining District Industrial Access Project Road

The Project assumes that the Ambler Mining District Industrial Access Project (AMDIAP) road will be constructed prior to commencing construction of the Arctic Project.

There is currently no developed surface access to the Arctic Project area and beyond. Access to the Arctic Project is proposed to be via AMDIAP, a road approximately 340 km (211 miles) long, extending west from the Dalton Highway where it would connect with the proposed Arctic Project area.  The final terminal for the road has not yet been determined.

The AMDIAP road is being permitted as a private road with restricted access for industrial use and has just completed the scoping phase of the environmental impact statement (EIS). The notice of intent currently states that the draft EIS will be completed by March 29, 2019 and the final EIS will be completed by December 30, 2019. Figure 18-3 shows the proposed route of the AMDIAP road.

Figure 18-3     Proposed Route of AMDIAP Road (Ambler Access Website 2018)

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18.2.2 Arctic Access Road

To connect the Arctic Project site and the existing exploration camp to the proposed AMDIAP road, a 30.7 km access road will need to be built. Existing roads can be upgraded for 13.8 km of the required 30.7 km.

The Arctic access road will be built to accommodate an 80 t vehicle, with a speed limit of 40 km/h. The road will have a maximum sustained grade of 8%. The design life of the road is 15 years. It is anticipated that the road construction could be completed within one-year. Figure 18-4 shows the proposed Arctic access road.

Figure 18-4     Arctic Access Road (AllNorth 2017)

18.3 Airstrip

The Dahl Creek airport is situated approximately 32 km south of the Arctic Deposit. The airport would need to be upgraded with a lighting system and an automated weather observation system to be functional for the purposes of the Project. These upgrades would support the use of Dash 8 aircrafts for transporting crew to and from Fairbanks International. The airport is publicly owned, and it is assumed that the upgrades would be accepted if funded by the Project.

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18.4 Camps

The project will require the use of three different accommodation camps. Each camp will be self-contained and have its own power generation and heating capabilities, water treatment plant, garbage incineration and sewage treatment plant.

18.4.1 Bornite Exploration Camp

There is an existing camp at Bornite, which is currently used as an exploration camp. This camp can hold approximately 70 people with some minor upgrades. This camp will be used to support the construction of the Arctic access road.

18.4.2 Temporary Construction Camp

Due to the remote location of the Project site, a construction camp will be required. This camp will be constructed at the intersection of the AMDIAP road and Arctic access road. The camp will be constructed to provide room and board for 200 people. After the Project construction is completed this camp will be removed.

18.4.3 Construction / Operations Permanent Camp

The permanent camp will be constructed along the Arctic access road, closer to the processing facility. The permanent camp will be constructed ahead of operations to support the peak accommodation requirements during construction. The camp will be constructed to provide room and board for 450 people.

18.5 Fuel Supply, Storage and Distribution

The processing facility will use liquefied natural gas (LNG) that will be supplied via existing fuel supply networks near Point Mackenzie, Alaska. The existing commercial supply of LNG is typically delivered to industrial consumers sites in 16 m long overall LNG transport tractor-trailers. The trailer is pumped out by dedicated cryogenic unloading pumps at the reception site and then is returned to the supplier for refill.

Shop fabricated steel tanks will be used for LNG storage. A total of six vertical tanks, each 19 m high by 3.4 m diameter would provide approximately 702 m³ net storage, which equates to approximately five days of storage.

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18.6 Power Generation

Power generation will be provided by six LNG generators. This arrangement will allow for four in operation, one in stand-by, and one undergoing maintenance. Each generator will be rated for 4.4 MW of output, yielding 17.6 MW of power supply. All six generators will be housed within a shared building. The supply voltage will be 4.16 kV.

18.7 Electrical System

The plant power distribution from the powerhouse will be via buried conduits wherever possible. The distribution voltage to the local electrical rooms will be 4.16 kV.

There will be five electrical rooms. There will be one at the powerhouse, one at the tailings management facility, one at the crushing plant area, one at the main process area and one at the main process building. The electrical rooms will be pre-fabricated and loaded with electrical equipment prior to delivery to site.

The total connected load will be 17.5 MW with an average power draw of 12.6 MW.

18.8 Surface Water Management

The proposed mine development is located in valley of Sub-Arctic Creek, a tributary to the Shungnak River. The catchment area of Sub-Arctic Creek above the monitoring location at the mouth of Sub-Arctic Valley is 26.5 km². All mine infrastructure and mine affected water is within the Sub-Arctic Creek valley. A surface water management system will be constructed to segregate contact and non-contact water. Non-contact water will be diverted around mine infrastructure to Sub-Arctic Creek. Contact water will be conveyed to treatment facilities prior to discharge to the receiving environment.

SRK (SRK Consulting, 2018) describes the site water management plan in detail. The objectives of the water management system are to:

· Ensure sufficient water quantity is available to support processing.

· Manage process water quality such that metal recoveries are not impaired.

· Manage contact and non-contact water separately to minimize volume of contact water collected on site.

· Collect and treat impacted water that would otherwise impair water quality of the receiving streams.

· Minimize erosion of natural soils and mine infrastructure to reduce suspended solid loading in surface runoff discharge.

Water on site is classified into three types:

· Non-contact water is diverted away from mine infrastructure to the extent possible to reduce infiltration into the waste rock dump and pit inflows.

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· High-sediment water (i.e. surface flows from pads and topsoil stockpiles) is collected and conveyed to downstream sedimentation ponds to treat for total suspended solids prior to discharge in the Sub-Arctic Creek.

· Contact water from the waste rock dump or the TMF.

Figure 18-5 illustrates the water management plan for the maximum footprint of the project during operations.

Figure 18-5     Surface Water Management Plan during Operations

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18.8.1 Process Water Supply

Process water will be supplied primarily from reclaim in the TMF. The TMF has a positive water balance and excess water will need to be discharged from the TMF annually. Gland water will be withdrawn from the Waste Rock Collection Pond (WRCP). The TMF dam will be constructed prior to the processing of ore. Water will be allowed to collect in the TMF initially to have an adequate volume (2 Mm³) to supply the mill during startup.

An additional water management criterion for the TMF is to provide at least 100 days residence in the reclaim pond. The rationale for this is to allow enough time to allow the degradation of mill reagents such the reuse of this water does not compromise metal recoveries by floating metals in during processing. To meet this 100 day residence time criteria, a minimum volume of 2,100,000 m³ of free water will be kept in the TMF. The 100 day requirement to reduce the recycling of flotation reagents will be evaluated in subsequent metallurgical characterizations as the project advances.

18.8.2 Water Management Infrastructure

There three types of water conveyance and storage infrastructure:

· Diversion and collection channels

· Culverts

· Containment structures.

Diversions and Collection Channels

Diversion channels will be constructed at the maximum extents of the waste rock dump and around the edge of the pit to divert non-contact water away from these areas to reduce contact water volume. Runoff from the pit walls and pit water will be pumped to WRCP. All surface runoff and seepage from the waste rock dump will flow into the WRCP.

Collection channels will be positioned downgradient of mine infrastructure pads and cover soil stockpiles to convey surface runoff to settling ponds prior to discharging to the Sub-Arctic Creek. Collection channels will convey water from the toe of the plant pads, overburden and topsoil stockpiles to the settling ponds above Sub-Arctic Creek. The collection channels will be excavated into existing ground at the maximum extent of the stockpiles and infrastructure pads. Water from the WRCP will be pumped to the treatment plant, treated and discharged to the Shungnak River during the open water season.

Non-contact water diversions and contact water diversions are designed to convey the peak flow from a 100-year and 10-year 24-hour rain-on-snow event respectively.

All channels will have a minimum depth and width of 1.0 m and will be lined with a nonwoven geotextile, followed by a layer of rip rap for erosion protection. Rip rap was sized to ensure velocities in the channel are below 5 m/s. The riprap thickness is specified at two times the D₅₀of the riprap and the geotextile fabric is intended to limit erosion, to reduce sediment loading downstream, and reduce the chance of channel embankment failure. Channel depths were designed to have a minimum freeboard of 0.3 m above the design peak flow water level in the channel to allow for sediment accumulation in the channel, unforeseen climatic event or other uncertainties.

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Culverts

Culvert crossings will be required to maintain drainage patterns across haul roads and access roads using circular corrugated steel pipes (CSP). The culverts will convey flows from diversion channels and collection channels and are sized for the 100-year 24-hour rain-on-snow event. The culverts will be placed in a compacted fine-grained engineering fill according to the manufacturers’ specifications.

Containment Structures

Runoff from the project will be routed to ponds prior to treatment and/or discharge to Sub-Arctic Creek or the Shungnak River. Two types of ponds were sized for the Arctic Project: sedimentation ponds and the WRCP.

Runoff from the plant area, overburden stockpile and topsoil stockpile will be routed to downstream sedimentation ponds. These ponds will settle suspended solids and attenuate peak flows. The ponds will have a minimum depth of 2 m and at least a 3:1 length to width ratio. These sedimentation ponds were designed to contain the volume from the 10-year rain-on-snow event with an overflow spillway designed to convey the peak flow from a 200-year storm.

Flocculent addition, filtration curtains and/or or baffles are contingency measures that can be used to meet enhance settling. Sediment will accumulate within the pond and will be removed periodically to maintain the design capacity.

18.8.3 Waste Rock Collection Pond

The WRCP will be located directly below the WRD to collect runoff from the dump, seepage from the rock drain and water from dewatering the pit. Gland water for the process plant will be withdrawn from the pond. Excess water will be treated during the open water season to remove metals prior to discharge to the Shungnak River or recycled back to the TMF when the treatment plant is not operating. The pond was sized to store the 100-year 24-hour rain-on-snow storm volume with additional capacity water (~8,000 m³) for processing. The total pond capacity will be 132,000 m³ to the spillway invert of 643.5 masl with a crest height of 645 masl.

The dam for the pond will be constructed using the overburden material stripped from the TMF footprint and placed on the existing colluvium overburden material in compacted 1 m lifts. The down-stream face of the dam will be constructed to 3.5H:1V while the upstream face will be 2.5H:1V. The final upstream face and the full bottom of the pond will be lined with an HDPE geomembrane.

18.8.4 Site Water and Load Balance

A water and load balance model was built using the dynamic system modeling software Goldsim (version 12.0). The model is based on the water management plan described in Section 18.8.

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The water balance model predicts water use, surpluses and deficits for the site and mine water management infrastructure (TMF reclaim pond and WRCP) over the 12 year mine life. Mine facility footprints vary according to the mine plan over the mine life. Only average hydrological conditions were evaluated by the model.

The load balance integrates source terms with the water balance. Source terms which define water chemistry of each water type were developed and assigned to flows in the model. The source terms consist of concentrations that are either historical statistics, or estimates of concentrations based on preliminary geochemical test work. Mass loads are calculated by multiplying the source term concentration by the flow rate. Constituent mass loads and flows are mixed to estimate concentrations at locations on the site and in the receiving environment.

The results from the preliminary water and load balance indicate that excess water from the WRCP and TMF will needed to be treated prior to discharge to the receiving environment. The TMF has high ammonia and cyanide concentrations requiring treatment. The WRCP collects water that is high in dissolved metals.

18.9 Water Treatment Plant

The results from the preliminary water and load balance were used to develop a water treatment strategy. Contact water will be treated in two parallel treatment systems: a metals treatment plant receiving water from the waste rock pond and an ammonia oxidation biological plant receiving water from the TMF.

The water treatment plants will only operate during the open water season from May through September. The effluent from both treatment plants will be combined and discharged via an 11 km pipeline to the Shungnak River.

Selenium concentrations in the process water and waste rock runoff are predicted to be high. The water treatment systems are unlikely to remove appreciable amounts of selenium. Discharging the combined effluent to the Shungnak River is the proposed selenium management option. The water quality in the Shungnak River below the discharge point is predicted to meet water quality criteria after a mixing zone. The discharge location in the Shungnak River and mixing zone will both require regulatory approval. Obtaining a permit to discharge into and a mixing zone in the Shungnak River is a regulatory risk for the Project.

18.10 Tailings Management Facility

18.10.1 General Description

The TMF will be located at the headwaters of the Sub-Arctic Creek, in the upper-most portion of the creek valley (refer to Figure 18-5). The maximum storage capacity of the facility will be about 33 Mm³ (43.5 Mt). The 58.6 ha footprint of the TMF will be fully lined with an impermeable liner (HDPE).

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Tailings containment will be provided by an engineered dam that will be buttressed by the WRD constructed immediately downstream of the TMF and the natural topography on the valley sides. A starter dam will be constructed to elevation 830 m. Three subsequent raises will bring the final dam crest elevation to 890 m, which is the same as the final elevation of the waste rock dump. The TMF is designed to store approximately 27.3 Mm³ (38.7 Mt) of tailings plus 3.0 Mm³ of water produced over the 12 year mine life as well as the PMF and still provide 2 m of freeboard.

Tailings will be deposited as conventional slurry from the dam crest. The settled tailings density is assumed to be 1.4 t/m³ based on preliminary data provided by Ausenco. Beach slope angles are assumed to be 2% for sub-aerially deposited tailings. The reclaim pond will be forming against the natural terrain upstream of the dam, to reach a maximum design elevation of 880.8 m.

18.10.2 Design Criteria

The basis of the TMF design is provided in Table 18-1. Values were determined from project-specific information, judgment and experience with other projects.

Table 18-1     TMF Design Parameters and Design Criteria

Design Item	Criterion	Reference
Operational life of TMF	12 years	Trilogy
Total tailings	38.7 Mt	Ausenco
Annual Tailings Production	2.36 Mm3/yr	Ausenco
Tailings percent solids (in the slurry pipeline)	30% (Ms/(Ms+Mw))	Ausenco
Tailings solids specific gravity	3.1	Ausenco
Tailings settled dry density	1.41 t/m3	Ausenco
Expected tailings beach	2.0%	SRK
Tailings Deposition	Spigots from dam crest	SRK
Target storage requirement	30.4 Mm3 (including 3 Mm3 of free water)	SRK
Dam Hazard Classification	Class II	ADSP
Minimum Freeboard above tailings	2 m	SRK
Tailings dam crest width	30 m	SRK
Tailings dam max elevation (height)	890 masl (150 m)	SRK
Design Earthquake	1:2475 year, PGA = 0.2645	ADSP
Stability Factor of Safety (FOS)	1.0 (Seismic) to 1.5 (Static)	ADSP
Dam construction materials	Waste rock compacted in 1 m lifts	SRK
Dam downstream slope (prior to WRD placement)	3H:1V	SRK
Dam upstream slope	2½ H:1V	SRK
Starter dam storage capacity (elevation)	5.2 Mm3, approaching yr 1, (830 masl)	SRK
Dam Raise 1 storage capacity (elevation)	14.1 Mm3, approaching yr 5, (855 masl)	SRK
Dam Raise 2 storage capacity (elevation)	28.1 Mm3, approaching yr 10, (880 masl)	SRK
Dam Raise 3 storage capacity (elevation)	35.0 Mm3, end of mine, yr 12, (890 masl)	SRK

Notes: Dam classification follows Guidelines for Cooperation with the Alaska Dam Safety Program (ADSP, 2005, 2017)

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18.10.3 Overburden Geotechnical Investigation

An overburden geotechnical investigation was carried out in the summer of 2017 to provide overburden characterization in support of the waste facility siting evaluation. This investigation was carried out using both sonic drilling and test pitting.

A total of seven test pits and one sonic borehole were completed at the Site 1 location, which was ultimately chosen to host the mine waste facilities. The overburden is characterized as shallow well-graded silty sand or gravel overburden overlying fractured or weathered bedrock at depths between 2.4 m and 6.1 m.

The groundwater level was measured immediately after completion of the drill hole and a monitoring well was installed for future field testing and monitoring. Laboratory tests were subsequently completed on select samples obtained during the program. Field investigation logs were prepared for each location and the laboratory test results were added to the respective log (SRK 2017 Investigation Report).

18.10.4 Site Selection

The site of the TMF was selected in August 2017 during a workshop to evaluate locations for mine waste facilities (TMF and WRD). Workshop participants included representatives from the PFS team (Trilogy, Ausenco, Amec Foster Wheeler, & SRK). The workshop was facilitated by a representative from Ausenco who subsequently prepared a report summarizing the workshop process and its key outcomes (Ausenco, 2017). The selection of the TMF site was influenced by:

· The minimum footprint available to accommodate both the WRD and TMF

· Field investigations to evaluate foundation conditions

· Proximity to the pit and mill facilities

A weighting system was applied to four broad categories including environmental concerns, permitting, capital costs, and operating costs.

18.10.5 Starter Dam

The foundation for the starter dam will be cleared, topsoil removed, and overburden stripped to bedrock. Along the valley bottom the overburden thickness is expected to be approximately 4 m, thinning to approximately 3 m on average on the abutment valley side slopes. The topsoil and overburden material will be stockpiled near the pit for use in future reclamation of the waste facilities.

The starter dam at TMF will be constructed to crest elevation 830 m, which will allow for storage of the pre-production water and approximately 1 year of storage after mill production begins. The upstream and downstream faces of the starter dam will be constructed at 2.5:1 and 3:1 respectively and entirely of waste rock from the open pit. Waste rock will be placed in 1m lifts and compacted. Figure 18-6 shows a cross section through the starter dam and the abutting WRD.

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The tailings impoundment footprint and the upstream face of the dam will be lined with a single side textured 60 mil HDPE geomembrane, placed over 1 m of bedding material. To prepare for liner installation, the impoundment area will be cleared, grubbed and stripped of topsoil prior to bedding material placement. The bedding material will protect the liner against puncture from potential sharp clasts in the overburden and exposed bedrock within the impoundment footprint.

Figure 18-6     Cross Section through the TMF Starter Dam to Elevation 830 m

18.10.6 Dam Raises and Final Dam

Three dam raises will be completed to reach the final dam height of 890 m. The construction for these raises will be completed in years 1, 5, and 10 to elevations 855 m, 880 m and 890 m respectively. Construction will be completed using the downstream method and will connect the structural fill in the tailings dam with the uncompacted WRD. In order to achieve this, tailings deposition will be required from the perimeter of the TMF during the construction sequence of the dam crest. The dam crest will be constructed to a width of 30 m at the end of each construction campaign.

An upstream slope of 2.5:1 will be maintained throughout the construction of the dam to facilitate the placement of the bedding layer and the installation of the liner. The downstream portion of the dam will abut the WRD at each stage of dam raising. Construction material for the dam will be waste rock from compacted in 1 m lifts.

Additional liner will be placed within the expanded tailings impoundment footprint using the same procedures as for the initial starter dam. The impoundment area will be cleared, grubbed and bedding material placed prior to the single side textured 60 mil HDPE geomembrane being laid down and seamed to the existing liner.

Figure 18-7 shows a cross section at the final design elevation.

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Figure 18-7     Cross Section of the TMF and WRD at Final Design Elevation

18.10.7 TMF Water Pool and Water Return

Under normal situations, the pool will be maintained in a manner that will maintain a beach to provide separation between the pool and the perimeter dam. Recycle water will be obtained using a floating barge and pipeline system situated on the pool. The free water volume will be maintained at a minimum of 2 Mm³ to allow sufficient volume for conditioning time of the reagents prior to water being recirculated back to the mill.

Tailings deposition in winter will be managed to limit ice formation in the tailings, however an allowance for up to 10% ice entrainment can be accommodated by the current TMF capacity.

18.11 Tailings Delivery and Return System

The tailings delivery system will transport slurried tailings from processing plant to the TMF. The delivery system will be sized initially on the basis of a 10 kt/d operation. This will consist of one 400mm (16 inch) diameter carbon steel rubber lined pipeline, approximately 3 km long. This pipeline will transport 1,050 m³/h of tailings to the TMF.

The return water delivery system for recycle water from the TMF has been sized on the basis of 770 m³/h of water being pumped from the TMF to the process water pond, for the 10 kt/d operation. This system will consist of a barge pump and one 300mm (12 inch) diameter carbon steel pipeline, approximately 3 km long, adjacent to the tailings pipeline. This pipeline will be heat traced to prevent freezing.

18.12 Waste Rock Dump and Overburden Stockpiles

A large WRD will be developed north of the planned Arctic pit in the upper part of the Arctic valley (Figure 18-8). The waste rock placed in the most northern portion of the WRD will be compacted to provide the structural fill for the TMF tailings dam as described in Section 18.10. There will also be two small overburden stockpiles to store the stripped topsoil and overburden from the TMF footprint.

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Figure 18-8     General Location of WRD and Stockpiles

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18.12.1 Waste Rock Dump

The waste rock from the pit will be placed in a single large WRD located in the Sub-Arctic Creek valley, upstream of the proposed pit (refer to Figure 18-8). The total volume of waste rock is expected to be 145.6 Mm³ (296 Mt); however, there is potential for expanded volume in the waste if placement density is less than 2.0 t/m³. Most of the waste rock is anticipated to be PAG and there will be no separation of waste based on acid generation potential. Rather, seepage from the WRD will be collected and treated.

The dump will have a final height of 245 m to an elevation of 890 masl and is planned to be constructed in 20 m lifts with benches at 23.5 m on average at the dump face, to achieve an overall slope of 2.7H:1V. The final design is expected to have variations in the slope aspect, to best manage overland flow, resulting in a transitioning convex to concave shape of the dump face.

A rockfill underdrain will be constructed under the WRD in the current Sub-Arctic Creek channel bed. The underdrain will be excavated into the overburden prior to dump construction and will be capable of handling base flow through the Sub-Arctic Creek valley. Water will be collected in a pond at the base of the WRD and held for treatment.

18.12.2 Overburden and Topsoil Stockpiles

Two stockpiles will be developed on the western side of the planned pit to store topsoil and overburden materials for use in final reclamation of the site. The topsoil stockpile will be placed in between the haul roads and will store up to 225,000 m³ of material while the overburden stockpile will be located below the lower haul road between the pit and the mill site with storage capacity up to 650,000 m³.

Collection channels have been designed below both of the stockpiles and will convey water into two different sedimentation ponds before discharging the water into Sub-Arctic Creek.

18.13 Compressed Air Supply

High-pressure compressed air will be provided by two duty and one standby screw compressors and a duty plant air receiver.

There will be three high-pressure air uses: instrument air, plant air and concentrate filter air. The instrument air will be dried and then stored in a dedicated air receiver. The plant air and filtration air will be fed straight from the plant air receiver.

Low-pressure air for flotation-cell air requirements will be provided by two duty and one standby centrifugal blowers.

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18.14 Site Communications

On-site communications will consist of inter-connected mobile and fixed systems, including landline telephone network, radios and internet. It is assumed that a network connection will can be made via satellite.

18.15 Fire Protection

The firewater distribution network will be maintained under a constant pressure with a jockey pump and will be looped and sectionalized to minimize loss of fire protection during maintenance. Where run outside buildings, fire water piping will be above ground and be heat traced and insulated.

Yard hydrants will be limited to the fuel storage tank area. Wall hydrants will be used in lieu of yard hydrants, and these will be located on the outside walls of the buildings in heated cabinets.

Fire protection within buildings will include standpipe systems, sprinkler systems and portable fire extinguishers. Standpipe systems will be provided in structures that exceed 14 m in height and additionally where required by regulations, local authorities or the insurance underwriter.

Camp modules will be purchased with fire detection; fire rated walls and will use separation as a means of fire protection. Handheld extinguishers will be located throughout the buildings.

Fire protection of the generators will be provided by a water mist system. Gas detection will be provided to detect dangerous levels of LNG gas within the generator room.

18.16 Plant Buildings

18.16.1 Gatehouse

The guardhouse will be located on the site access road where security staff can control entry to the mine and process plant areas. The building will be constructed as a single-storey wood frame building with nominal ablution facilities to support the occupants on shift. This is recommended to be modular construction on skids to simplify construction and facilitate relocation in the future. This building will be heated using electric unit heaters.

18.16.2 Mine Infrastructure Area

The mine infrastructure area is designed at 2,000m² and will be positioned on an upper pad next to the fuel storage area. This building will consist of the truck workshop, truckwash, mine offices, mine dry and warehouse. The truck workshop and warehouse will have lifting and handling activities fulfilled by an overhead gantry crane. This building will be a pre-engineered steel frame and metal-clad building. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

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18.16.3 Laboratory

The laboratory will be 350 m² in area and will be situated adjacent to the process building. The building will house all laboratory equipment for the daily operational process control including the metallurgical and environmental requirements. The site metallurgist’s office will also form part of the building. Any mechanical items associated with the dust collection equipment will be located external to the building. The building will be constructed as a single-storey modular wood-frame building and will be connected to the process building with arctic corridor. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.4 Administration Building

The administration building will consist of plant offices for the process plant area. The building will have an approximate area of 600 m². The building will be constructed as a single-storey modular wood-frame building. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.5 Mill Dry Facility

The mill dry facility will consist of plant change rooms for the process plant area. The building will be approximately 100 m² in area. These facilities will have clean and dirty areas and will be complete with showers, basins, toilets, lockers and overhead laundry baskets. The building will be constructed as a single-storey modular wood-frame building. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.6 Plant Workshop and Warehouse

The plant workshop will be used to perform maintenance on process equipment and equipment spares. The plant workshop will be approximately 790 m² in area. This building will be a pre-engineered steel frame and metal-clad building. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.7 Primary Crushing

The primary crushingbuilding will be used to house the crusher equipment. The building will have an approximate area of 200 m². This building will be a pre-engineered steel frame and metal-clad building.

18.16.8 Fine Ore Stockpile

The fine ore stockpile will be used to keep the ore dry during the winter. The building will be approximately 2,700 m² in area. This building will be a pre-engineered steel-frame building. Cladding and insulation of a building this large may include insulated steel, or a tensioned fabric system with integral insulation.

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18.16.9 Process Plant

The process plant building will have an approximate area of 4,880 m² (in two sections: 30 m wide x 45 m long and 36m x 98m), and will house all the milling, flotation and concentrate thickening and filtering equipment. The building will be divided into two sections. The first section will contain the mill; the second section will contain the flotation and filtration equipment. Both sections will be serviced by overhead cranes. This building will be a pre-engineered steel frame and metal-clad building with internal insulation to reduce heat loss. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.10 Concentrate Loadout

The concentrate loadout building will house process equipment and provide a covered area for loading the concentrate on to the trucks. This building will be separated from the process plant building to minimize the building volume that requires process ventilation equipment. The building will be approximately 1,000 m² in area. This building will be a pre-engineered steel frame and metal-clad building.This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.11 Reagent Storage and Handling

The reagent storage and handling building will be located outside the process plant building, and will be connected with an arctic corridor. The building will have an approximate area of 720 m². This building will be pre-engineered steel frame and metal-clad building. This building will be heated using an air handler system that utilizes waste heat produced from the LNG generators.

18.16.12 Raw Water Supply

The raw water supply building will be located at the fresh water source and house the pumping equipment. The building will be 36 m² in area and will be a pre-engineered modular building. This building will be heated using electric unit heaters.

18.17 Concentrate Transportation

Concentrate will be shipped from the Arctic mine site to the Port of Alaska in Anchorage in specialized 6 m (20 ft) intermodal bulk shipping containers for direct loading into bulk carrier vessels for ocean transport to the smelter or refinery. Containers will be trucked to Fairbanks, AK, and then transferred to rail for delivery to the Anchorage port terminal.

A concentrate trucking contractor will be responsible for loading the containers in the Arctic concentrate storage building using a wheeled loader. Containers would be loaded with net 27.2 WMT (30 SWT) of concentrate, resulting in 29.9 MT (33 Short Tons) gross weight per container. Based on a daily production of 1,513 WMT of concentrates, approximately 56 containers will be loaded per day and shipped from the Arctic site.

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The base for the trucking operation would be at Arctic Mine Junction (AMJ), located where the road to the Arctic Mine intersects the AMDIAP road. This would be the primary laydown yard for concentrate containers. The trucking contractor would have a maintenance facility at AMJ, and mobile maintenance trucks along the AMDIAP; however, most major equipment maintenance work is expected to be performed at the Contractor’s Fairbanks operating base. The AMJ pad would be initially constructed for the temporary construction camp.

The trucking of the containers to Fairbanks would be in three stages:

1. Load the containers at the Arctic mine concentrate loading building and then shuttle the container to a staging area at the AMJ area. The Contractor would supply and maintain the loader and forklifts at the Arctic site, and Trilogy Metals would have backups equipment available if required.

2. Truck the containers from AMJ on the AMDIAP road to the junction with the Dalton Highway using a B-train (tri-tri) configuration with each truck hauling two containers. At the Dalton Highway Junction, the driver would disconnect the two trailers and return to the mine with two trailers with empty containers. This would require approximately 26 trips per day. Drivers would be based at the Arctic site and would complete one trip per day and would return to sleep at the Arctic Mine at the end of their shift.

3. Fairbanks-based trucks and drivers would then move the trailers in a single trailer configuration from the Dalton Highway to a depot in Fairbanks.

In Fairbanks the containers would be loaded on railcars for transport to the Port of Alaska. At the port, the containers would be staged for direct loading into marine vessels using fixed shore cranes and a container rotator attachment. Concentrate would be shipped from Alaska in 10,000 dmt parcels.

Table 18-2provides details of the planned concentrate movement.

Table 18-2     Mode of Transport and Distances for Concentrate Shipping

Segment		Mode	Distance (km)	Trips/Day	Trips/Week
1	Arctic Site to AMJ	Truck – Single Trailer	19	56	389
2	AMJ to DHJ (via AMDIAP)	Truck – Double Trailer	322	28	195
3	DHJ to Fairbanks	Truck – Single Trailer	402	56	389
4	Fairbanks to Port of Alaska	Rail	573		3
5	Port of Alaska to Asian Port	Marine Bulk Carrier	9,000

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Concentrate shipping containers would be sourced from one of several suppliers and leased. Based on a production rate of 552,200 WMT of concentrates per year, and loading 10,000 DMT packages at the Port of Alaska, it is expected that a fleet of approximately 1,300 containers would be required.

The concept and cost development for the truck transport from the Arctic Mine to Fairbanks was developed by Lynden Logistics LLC, a company with experience in Alaska and Northern Canada as contract mineral concentrate truckers. Costs for the rail transport from Fairbanks to the Port of Alaska were provided by the Alaska Railroad Corporation. Development of concepts and costs for ocean transport of concentrates from Alaska to a typical Asian port was provided by Jim Vice, a consultant engaged by Trilogy Metals.

It is expected that the concentrate trucking would be integrated with the trucking of operating supplies and freight for the Arctic Mine, and most likely handled by a single contractor.

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19.0 Market Studies and Contracts

19.1 Metal Prices

Trilogy Metals provided Ausenco with the metal price projections for use in the Pre-Feasibility Study on which the Technical Report is based. Trilogy Metals established the pricing using a combination of two year trailing actual metal prices, and market research and bank analyst forward price projections, prepared in early 2018 by Jim Vice of StoneHouse Consulting Inc.

The long-term consensus metal price assumptions to be used in the Pre-Feasibility Study are:

· Copper: $3.00/lb.

· Lead: $1.00/lb.

· Zinc: $1.10/lb.

· Gold: $1,300/oz.

· Silver: $18.00/oz.

19.2 Markets and Contracts

No formal marketing studies have been completed for the concentrates proposed to be produced from the Trilogy deposit.

Assumptions as to concentrate quality are based on the results of inductively—coupled plasma (ICP) analysis of copper and lead concentrates produced from locked cycle tests at ALS Metallurgy and the zinc concentrate from locked cycle tests at SGS. The samples are thought to represent the expected concentrate quality. On the basis of these analysis, is has been assumed that the concentrates will be sent to an Asian port for smelting and refining.

No contracts have been entered into at the Report effective date for mining, concentrating, smelting, refining, transportation, handling, sales and hedging, and forward sales contracts or arrangements. It is expected that the sale of concentrate will include a mixture of long-term and spot contracts.

· Most concentrate is traded on the basis of term contracts. These frequently run for terms of one to 10 years, although many long-term contracts are treated as evergreen arrangements which continue indefinitely with periodic renegotiation of key terms and conditions. Generally, a term contract is a frame agreement under which a specified tonnage of material is shipped from mine to smelter, with charges renegotiated at regular intervals (typically annually).

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· Spot contracts are normally a one-off sale of a specific quantity of concentrate with a merchant or smelter. The material is paid for in much the same way as a concentrate shipped under a term contract. Merchant business is a mixture of one-off contracts with smelters and long-term contracts with both miners and smelters.

Often terms of sale for a term contract between miners and smelters are at “benchmark terms”, which is the consensus annual terms for the sale of concentrate, and negotiated annually. Spot sales are made at spot terms, negotiated on a contract by contract basis.

19.3 Smelter Term Assumptions

Smelter terms were applied for the delivery of copper, zinc and lead concentrate. It was assumed that delivery of all concentrates would be to an East Asian smelter at currently available freight rates. These terms are considered to be in line with current market conditions.

Copper

The contracts for the copper concentrate will generally include the following payment terms:

Copper: pay 96.5% of the content, subject to a minimum deduction of 1 unit, at the London Metal Exchange (LME) price for copper less a refining charge of $0.08 per payable pound.

Treatment charge: $80 per dmt of concentrates

Gold credit: If the gold grade is greater than 1 gr/dmt, then payment is 90% of the content less $5 per payable ounce.

Silver credit: If silver content is above 30 gr/dmt, then the payment is 90% of the content less a refining charge of $0.50 per payable ounce.

Penalty charge: $2.00 for each 1% that the zinc grade plus the lead grade exceeds 3%.

Zinc

The contracts for the zinc concentrate are assumed to generally include the following payment terms:

Zinc: pay 85% of content, subject to minimum deduction of eight units at the LME price

Treatment charge: $200/dmt of concentrate delivered; no price participation

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Gold credit: none

Silver credit: deduct 3 oz. /dmt from the Ag grade and pay for 70% of the balance

Penalty charge: none.

Lead

The contracts for the lead concentrate are assumed to include the following payment terms:

Lead: pay 95% of the content subject to a minimum deduction of three units.

Treatment charge: $180/dmt of concentrate delivered

Gold credit: pay 95% of content, subject to a minimum deduction of 1 g/dmt, less a refining charge of $10/payable oz.

Silver credit: pay 95% of content, subject to a minimum deduction of 50 g/t less a refining charge of $0.80/payable oz.

Penalty charge: $1.50 per each 0.1% that the arsenic grade exceeds 0.5% arsenic.

19.4 Transportation and Logistics

Transportation cost assumptions for the concentrate are summarized in the following table.

Table 19-1     Concentrate Transport Costs

Description	US$/dmt
Arctic Mine to Fairbanks	$175.84
Fairbanks to Port of Alaska	$29.32
Port Terminal & Handling	$21.74
Ocean Freight to Asian Port	$43.48
Total Transport Costs	$270.37

19.5 Insurance

An assumed insurance rate of 0.15% was applied to the recovered value of the concentrates less refining, smelting, penalties, and treatment charges.

19.6 Representation and Marketing

An allowance of $2.50/wmt of concentrate was applied as an allowance for marketing and representation.

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19.7 QP’s Opinion

The QP has reviewed the marketing, smelter terms, and transport costs, and considers them to be acceptable for use in financial modelling.

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20.0 Environmental Studies, permitting and social or community Impact

This section summarizes the existing environmental information for the Arctic Project area, describes the major mine permits that may be required to develop the project into a mine, and describes social and community considerations for the project.

20.1 Environmental Studies

The Arctic Project area includes the Ambler lowlands and Sub-Arctic Creek within the Shungnak River drainage. To date, a moderate amount of baseline environmental data collection has occurred in the area including surface water quality sampling, surface hydrology monitoring, wetlands mapping, stream flow monitoring, aquatic life surveys, avian and mammal habitat surveys, cultural resource surveys, hydrogeology studies, meteorological monitoring, and acid base accounting studies. The existing data are summarized in Sections 20.1.1 to 20.1.9.

20.1.1 Hydrology

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, Sub-Arctic Creek, Arctic Creek, and the Kogoluktuk River (Shaw, 2007, 2008, 2009).

In July 2010, Tetra Tech performed baseline studies to characterize flow and water quality in streams that could be potentially impacted by construction and operation of a proposed access road between the Bornite and Arctic airstrips, and the existing road between the Arctic airstrip and the Arctic deposit. Tetra Tech collected water quality and flow data at 14 sites (Tetra Tech, 2010a).

Two hydrologic gauging stations were installed on the Shungnak River (SRGS) and Subarctic Creek (SCGS) respectively by DOWL HKM in July, 2012. Each station is powered by dual solar panels and a battery, and continually measures and records water temperature, pH, conductivity, and water depth. A third hydrologic gauging station was established at the Lower Ruby Creek Gauging Site (RCDN) by WHPacific in June, 2013. The RCDN station was moved upstream in 2017 due to a beaver dam,

Trilogy Metals staff also performed instantaneous stream flow measurements and measured other standard field parameters (YSI 556 multi-parameter unit) during seasonal sampling events from 2013 to 2017 on Sub-Arctic Creek, Ruby Creek, the Shungnak River, and select tributaries to these drainages. The baseline water quality and hydrology program was expanded considerably in 2016 when Trilogy Metals increased the program from seven to 17 sites, adding sample sites in Cabin Creek, Riley Creek, Wesley Creek, and the Kogoluktuk River (Craig, 2016). The water quality and hydrology sites are shown on Figure 20-1.

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Figure 20-1     Current Water Quality and Hydrology Stations Location Map

In 2017, four groundwater monitoring wells were installed in the Sub-Arctic 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. Groundwater quality data have been incorporated into the site water and load balance model (Section 18.8.4). Data collection is on-going.

SRK (SRK Consulting, 2018) reviewed existing baseline hydrological data for the Project site and prepared a regional analysis of hydrology. The major findings are summarized as follows.

Site and publicly-available data were used to estimate MAP at the Project and to extend the available period of record. This analysis predicted a mean annual precipitation of 942 mm for Sub-Arctic Valley. This estimate of mean annual precipitation is greater than the estimate used in the 2013 Tetra Tech PEA. Evaporation was also predicted for the site using climatic records and an empirical relationship. The higher precipitation changes the site water balance from being in deficit to having a surplus of water.

Unit measured flows at nearby regional station Dahl Creek for the concurrent monitoring period of the Shungnak River and Sub-Arctic Creek were compared. Flows measured in Dahl Creek closely resemble unit flows at the Project, which supports the increased mean annual precipitation for Sub-Arctic Creek.

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20.1.2 Water Quality

Environmental baseline monitoring has been conducted in the area over the past seven years. The primary objectives of the baseline data are to characterize the natural environment of potentially impacted areas and identify reference locations for comparison throughout the Project life to assess impacts. The baseline monitoring data was supplemented with publicly available regional data to evaluate long-term trends.

In July 2010, Tetra Tech performed baseline studies to characterize flow and water quality in several streams 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. Water quality sampling was conducted by Trilogy from 2012 to the present. Small sampling programs were performed from 2012-2015 during the summer field season. The sampling program was expanded in 2016 and water quality samples to include more sample locations on the Shungnak River, Subarctic Creek, Ruby Creek, Riley Creek, Wesley Creek, and the Kogoluktuk River and sampling in the throughout the year. Several seeps in the Sub-Arctic Creek drainage near the Arctic Project have also been sampled.

Samples were analyzed for metals, mercury, cyanide, chloride, fluoride, nitrates, sulfate, acidity, alkalinity, total suspended solids, conductivity, pH, total dissolved solids, total organic carbon, and total phosphorus. The information will be used in the permitting and design of facilities related to exploration and potential mining at the Arctic Project. These data can be used to characterize baseline water quality conditions prior to any site disturbance and were used in the water and load balance model to make water quality predictions during mining and closure.

20.1.3 Wetlands Data

Tetra Tech 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 summarized as follows.

The area between the Bornite and Arctic airstrips consists of a wide valley containing the Ambler lowlands and the Shungnak River. Wetlands are prevalent throughout much of the Ambler lowlands. Most of the wetlands within the area occur within tundra vegetation 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.

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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 a 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.

In 2015, Trilogy Metals engaged DOWL to perform additional wetlands mapping and generate two preliminary wetlands determinations for a 5,910-acre study area (DOWL, 2016). The study area included the entire Sub-Arctic Creek drainage and most areas that could be directly impacted by the proposed Arctic open pit and mine facilities. The work determined that the broad study area is comprised of 715 acres of potentially jurisdictional wetlands, 40 acres of Waters of the United States WOTUS and 5,155 acres of non-jurisdictional uplands. According to DOWL (2016), the field work was performed in accordance with Part IV of the Corps of Engineers 1987 Wetlands Delineation Manual and the Regional Supplement to the Corps of Engineers Wetland Delineation Manual: Alaska Region (Version 2.0, 2007). Wetlands were classified and grouped according to the Class Level and system guidelines outlined in Classification of Wetlands and Deepwater Habitats of the United States (1979). The functional rating of potentially jurisdictional areas was determined using the criteria outlined in the 2009 Alaska Regulatory Guidance Letter, ID No. 09-10, the Cowardin Class, and observed hydrology. Ten ecological attributes were examined to subsequently rank wetland habitats as having low, moderate, or high functional ecological services. Riverine habitats (rivers and streams) perform vastly different functions compared to wetlands. Accordingly, riverine systems were evaluated based on the presence or absence of 17 functions according to the criteria outlined in the U.S. Department of the Interior Bureau of Land Management, Technical Report 1737-15, Riparian Area Management: A User Guide to Assessing Proper Functioning Condition and the Supporting Science for Lotic Areas.

20.1.4 Aquatic Life Data

Tetra Tech performed aquatic life studies in 2010 in the area between the Bornite and Arctic airstrips, and along the Arctic deposit road in Sub-Arctic Creek. The purpose of this study was to characterize the aquatic life within the Shungnak River and select 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 summarized as follows.

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 including Slimy sculpin, Alaska blackfish, Northern Pike, Round whitefish, Dolly Varden and Arctic grayling. 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 up to 9 m tall, situated downstream of the Arctic Project area.

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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 by Tetra Tech (2011). Based on these metrics, Tetra Tech (2011) concluded that six monitoring sites in Sub-Arctic Creek, Ruby Creek and an unnamed creek 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 suggest these streams exhibit lower levels of stream health and would likely show less sensitivity to effects from development.

In 2016 Trilogy Metals engaged Alaska Department of Fish and Game ADFG to complete aquatic studies as an extension of the work done by consultants in prior years. As illustrated in Figure 20-2, ADFG performed water quality sampling, periphyton sampling, fish tissue sampling, and aquatic invertebrate and fish surveys (minnow traps) at five sites in the Shungnak River drainage and one site on upper Riley Creek, a tributary of the Kogoluktuk River. In addition, they completed fish surveys using fyke nets at lower Ruby Creek and lower Sub-Arctic Creek.

Some of ADFG’s results show that Sub-Arctic Creek, which drains the Arctic deposit area contained some the lowest background concentrations of zinc and copper, as well as the lowest total dissolved solids (TDS), but the broadest range of TDS values, compared to other drainages (Bradley, 2017). Upper Subarctic Creek also had the highest number of aquatic invertebrates but also had the lowest species richness with a total of 11 taxa identified. Lower Sub-Arctic Creek had a significantly lower average number of aquatic invertebrates but more diversity than upper Sub-Arctic Creek.

ADFG retained a number of fish for element analysis. They note that the same species were not captured at every location, making direct comparisons between sample sites difficult. Additionally, sample sizes were low at some locations. However, results provide a good start for a baseline data set regarding metals concentrations in fish. Finally, Bradley (2017) concluded that despite being isolated from the Kobuk River by a large waterfall, the Shungnak drainage supports self-sustaining populations of Arctic grayling, Dolly Varden, round whitefish and slimy sculpin. Upper Sub-Arctic Creek was unique as the fish catches were dominated by Dolly Varden, and the catchment contained the highest density of aquatic insects. According to Bradley (2017) it is likely the Dolly Varden move into the upper reaches of Sub-Arctic Creek to feed on the abundant aquatic insects.

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Figure 20-2 shows the sites sampled during the 2016 aquatics survey.

Figure 20-2     2016 Aquatics Survey Sample Sites

20.1.5 Hydrogeology Data

In 2012, BGC Engineering (BGC) completed packer testing and installed vibrating wire piezometers in five drill holes. A total of 14 tests in five holes were completed for determination of hydraulic conductivity. Drill hole AR11-0129 was also instrumented with a shallow thermistor cable to monitor ground temperatures. Results from the drilling program were supported by a desk study to develop a conceptual hydrogeologic model. The following summary is an excerpt from BGC (2012). The 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 RQD intervals and/or in intervals with 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 range 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 indicate that groundwater flow is topographically driven by recharge in the uplands and discharge in the valley bottom. In permafrost terrain, groundwater recharge is limited by the frozen ground which acts as an aquitard. BGC concluded that in order to develop a groundwater flow model for the Project, it would be necessary to understand the distribution of permafrost and talik zones in the project area to estimate recharge rates and delineate the areas where recharge occurs.

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In 2015 and 2016, SRK conducted hydrogeologic work including instrumenting drill holes and collecting hydrogeological information and compiled the data to generate a conceptual model of the groundwater flow system in the vicinity of the Arctic Deposit. SRK’s field work included drill monitoring, packer-based hydraulic testing (water injection and airlift tests), installation of vibrating wire and standpipe piezometers, measuring ground temperatures and performing a seepage and spring survey. The data from this work were analyzed to characterize the hydrological groundwater regime in the Arctic Deposit area. SRK (2016) interpreted the data to conclude that the Arctic deposit may be permafrost-free except for patches near the ridge top, and measured hydraulic conductivity in bedrock ranging from 3 x 10-9 and 3 x 10-6 m/s, and interpreted the conductivity to be primarily related to fractures. SRK also observed water level effects in adjacent drill holes up to 6m away and interpreted this to suggest that fault zones and/or a localized, well- connected fracture network with higher hydraulic conductivity may be present. This interpretation is quite similar to BGC’s 2012 conclusion that higher inflow rates from some fault zones should be anticipated.

Groundwater level data from piezometers led SRK (2016) to suggest that the groundwater system in the Arctic Deposit is compartmentalized and that at least one perched system exists, but concluded that the root causes for the compartmentalization remain uncertain; talc layers are considered the most likely cause. SRK also observed seasonal variations in groundwater levels indicating an increase during spring freshet, a second increase in summer, and decreasing water levels in fall and winter.

SRK (2016) concluded that the Arctic Deposit has three potential aquitards and three potential aquifers. Aquifers include permeable overburden, fractured bedrock and fault conduits. They summarized the groundwater flow as being topographically driven and recharged by a range of 5 to 10% of the mean annual precipitation which is variably interpreted to be 395 to 949 mm/yr. (Tetra Tech, 2013; BGC, 2012; SRK, 2016 and SRK 2018 (personal communication)).

Finally, SRK (2016) proposed two conceptual groundwater models, with one incorporating upper and lower water tables possibly separated by a steep fault, and a second model that includes a single water table with a high conductivity zone that depresses the water table in the approximate middle of the proposed mine pit as it was defined in 2016.

20.1.6 Cultural Resources Data

In 2016, Trilogy engaged consultant WHPacific to perform a cultural resource assessment of the Arctic Project area. WHPacific (2016) noted that all 2,327 acres of the survey area were flown by helicopter at low elevation for observation by the archaeologists. The result of this flyover was the determination that the majority of the Project area had a low probability of containing cultural resources and had very low surface visibility. Of the total Project acreage, 530 acres were traversed on foot. These areas included the lower valley slopes, ridges, flat areas overlooking valleys, and terraces along the waterways. No cultural resources were found in the survey a rea. As part of their work WHPacific also completed a literature review, archival research and held stakeholder meetings in the communities of Shungnak and Kobuk. The results of that work include reviewing confidential Alaska Heritage Resource Survey Site information and is not discussed here except to say that no archaeologic sites have been recorded in the Project area (WHPacific, 2016). Late 20^th century mining exploration is in evidence in the Project area as seen by roads, abandoned equipment, and the airfield (outside of the direct survey area). These are of an age that is on the cusp of being considered historical period resources (50 years or older) according to the National Historic Preservation Act. These resources are not unique within northern Alaska or for 20^th century mining materials.

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Local community members communicated that the region was not one that was used by local residents in the past due to its lack of resources and passages to areas north, west, and east, where other resources and trading opportunities existed.

WHPacific (2016) recommended no further cultural resources work and that Project work associated with the proposed Arctic pit, facilities, tailings, and access road corridor project should proceed as planned.

20.1.7 Subsistence Data

Access to the Arctic Project area includes travel over private lands owned by NANA Regional Corporation. Trilogy Metals acknowledges the importance of subsistence to local residents, and as a result, a Subsistence 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. Representatives from NANA and Trilogy Metals facilitate the meetings and report a summary of the discussions and recommendations provided by the Subsistence Committee to the Oversight Committee, as defined by the “NANA/NovaCopper Exploration Agreement” (2011).

A formal subsistence survey has not been performed in the immediate vicinity; however, Trilogy Metals 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 the 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 access road project, and identify what, if any, additional information (i.e., data gaps) needed to be collected 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 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 Trilogy Metals 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 ADFG also has information about caribou population density and migration patterns in the area (ADFG, 2011).

The Northwest Arctic Borough (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, Trilogy metals has communicated with the ADFG wildlife biologists, who monitor caribou herd movements in the spring and fall in proximity to the Arctic Project by using radio-collared caribou. 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 for the Western Arctic caribou herd. The nearest herd is approximately 48 km west of the Arctic Project area. According to ADFG, (2015) caribou herd population numbers for this herd are decreasing and have decreased from approximately 325,000 caribou in 2011 to 235,000 in 2013, triggering a change from liberal to conservative management levels (ADFG, 2015).

Trilogy Metals reports (Cal Craig, personal communication, 2017) that caribou are rarely seen in the Project area (during the summer field season) and only in small numbers when they are. ADFG’s migration maps illustrate that the migration routes for the Western Arctic and Porcupine herds are to the west and east of the Project area, respectively.

DOWL (2016) performed a large mammal habitat survey in the Project area. They include historic maps of caribou migration in their reports that also show that the Project area is outside of main corridor routes and calving areas, but that data from 1988-2007 suggested that the area may be used for wintering habitat.

20.1.8 Endangered Species, Migratory Birds, and Bald and Golden Eagle protection

In 2016, Trilogy Metals engaged WHPacific, through subcontractor ABR, Inc., to perform aerial surveys of nesting raptors in the Project area, including the Bornite area located some 24 km southwest of the Arctic Project area. ABR (2017) identified a total of 26 nests; 18 were in the Bornite area and eight were in the Arctic area. Fifteen of the total were occupied in the initial occupancy survey; nine were occupied by Rough-legged hawks, with three Peregrine Falcon nests and three Raven nests. In the later productivity survey ABR observed that only one Rough-legged Hawk nest had a (single) nestling, one Peregrine Falcon nest had two young and an unhatched egg, and two Raven nests had young (not counted).

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In 2017, Trilogy engaged WHPacific to review requirements that would be necessary to comply with the Endangered Species Act, the Migratory Bird Treaty Act and the Bald and Golden Eagle Protection Act. WHPacific (2017) concluded that there are no endangered species or critical habitat in the Project area. Further they reported that nine avian species of conservation concern are expected to occur or could potentially be affected by activities in the project area. They provided the timing guidelines for vegetation clearing that are meant to protect these species during nesting activities and advised that if impacts to migratory species are unavoidable for the project that the US Fish and Wildlife Service (USFW) must be consulted. They also recommended that Trilogy Metals request a project review from USFW when the Project is closer to initiation

20.1.9 Acid Base Accounting Data

Sampling efforts have been used to characterize the acid generation potential of the mine waste for the Arctic Project. In 1998, Robertson collected 60 representative core samples from the deposit for their acid/base characteristics; these samples provided a broad assessment of acid rock drainage (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 sulfide mineralogy in assessing ARD. The criteria used for classifying ARD potential also differs slightly from the Robertson era work.

Trilogy Metals retained SRK to provide on-going ML/ARD characterization services for the Arctic Project. Activities in 2016 focused on three objectives: 1) on-going monitoring of on-site barrel tests (kinetics), 2) on-going monitoring of parallel laboratory humidity cell tests (kinetics), and 3) expansion of the current acid-base-accounting (ABA) database (statics). Barrel test samples were routinely collected during 2016 and 2017 and analyzed by ARS Aleut Analytical. Humidity cell tests, initiated in 2015, were monitored on a weekly basis by Maxxam Analytics of Burnaby, British Columbia. Both the barrel test work and humidity cell test work is on-going. Trilogy Metals and SRK selected 1,119 samples to be analyzed for a conventional static ABA package with a trace element scan using the same methods as the exploration database. Samples were analyzed by Global ARD Testing Services of Burnaby, British Columbia. Upon completion of the laboratory test work, SRK began evaluating the use of proxies to support next steps in regards to the block modelling of ML/ARD potential. This work is on-going.

20.1.10 Additional Baseline Data requirements

In general, the baseline data for the Arctic Project is sufficient for the current stage of the project. Additional environmental data may be required to support future detailed mine design, development of an EIS, permitting, construction and operations. Trilogy Metals will continue to engage the regulatory agencies and project consultants to identify any additional data needs as the project design advances.

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20.2 Permitting

20.2.1 Exploration Permits

Trilogy Metals performs mineral exploration at the Arctic Deposit under State of Alaska and NWAB permits.

Trilogy Metals is presently operating under a State of Alaska Miscellaneous Land Use Permit (APMA permit) that expires at 2017 year-end, and an application to renew will be submitted by Trilogy to ADNR in early 2018. Cumulative surface disturbance for exploration activities on the Arctic Project remains less than 5 acres (excluding historic disturbance that includes roads and camp disturbances) and therefore there are currently no State requirements for reclamation bonding for the Arctic Project.

Trilogy metals reports that the NWAB Title 9 Conditional Use Permit authorizing exploration and bulk fuel storage expires on December 31, 2018, and the permit for landfill, gravel extraction, and bridge construction expires on December 31, 2019. No bonding is required for the borough permits.

Trilogy Metals has obtained several other permits for camp-support operations. These permits include a drinking water permit, a domestic wastewater discharge permit, 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 authorizations have been issued by the ADNR, a Title 16 Fish Habitat permit, and a wildlife hazing permit have been issued by the ADFG.

20.2.2 Major Mine Permits

The following discussion identifies the major permits and approvals that will likely be required for the Arctic Deposit to be developed into an operating mine.

Permits would be issued from Federal, State, and Regional agencies, including: the US Army Corps of Engineers (USACE), the Alaska Department of Environmental Conservation (ADEC), the Alaska Department of Fish and Game (ADFG), the Alaska Department of Natural Resources (ADNR), and the Northwest Arctic Borough (NWAB).  The State of Alaska permit for exploration on the project, the Annual Hardrock Exploration Activity (AHEA) Permit, is obtained and renewed every five years through the ADNR – Division of Mining, Land and Water.

The types of major mine permits required by this project are largely driven by the underlying land ownership; regulatory authorities vary depending on land ownership. The Arctic Project area includes patented mining claims (private land under separate ownership by Trilogy and NANA), State of Alaska land, and NANA land (private land). The mine pit would situate mostly on patented land while the mill, tailings and waste rock facilities would be largely on State land. Other facilities, such as the camp, would be on NANA land. Federal land would likely underlie 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 the State of Alaska and are not addressed in this report. A list of likely major mine permits is included in Table 20-1.

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Because the Arctic Project is situated to a large extent on State land, it will likely be necessary to obtain a Plan of Operation Approval (which includes the Reclamation Plan) from the ADNR. The Project will also require certificates to construct and then operate a dam(s) (tailings and water storage) from the ADNR (Dam Safety Unit) as well as water use authorizations, an upland mining lease and a mill site lease, as well as several minor permits including those that authorize access to construction material sites from ADNR.

The ADEC would authorize waste management under an integrated waste management permit, air emissions during construction and then operations under an air permit, and require an 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 review the USACE Section 404 permit to certify that it complies with Section 401 of the CWA.

The ADFG would have to authorize any culverts or bridges that are required to cross fish-bearing streams or other impacts to fish-bearing streams that result in the loss of fish habitat.

The USACE would require a CWA Section 404 permit for dredging and filling activities in WOTUS including jurisdictional wetlands.

The USACE Section 404 permitting action would require the USACE to comply with 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.

As part of the Section 404 permitting process, the Arctic Project will have to meet USACE wetlands guidelines to avoid, minimize and mitigate impacts to wetlands. The USACE will require Trilogy Metals to develop a compensatory wetlands mitigation plan for mitigating unavoidable wetlands impacts.

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 404permit application to the ACOE. The timeline includes the development and publication of a draft and final EIS and ends with a Record of Decision, and 404-permit issuance. 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 require between 1.5 to 3 years to complete, and could potentially take longer.

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Table 20-1     Major Mine Permits Required for the Arctic Project

Agency	Authorization
State of Alaska
ADNR	Plan of Operations Approval (including Reclamation 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
ADFG	Title 16 Permits for Fish Passage (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 generally define critical permitting path. Additional “minor” permits are also required.

20.3 Social or Community Considerations

The Arctic Project is located approximately 40 km northeast of the native villages of Shungnak and Kobuk, and 64 km east-northeast of the native village of Ambler. The population in these villages range from 156 in Kobuk (2016 Census) to 262 in Shungnak (2016 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. Trilogy Metals is working directly with the villages to employ residents in the ongoing exploration program as geotechnicians, drill helpers, and environmental technicians. Trilogy Metals and NANA have established a Workforce Development Committee, described below, to assist with developing a local workforce. In addition, Trilogy Metals has existing contracts with native-affiliated companies (such as NANA Management Services and WHPacific Inc.) that are providing camp catering and environmental services for the project, respectively.

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In October 2011, NovaCopper (now Trilogy Metals) signed a cooperative 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-affiliated consultants and contractors. Furthermore, the agreement formalized the Subsistence Committee to protect subsistence and the Iñupiaq way of life and an Oversight Committee, with equal representation from Trilogy Metals and NANA, to regularly review project plans and activities. The Workforce Development Committee also addresses 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 Trilogy Metals that promotes education for youth in the region. Trilogy Metals generally meets monthly, during summer months, with the residents of Kobuk, Shungnak and Ambler, the three villages closest to the Project area. Trilogy Metals also generally meets annually with eight other NANA region villages including Noatak, Kivalina, Kotzebue, Kiana, Deering, Buckland, Selawik and Noorvik, for updating residents on project plans and fielding their questions and concerns.

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. Trilogy Metals acknowledged these concerns and is taking substantive steps to address them during the current exploration stage of the Project.

Local community concerns will also be formally recognized during the development of the project EIS. Early in the EIS process, the lead federal permitting agency will hold scoping meetings in rural villages to hear and record the concerns of the local communities so that the more significant of these concerns can be addressed during the development of the EIS. In addition, the lead federal agency would have government-to-government consultations with the Tribal Councils in each of the villages, as part of the EIS process, to discuss the project and hear 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 if regional interests are met. Regional engagement by Trilogy Metals has also encountered a strong desire for the economic benefits that come with mining projects.

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20.4 Mine Reclamation and Closure

Mine reclamation and closure considerations are largely driven by State regulations (11 AAC 86.150, 11 AAC 97.100-910, and 18 AAC 70) and statutes (AS 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 from acid rock drainage or metal leaching on the site. A detailed reclamation plan will be submitted to the State 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 land, 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 Plan. In addition, private land owners must formally concur with the portion of the reclamation plan for their lands so that it is compatible with their intended post-mining land use. Generalized reclamation and closure strategies are presented in Section 20.4.1.

20.4.1 Reclamation and Closure Plan

A final reclamation plan for the Arctic Project will be developed as part of the formal mine permitting process in the future. A preliminary plan was developed by SRK for this PFS. The following is a general discussion of that preliminary reclamation plan.

Closure Objectives and Closure Criteria

The overall closure objective is to establish stable chemical and physical conditions that protect the environment and human health. To the extent practicable, rehabilitation efforts will endeavor to return the site to a condition which generally conforms with the surrounding terrain. The site will be monitored and maintained post-closure in order to demonstrably meet these conditions.

The following general closure objectives were considered:

· Demolish and remove all construction, camp and industrial facilities and reclamation of affected footprints;

· Achieve long-term slope stability of the pit, waste rock dump, and tailings facility;

· Meet water quality criteria for all mine water and seeps prior to discharge to the environment;

· Prevent intrusion and migration of tailings porewater and water from the pit into the regional groundwater;

· Prevent contact of humans and wildlife with the mine waste (waste rock and tailings);

· Establish adequate vegetation density to ensure erosion protection of the soil slopes; and

· Re-establish vegetation on areas returned to normal land use.

Closure Activities

Closure activities will be undertaken at the end of mine life to bring the mine facilities in a state consistent with the stated closure objectives and compliant with the regulations for closure and abandonment. Major activities planned for the various mine components and facilities are detailed as follows.

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Open Pit Workings

It is expected that the pit wall will be geotechnically stable in the long-term, and therefore no in-pit work is required at closure. The perimeter of the pit will be fenced in and clear signage will be installed to warn of existing dangers.

Once operations are completed, the water in the pit will be allowed to rise and a pit lake will be created. At closure runoff from the waste rock will report to the mine pit and accumulate there. Water from the pit lake will be treated seasonally and discharged to the environment. The pit lake will not be allowed to overflow; however, an emergency spillway will be constructed as a matter of best practice.

Waste Rock Dumps and Tailings Management Facility

The waste rock will all be managed as PAG and permanently placed between the mine pit and the tailings impoundment: all constructed up-gradient from the mine pit. The face of the waste rock dump will be regraded at closure for long-term stability. Efforts will be made to create an overall convex-concave shape on the dump face, which is the predominant shape of natural slopes in mature landscapes in the area. The entire WRD will then be covered with 0.3 m of growth medium salvaged prior to construction of the dump. The growth medium cover will be revegetated using a seed mix of native species.

Under the present mine plan, the tailings will be conveyed to a lined tailings facility located up-gradient from the open pit and the waste rock storage facility. The exposed tailings areas will be capped with a 0.3 m soil cover and revegetated using a seed mix of native species. A portion of the former supernatant pond can be backfilled with waste rock to create the needed surface grades for run-off. An engineered spillway will be created to ensure that water level will be maintained in the TMF such that the tailings remain saturated to eliminate oxidation and associated ARD/ML.

Material for covering the WRD and TMF will be taken from material stored in the stockpiles. After the material is used for reclamation of the waste facility the stockpile footprint and sedimentation ponds will be graded and revegetated.

Buildings and Equipment

The water treatment plant (WTP) will be left in place, together with all appurtenant facilities and utilities. All other steel frame buildings including the camp facilities, the mill, the truck shop, and the conveyors will be demolished selectively by removing the roof and siding and then dismantling the steel frames and trusses. Resulting debris will be disposed of in the landfill. Concrete walls, pillars, and beams will be demolished to the ground and concrete foundations will be covered in place. Controlled blasting may be used to help with the demolition. All sumps and cavities will be backfilled to ground level. The rock fill pads underlying all buildings and equipment will be re-graded to prevent permanent ponding.

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Non-hazardous and hazardous waste will be segregated. Hazardous waste will be placed in suitable containers and hauled to a licensed disposal facility, while non-hazardous waste will be placed in the landfill.

Any unwanted mobile or stationary equipment will be stripped of electronics and batteries, drained of all fluids (fuels, lubricants, coolants), decontaminated by power washing, and placed in the landfill for final disposal.

Mine Infrastructure

Mine support infrastructure is comprised of internal access roads, haul roads, and rock fill pads underlying the site buildings and facilities.

All bridges and culverts associated with the haul roads and internal access roads will be removed and natural drainage will be restored. Swales will be created where needed, to allow continued use of the roads into post-closure water treatment, monitoring, and maintenance. Roads that are not needed in post-closure will be ripped and re-vegetated.

The surface of the rockfill pads underlying some of the buildings and facilities on site will be re-graded and/or crowned as necessary to prevent ponding of water. The pads will be covered with 0.3 m of overburden and re-vegetated.

The site access road will be maintained as long as water treatment is occurring on site.

Landfill

An unlined non-hazardous landfill will be located in the WRD. Demolition waste and other non-hazardous waste will be placed in the landfill and consolidated to minimize the occupied volume. The waste will then be covered with at least 1 m of waste rock. The final surface will then be graded to prevent permanent ponding and will be covered similarly to the rest of the waste rock dump with a 0.3 m thick soil cover.

Water Management System

The water managementsystem consisting of the pit lake, the waste rock run-off collection pond, the tailings reclaim water pond, interception ditches, and various pipelines will be largely decommissioned. The dams and the footprint of the run-off management pond will be re-vegetated, while the sediment ponds will be backfilled. New diversion channels and interception ditches will be created to manage surface runoff on and around the TMF and the WRD, as well as intake and discharge pipelines between the pit and the WTP. Runoff from the WRD will report to the open pit and, as the pit fills naturally, water levels will be managed by treating and discharging treated water for the long-term

Water Treatment Plant

A WTP will be constructed during mine operations to treat mine waters from the pit lake and excess tailings supernatant water prior to discharge to the environment. Long-term water treatment will be required, possibly in perpetuity.

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Treatment of the excess tailings porewater for ammonia will cease five years after closure. The WTP will be operated on a seasonal basis, in the warm months of the year (May to September) and the treated water will be discharged into the Shungnak River. Sludges produced in the treatment process will be pumped back into the deeper levels of the pit lake.

Post-closure

Continued presence at site will be required in the post-closure period for as long as water treatment is necessary. This includes maintaining road access, a seasonal camp, and the water discharge pipeline to Shungnak River. Generators and mobile equipment will also be required for the WTP operations as well as sampling and monitoring activities.

The site-wide water balance will be updated and improved for final mine permitting to predict water shortfalls or surpluses at closure. The details of that updated water balance will help identify the need for any water use authorizations, surface water discharge permits and establish water treatment capabilities for the Arctic Project at mine closure.

In the short-term (up to 10 years following closure) monitoring to confirm that the closure objectives are met will be based on the following requirements:

· The site should be visually inspected by a Professional Engineer annually for three consecutive years and less frequent thereafter for up to 10 years to ensure that erosion-prone areas have stabilized.

· The soil covers over the WRD and the TMF should be regularly inspected by a qualified inspector to ensure the physical integrity of the cover is maintained. Inspection intervals should be by-annually for the first 10 years after construction.

· The site should be inspected by a vegetation specialist to confirm suitability of the revegetation efforts. Inspections should be completed at the following intervals, unless otherwise recommended by the vegetation expert: Year 1, Year 3, Year 7 and Year 10 post-closure.

In the long-term, continuous water quality monitoring will be implemented at sampling frequencies prescribed by futuredischarge permits.

Maintenance will be performed on areas that monitoring identifies as needing repairs. Water treatment in perpetuity will likely require a well-defined set of discharge quality criteria to be determined at the later stages of permitting and design.

Closure Cost Estimate

The estimate cost of closure is based on unit rates used by SRK on other closure projects in cold environments. The indirect costs were included as percentages of the estimated direct costs based on guidelines for Alaska (DOWL 2015).

Long-term water treatment and maintenance of certain water management facilities were calculated separately, and an NPV value is provided for the first 200 years, at a discount rate of 4.3%.

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Reclamation costs have been estimated to be $65.3 million for this PFS, in 2017 undiscounted US dollars. Annual costs associated with long-term operations of the WTP are estimated to be about $1.27 million for the first five years and $0.96 million thereafter. A summary of estimated costs is provided in Table 20-2.

Table 20-2     Summary of Closure and Reclamation Costs

Cost Items	Subtotals Closure Costs (rounded to nearest thousand - USD)
DIRECT CLOSURE COSTS
Open Pit Mine Workings	$	142,400
Waste Rock and Overburden Piles	$	9,295,200
Tailings Management Facility	$	4,041,600
Buildings and Equipment	$	1,402,400
Roads	$	328,800
Landfill and Waste Disposal Area	$	12,000
Water Management Systems	$	2,340,800
Subtotal Direct Closure Cost	$	17,563,000
INDIRECT CLOSURE COSTS
Mob/demob	$	351,200
Contingency	$	7,024,800
Engineering Redesign	$	878,400
Contract Administration	$	1,229,600
Contractor Profit & Overhead	$	2,634,400
Performance Bond	$	527,200
Liability Insurance	$	88,000
Subtotal Indirect Closure Cost	$	12,734,000
Total Closure Cost	$	30,297,000
POST-CLOSURE COSTS (200 years)
Short-term Post-closure Monitoring	$	408,000
Water Treatment	$	283,786,400
Confirmatory Sampling	$	4,176,000
Reporting	$	6,400,000
Subtotal Post-Closure Costs (undiscounted)	$	294,770,000
Post-closure Costs NPV	$	35,050,000
Total Discounted Project Cost	$	65,347,000

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20.4.2 Reclamation and Closure Financial Assurance

In the absence of activities on Federal Land in the mine area and excluding the access road to the district from the Dalton Highway, there would not be any financial assurance requirements from the Federal government for the mine.

There are three 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 plan be submitted prior to mine development and requires financial assurance, typically prior to construction, to assure reclamation of the site. The ADNR Dam Safety Unit also requires a financial assurance sufficient to cover the cost of decommissioning dams or the cost for long term maintenance and monitoring of dams that will remain in-place. 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 costs, monitoring, and maintenance needs. The State requires that the financial assurance amount also include the property holding costs for a one-year period.

The final financial assurance amount will be calculated through the process of reviewing and approving the Arctic Project reclamation plan during the formal permitting process. In general, the approach is to combine the reclamation costs, post-closure monitoring costs and the long-term annual water treatment costs into a financial amount that includes deriving the NPV of the long-term costs and combining that with the reclamation cost.

Trilogy Metals may satisfy the State 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) for the dam- or ADEC-related obligation - any other form of financial assurance that meets the financial test or other conditions set in regulation by the ADNR or ADEC commissioners.

The adequacy of the reclamation 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 plan costs.

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21.0 Capital and Operating Costs

21.1 Capital Cost Summary

21.1.1 Capital Cost Introduction

The objective of the Arctic Project PFS was to develop a capital cost estimate with an accuracy of -20% to +30%, in accordance with Ausenco’s standards for a pre-feasibility study. This includes the cost to complete the design, procurement, construction and commissioning, of all of the facilities

This estimate collectively presents the entire costs for the project including the Amec Foster Wheeler mining scope, SRK scope, Owner’s scope (Trilogy Metals Inc.) and Ausenco’s scope.

The physical facilities and utilities for the Arctic Project include but are not limited to the following areas:

· Ausenco

o Process plant

o Surface infrastructure and ancillaries

o Off plant roads (by AllNorth)

o Existing airstrip upgrade (DOWL)

· Trilogy Metals

o Owners’ costs

· SRK

o TMF

o Waste rock facility

o WTP

o Site water management

· Amec Foster Wheeler

o Mine development

o Construction of mine roads

o Explosives magazine

o Mining fleet

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21.1.2 Project Execution

The estimate has been based on the traditional engineering, procurement and construction management (EPCM) approach where the EPCM contractor will oversee the delivery of the completed project from detailed engineering and procurement to handover of working facility. The EPCM contractor would engage and coordinate several subcontractors to complete all work within the given scopes. Typical vertical and/or horizontal contract packages are identified and aligned with different pricing models such as, but not limited to:

· Schedule of rates (unit price)

This contract pricing model is based on estimated quantities of items included in the scope and their unit prices. The final contract price is dependent on the quantities needed to complete the work under the contract

· Time and materials

Time and materials (T&M) fixes rates for labour and material expenditures, with the contractor paid on the basis of actual labour hours (time), usually at specified hourly rates, actual cost of materials and equipment usage, and an agreed upon fixed add-on to cover the contractor’s overheads and profit

· Design and construct

With this option one entity will provide design and construction services for an awarded scope of work. A higher degree of price certainty can be achieved when a lump sum arrangement is used; this method also provides a single point of accountability and an improved integration of the design with construction.

21.1.3 Work Breakdown Structure

The estimate has been arranged by major area, area, major facility, and facility. Each sub-area has been further broken down into disciplines such as earthworks, concrete etc. Each discipline line item is defined into resources such as labour, materials, equipment, etc., so that each line comprises all the elements required in each task.

The work breakdown structure (WBS) has been developed in sufficient detail to provide the required level of confidence and accuracy and also to provide the basis for further development as the project moves into execution phase.

21.1.4 Estimate Summary

The estimate is derived from a number of fundamental assumptions as shown on the process flow diagrams (PFDs), drawings, scope definition and WBS, it includes all associated infrastructure as defined within the scope of works.

The capital cost estimate has been summarized at the levels indicated by the following tables and stated in United States dollars (US$) with a base date of 4th quarter 2017 and with no provision for forward escalation.

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Table 21-1     Estimate Summary Level 1 Major Facility

Cost Type	Description	US$M
Direct	Mine	281.1
	Crushing	18.3
	Process	113.8
	Tailings	30.3
	On-Site Infrastructure	84.5
	Off-Site Infrastructure	15.6
	Direct Subtotal	543.8
Indirect	Indirects	121.9
	Contingency	92.0
	Owners Costs	21.9
	Indirect Total	235.8
Project Total		779.6

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Table 21-2     Initial Estimate by Major Discipline

Description	US$M
Site Development	2.9
Earthworks	6.1
Concrete	27.3
Structural Steel	7.4
Architectural	23.8
Platework	5.9
Mechanical Equipment	61.4
Mobile Equipment	8.1
Piping	26.3
Electrical Equipment	35.3
Electrical Bulks	6.4
Instrumentation	2.9
Third Party Estimates	327.6
Subtotal Direct Costs	541.2
Field Indirects	77.6
Spares & First Fills	5.4
Vendors	2.8
EPCM	38.7
Contingency	92.0
Owners Costs	21.9
Indirect Costs	238.3
PROJECT TOTAL	779.6

21.1.5 Definition

Definition of Costs

The estimate is broken out into direct and indirect initial capital.

Initial capital includes all costs incurred during the pre-production period.

Sustaining capital is the capital cost associated with the periodic addition of new plant, equipment or services that are required to maintain production and operations at their existing levels

Direct costs are those costs that pertain to the permanent equipment, materials and labour associated with the physical construction of the process facility, infrastructure, utilities, buildings, etc. Contractor’s indirect costs are contained within each discipline’s all-in rates.

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Indirect costs include all costs associated with implementation of the plant and incurred by the owner, engineer or consultants in the design, procurement, construction, and commissioning of the project.

General Methodology

The estimate is developed based on a mix of material take-offs and factored quantities and costs, semi-detailed unit costs and defined work packages for major equipment supply.

The structure of the estimate is a build-up of the direct and indirect cost of the current quantities; this includes the installation/construction hours, unit labour rates and contractor distributable costs, bulk and miscellaneous material and equipment costs, any subcontractor costs, freight and growth.

Exchange Rates

Trilogy Metals provided Ausenco with the exchange rate projections for use in the Pre-Feasibility Study on which the Technical Report is based. Trilogy Metals established the projections using information sourced from the XE website as at August 26, 2017.

The following exchange rates are to be used to develop the capital cost estimate for the Pre-Feasibility Study on which the Technical Report is based.

Table 21-3     Estimate Exchange Rates

Forex Rate	US$
AUD 0.74	1
EUR 1.20	1
CAD 0.80	1
GBP 1.29	1

Notes: AUD = Australian Dollar, EUR = Euro, CAD = Canadian Dollar, GBP = Great British Pound

Market Availability

The pricing and delivery information for quoted equipment, material and services was provided by suppliers based on the market conditions and expectations applicable at the time of developing the estimate.

The market conditions are susceptible to the impact of demand and availability at the time of purchase and could result in variations in the supply conditions. The estimate in this report is based on information provided by suppliers and assumes there are no problems associated with the supply and availability of equipment and services during the execution phase.

21.1.6 Road Construction

The total estimate for the main site access road is $14.6M. The access road includes approximately 8km of road between Bornite and the Arctic intersection. The estimate includes the access road design and construction package, stream crossings and drainage structures, road surfacing, project delivery costs, and contingency.

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21.1.7 Basis of Mining Capital Cost Estimate

The scope of the mining cost estimate includes the purchase of initial mining fleet, maintenance, and mine support equipment; wages for hourly and salary personnel for pre-production mine operation; haul road construction; and miscellaneous equipment.

Estimates for mining equipment were based either on mining fleet equipment schedules and equipment pricing provided by vendors for supply, delivery, assembly, and testing or from historical data.

Mining quantities were derived from first principles and mine-phased planning to achieve the planned production rates. Mining excavation estimates were based on geological studies, mine models, drawings, and sketches. Costs include pre-production stripping and haul road construction by the mining fleet.

Fuel consumption was estimated from vendor-supplied data for each type of equipment and equipment utilization factors.

21.1.8 Mining Capital and Sustaining Capital Costs

Total mine capital costs estimated to develop the Arctic mine are $281.2 M (Table 21-4). The capital costs are inclusive of $158.6 M in preproduction operating costs for a two-year preproduction period and $115.7 M in initial capital expenditures primarily for mobile equipment. Included within the initial capital costs are $6.7 M in initial equipment spares and $1.0 M for shop tooling. Following preproduction capital spending, an additional $45.0 M is required primarily for additions to the CAT 777 mining fleet, and to replace equipment beyond its manufacturer-recommended service life.

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Table 21-4    Mine Capital Costs

Cost Area	US$ 000’s
Pre-Production Mining	158,642
Engineering and Management	10,239
Mining Infrastructure and Earthworks	6,900
Drilling	8,856
Blasting	348
Loading	17,657
Hauling	55,620
Support	18,960
Dewatering	156
Maintenance	3,808
Total Initial Mine Capital	281,186
Sustaining Capital	45,048
Total Mine Capital and Sustaining Capital	326,234

21.1.9 Tailings Management Facility

The capital cost estimate for the TMF makes provision for constructing the initial starter dam of the TMF to an elevation of 830 m, which is sufficient to store the first year of tailings production and enough water to provide a 100 days residence for the mill reclaim water. The tailings dam would be constructed using waste rock material for the adjacent WRD and compacted in 1m lifts. As the rock is already being delivered to the WRD, the only cost included for placement in the estimate is to cover the incremental compaction costs. No allowance was provided for spreading the material as it is assumed that the dozers already on the WRD will handle that activity.

An allowance has also been made for excavating the overburden encountered beneath the starter dam footprint. This material would be either stockpiled or used in construction of the WRCP. Costs were also estimated for the general foundation prep within the footprint of the tailings impoundment in advance of the liner placement. Supply and installation of the HDPE geomembrane was only considered in the estimate for placement up to the starter dam elevation of 830 m.

The storage capacity of the TMF will be increased through three additional raises of the dam in years 1, 5 and 10 to an ultimate elevation of 890 m. Sustaining capital has been estimated for each of these raises to accommodate compaction of the waste rock in the compacted dam zone as well as placement of the geomembrane.

21.1.10 Water Treatment Plant

Capital cost estimate for the high-density sludge lime treatment plant and the ammonia/cyanide oxidation treatment plant were prepared. Costs for major equipment items were based on quotes collected in the past year. Costs for other items to construct the plants were scaled from the total costs for major equipment. The cost estimate includes a 2,000 m² heated building to house both the high-density sludge lime treatment plant and the ammonia/cyanide oxidation plant.

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The influent annual average flow rate to each plant was increased by 20% to add a level of conservativism to the cost estimate. A 25% contingency was applied to the capital cost estimate and a 5% escalation factor was applied to account for uncosted, undesigned or omitted items.

21.1.11 Sustaining Capital and Closure Costs Summary

Other non-mining sustaining capital costs include expenditure related to the tailings management facility, and an allowance for miscellaneous general and administrative (G&A) sustaining capital costs. Table 21-5 outlines these expenditures.

Table 21-5    Sustaining Capital and Closure Costs

	Sustaining Capital (US$M)
G&A	0.9
Tailings	19.9
Mining	45.1
Total Sustaining Capital	65.9
	Closure Cost (US$M)
Closure Costs	65.3

21.2 Operating Cost Estimate

21.2.1 Operating Cost Summary

An average operating cost was estimated for the Arctic Project based on the proposed mining schedule. These costs included, mining, processing, G&A, surface services, and road toll costs. The average LOM operating cost for the Arctic Project is estimated to be $46.81/ t milled.

The processing plant throughput is designed to operate at approximately 10,000 tonnes/day, or 3,650,000 t/a. The proposed mining schedule ramps up in Year 1 and ramps down in Year 12 resulting in an LOM average of approximately 3,586,500 t processed per year. Total throughput was estimated to be 43,038,000 t over the 12-year life of mine. The breakdown of costs in Table 21-6 is estimated based on the average LOM mill feed rate.

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Table 21-6    Overall Operating Cost Estimate

Description	LOM Average Unit Operating Cost ($/ t milled)	Percentage of Total Annual Operating Costs
Mining*	20.47	44%
Processing	15.09	32%
G&A	5.60	12%
Surface Services	0.95	2%
Road Toll and Maintenance	4.70	10%
Total Operating Cost	46.81	100%

^*Excludes pre-production costs.

21.2.2 Mining Operating Cost Estimate

Operating costs average $3.06/primary tonne mined including stockpile rehandle. Excluding the preproduction period the average mining cost is $3.09/t. Total tonnage moved includes 0.63 Mt of stockpile rehandle and 339.5 Mt of ore and waste production. During PP -2, the mining costs are above average due to mining ramp-up and the lower than average tonnage mined. Starting in PP -1, mining costs are below average at $2.69/t mined due to the short hauls and high mining rates. Following PP -1, mining costs increase as the haul cycles increase with the deepening of the Phase I and Phase 2 laybacks. Mining costs then decrease in Year’s 6 and 7 with the completion of Phase 1 and 2 mining, and then increase with the deepening of the final Phase and a reduction in total tonnes mined. Table 21-7 shows the projected mining operating costs per period.

Table 21-7    Life of Mine Mining Cost

Period	Mining Cost	Primary Production		Total Mined1
	$US (000's)	(kt)	US$/t	(kt)	US$/t
PP -2	72,693	23,090	3.15	23,090	3.15
PP -1	85,949	32,002	2.69	32,002	2.69
Yr1	93,240	32,057	2.91	32,684	2.85
Yr2	93,859	31,976	2.94	31,976	2.94
Yr3	95,091	32,098	2.96	32,098	2.96
Yr4	100,698	32,483	3.10	32,483	3.10
Yr5	97,973	29,880	3.28	29,880	3.28
Yr6	86,140	28,297	3.04	28,297	3.04
Yr7	67,070	24,598	2.73	24,598	2.73
Yr8	61,808	20,578	3.00	20,578	3.00
Yr9	55,958	17,463	3.20	17,463	3.20
Yr10	54,618	15,877	3.44	15,877	3.44
Yr11	41,250	11,738	3.51	11,738	3.51
Yr12	33,327	7,344	4.54	7,344	4.54
Total	1,039,674	339,482	3.06	340,109	3.06

¹Total Material Mined includes low-grade stockpile rehandle.

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21.2.3 Processing Operating Cost Estimate

The LOM average process operating cost is $15.09/t milled. This estimated unit cost is based on the designed 10,000 tonnes per day throughput. Table 21-8 summarizes the processing operating cost estimates.

Table 21-8    Summary of Processing Operating Cost Estimates

Description	Annual Operating Costs ($M)	Annual Operating Costs ($/t milled)
Plant Operations Labour	8.49	2.33
Plant Maintenance Labour	13.45	3.69
Power Supply (Mill and Tailings)	19.10	5.23
Processing Consumables	12.01	3.29
Maintenance Supplies	1.80	0.49
Light Vehicles & Mobile Equipment	0.24	0.07
Total (Processing)	55.10	15.09

Plant operations labour costs is estimated to be $2.33/t milled and plant maintenance labour costs is estimated to be $3.69/t milled. The estimated labour force for plant operations and plant maintenance was estimated at 71 and 92 people respectively. Annual salaries and wages were supplied by Trilogy Metals. The estimate is based on providing a labour force to support continuous operations at 24 hours per day, 365 days per year.

The power supply cost of $5.23/t milled is based on an average use of 110,395 MWh per year and an electric energy price of $0.173/kWh for electric power generated by LNG on site.

Processing consumables costs include primary crushing liners and screens, grinding media, reagents, and other plant consumables. Consumable rates are estimated based on obtained quotes from suppliers. All costs include freight charges to site.

Annual maintenance supplies costs are estimated as a percentage of major capital equipment costs plus an allowance for freight charges.

21.2.4 General and Administrative and Surface Services Cost Estimates

The LOM average G&A costs are estimated to be $5.60/t milled. These estimated costs were developed with Trilogy Metals and Ausenco and include:

· Labour cost for the 38 administrative staff (12 hourly, and 26 salaried).

· Service cost for safety, training, medical and first aid expenditure, computer supplies and software, human resources services, and entertainment/membership.

· Asset operations costs including operating vehicles, and warehouse costs.

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· Contract services expenditures, including insurance, consulting, relocation expenses, recruitment, auditing and legal services.

· Camp costs and personnel transport.

· Other costs, including liaisons to local communities, sustainability costs, and an allowance for regional taxes and licenses.

Table 21-9 shows a summary of the G&A cost estimates.

Table 21-9    G&A Cost Estimates

Cost Description	Annual Costs ($)	Average Unit LOM Cost ($/t)
Labour Costs – Salaried Staff	3,545,220	0.97
Labour Costs – Hourly Staff	1,387,029	0.38
Airfare from Fairbanks	2,748,000	0.75
General Office Expense	250,000	0.07
Medical and First Aid	100,000	0.03
Environment	600,000	0.16
Travel	150,000	0.04
Training and Safety	200,000	0.05
Computer Supplies Including Software	100,000	0.03
Entertainment/Membership	50,000	0.01
Vehicles	186,494	0.05
Warehouse	200,000	0.05
Regional Taxes and Licenses Allowance	300,000	0.08
Communications	500,000	0.14
Insurance	1,500,000	0.41
Consulting/External Assays	200,000	0.05
Relocation Expense	50,000	0.01
Recruitment	100,000	0.03
Audit	100,000	0.03
Legal Services	100,000	0.03
Camp Cost Including Heating	7,078,000	1.94
Liaison Committee/Sustainability	500,000	0.14
Other	500,000	0.14
Total (G&A)	20,444,743	5.60

The surface service costs are estimated to be $0.95/t milled. These costs include:

· Labour costs for the 16-person surface services crew (one salaried and 15 hourly).

· Asset operations including heating, site services for general maintenance, general road maintenance, and ground transportation.

· Operation (labour, reagents, power) and maintenance costs for water treatment (annual maintenance costs were estimated as 5% of the capital cost).

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· Other expenses including airport operations and maintenance, road dust suppression.

Table 21-10 shows the surface services cost estimates.

Table 21-10    Surface Services Cost Estimates

Cost Description	Annual Costs ($)	Average Unit LOM Cost ($/t)
Surface Service Labour Costs	1,684,034	0.46
Asset Operations	358,102	0.10
Other	111,200	0.03
Water Treatment	1,273,600	0.36
Total (Surface Services)	3,426,936	0.95

21.2.5 Road Toll Cost Estimate

Ausenco has not independently estimated the road toll cost that Trilogy will pay to use the AMDIAP proposed to be built by AIDEA, a State-owned private corporation owned by the Government of Alaska.  AIDEA anticipates that there will be multiple users of the access road for multiple purposes with significantly different levels of use.  These variables make it difficult to accurately project, at this time, what the estimated road toll will be for the Arctic Project. Trilogy Metals provided Ausenco with the company’s view of the likely costs, and provided support for the assumptions using points raised during email discussions with AIDEA (Trilogy Metals, 2018). The email discussions support Trilogy’s interpretation as to the cost estimate for the purposes of this Report.

There is currently no developed surface access to the Arctic Project area and beyond. Access to the Arctic Project is proposed to be via AMDIAP, a road approximately 340 km (211 miles) long, extending west from the Dalton Highway where it would connect with the proposed Arctic Project area.  The final terminal for the road has not yet been determined. Although the capital costs of the road are not yet final, an estimate of approximately $300 million has been used in the PFS.  Although Trilogy has been in discussions with AIDEA who had been investigating alternatives to reduce the cost to construct the AMDIAP, the final cost of the road could be higher than $300 million.  The working assumption of the PFS 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 the Arctic Project) in order to pay back the costs of financing the AMDIAP.  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 and other users of the road which could also include commercial transportation of materials and consumer items that would use the AMDIAP to ship concentrates to the Port of Anchorage in Alaska and possibly provide goods and commercial materials to villages in the region.

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For the purposes of the PFS study, AIDEA and Trilogy Metals reviewed current bonding ability of AIDEA based on a $300 million 30-year bond rate of 6.00% compounded semi-annually and a $300 million 15-year bond rate of 5.50% compounded semi-annually.  Although the final toll payments will be negotiated with AIDEA and the Public-Private Partnership owners of the access road sometime in the future, it has been assumed that a toll would be paid based on the Arctic Project paying approximately $9.7 million each year for its 12-year mine life. The toll payments are assumed in the PFS to commence in Year 1 of production.  In addition, a road maintenance fee of $2/tonne milled processed has been assumed.

The toll payments equate to a LOM unit cost of $2.70/tonne milled, resulting in a total road toll and maintenance LOM unit cost of $4.70/tonne milled.

There is a risk to the capital and operating cost estimates if the toll road is not built in the time frame required for the Arctic Project, the design basis for the road cost estimate changes (for example from a single lane roadway as assumed in this Report to a dual lane), or if the annual toll charges that will be levied are significantly different from what was assumed.

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22.0 Economic Analysis

22.1 Introduction

An economic analysis was undertaken to determine the IRR, NPV and payback on initial investment of the Arctic Project. To complete this analysis, Ausenco developed a pre-tax cash flow model based on the production schedule of the Project.

The pre-tax financial results are:

· 38.0% IRR

· $1,935.2 million NPV at an 8% discount rate

· 1.9 year payback period, on the initial capital costs of $779.6 million.

Trilogy engaged the Canadian firm of Ernst & Young LLP (EY) in Vancouver to provide the basis for and prepare tax calculations for use in the post-tax economic evaluation of the Project with the inclusion of United States Federal Alaska State income taxes, and Alaska Mining License Tax (EY Information).

The post-tax financial results are:

· 33.4% IRR

· $1,412.7 million NPV at an 8% discount rate

· 2.0 year payback period, on the initial capital costs of $779.6 million

A sensitivity analysis was carried out to determine the sensitivity of the project’s NPV to key economic assumptions and factors.

22.2 Cautionary Statement and Forward Looking Information

The results of this economic analysis, represents forward looking information. The results depend on the inputs that are subject to a number of known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented in this section. Information that is forward looking includes mineral resource and mineral reserve estimates, commodity prices, the proposed mine production plan, construction schedule, projected mining and metallurgical recovery rates, proposed capital and operating cost estimates, closure cost estimates, toll road cost estimates, and assumptions on geotechnical, hydrogeological, environmental, permitting, and social information.

22.3 Inputs to the Cash Flow Model

The Project will consist of a three-year pre-production construction period, followed by 12 years of production. The NPV and IRR were calculated at the beginning of the construction period in Year -3. All currency is in US dollars (US$) unless otherwise stated.

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The cost and revenue estimates were assembled using real dollars, treating Year -3 as the base year. No escalation was applied to any of the estimates beyond this date.

Trilogy Metals provided Ausenco with the metal price projections for use in the Pre-Feasibility Study on which the Technical Report is based. Trilogy Metals established the pricing using a combination of two year trailing actual metal prices, and market research and bank analyst forward price projections, prepared in early 2018 by Jim Vice of StoneHouse Consulting Inc.

The long-term consensus metal price assumptions to be used in the Pre-Feasibility Study are:

· Copper: $3.00/lb.

· Lead: $1.00/lb.

· Zinc: $1.10/lb.

· Gold: $1,300/oz.

· Silver: $18.00/oz

These assumptions were supplied by Trilogy Metals, and reflect a consensus pricing based on current metal prices and forecasts in the public domain.

The life of mine material tonnages and payable metal production used in the cash flow model can be found in Table 22-1.

The estimated life of mine average annual payable metal produced is approximately 159 million pounds of copper, 199 million pounds of zinc, 33 million pounds of lead, 30,600 ounces of gold, and 3.3 million ounces of silver.

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Table 22-1    Mine and Payable Metal Production for the Arctic Mine

Description	Units	Value
Total Tonnes Mined	ktonnes	340,109
Mill Feed	ktonnes	43,038
Concentrate
Cu Concentrate	ktonnes	2,961
Zn Concentrate	ktonnes	2,163
Pb Concentrate	ktonnes	354
Payable Metal
Payable Cu	('000'lb)	1,908,493
Payable Zn	('000'lb)	2,399,128
Payable Pb	('000'lb)	405,727
Payable Au	('000'oz)	368
Payable Ag	('000'oz)	40,238

22.4 Basis of Pre-Tax Financial Evaluation

The pre-tax financial model incorporated the production schedule and smelter term assumptions to produce annual recovered payable metal, or gross revenue, in each concentrate stream by year. Off-site costs, including the applicable refining and treatment costs, penalties, concentrate transportation charges, and marketing and representation fees, and royalties were then deducted from gross revenue to determine the NSR. Further details of the smelter terms used to calculate the recovered metal value and off-site operating costs can be found in Section 19.0. Royalties are discussed in Section 4.0.

The operating cash flow was then produced by deducting annual mining, processing, G&A, surface services, and road toll charges from the NSR.

Initial and sustaining capital was deducted from the operating cash flow in the years they occur, to determine the net cash flow before taxes.

Initial capital cost includes all estimated expenditures in the construction period, from Year -3 to Year -1 inclusive. First production occurs at the beginning of Year 1. Sustaining capital expenditure includes all capital expenditures purchased after first production, including mine closure and rehabilitation.

Ausenco has relied on Trilogy Metals as to how the royalties applicable to the Arctic Project are applied in the economic analysis and the related text below.

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Under the NANA Agreement, NANA has the right, following a construction decision, to elect to purchase a 16% to 25% direct interest in the Arctic Project or, alternatively, to receive a 15% Net Proceeds Royalty (refer to Section 4.0).  This financial analysis was carried out on a 100% ownership basis and does not include any future potential impact on Trilogy Metals’ Project interest if NANA elects to purchase an interest of between 16% and 25% in the Arctic Project under the NANA Agreement or, alternatively, the impact on Trilogy Metals’ Project interest if the 15% net proceeds royalty becomes applicable.  The financial analysis does include the 1.0% NSR to be granted to NANA under the NANA Agreement in exchange for a surface use agreement.

The financial analysis was carried out on a 100% ownership basis and does not include any future impact on Trilogy Metals’ Project interest should South32 elects to form the joint venture under the South 32 Option Agreement (refer to Section 4.0).

22.5 Pre-Tax Financial Results

A summary of the pre-tax financial results is provided in Table 22-2.

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Table 22-2    Summary of Pre-Tax Financial Results

Description	Unit	LOM Value
Recovered Metal Value
Copper	US$ million	5,725.5
Lead	US$ million	405.7
Zinc	US$ million	2,639.0
Gold	US$ million	477.8
Silver	US$ million	724.3
Total Recovered Metal Value	US$ million	9,972.3
Off-Site Operating Costs
Royalties, Refining and Treatment Charges, Penalties, Insurance, Marketing and Representation & Concentrate Transportation	US$ million	2,526.8
On-Site Operating Costs
Mining	US$/t milled	20.47
Processing	US$/t milled	15.09
G&A	US$/t milled	5.60
Surface Service	US$/t milled	0.95
Road Toll	US$/t milled	4.70
Total Operating Cost	US$/t milled	46.81
Total Operating Cost	US$ million	2,014.7
Capital Expenditure
Initial Capital	US$ million	779.6
Sustaining Capital	US$ million	65.9
Mine Closure & Reclamation	US$ million	65.3
Total Capital Expenditure	US$ million	910.8
Financial Summary
Pre-tax Undiscounted Cash Flow	US$ million	4,520.1
NPV at 8%	US$ million	1,935.2
Cash Costs, Net of By-product Credits	($US/lb Cu payable)	0.15
Capital Costs (per lb of Payable Cu)	($US/lb Cu payable)	0.48
IRR	%	38.0
Payback Period	years	1.9

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22.6 Post-Tax Financial Analysis

The following tax regimes were incorporated in the post-tax analysis as provided by EY as part of the EY Information: US Federal Income Tax, Alaska State Income 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, including the US Federal enactment of the Tax Cuts & Jobs Act (TCJA) on December 22, 2017.

22.6.1 US Federal Tax

For tax years ending in 2017 and prior, US Federal income taxes are determined annually as the higher of “regular” corporate income tax and Alternative Minimum Tax (AMT). Under the TCJA, the corporate AMT tax regime has been repealed for all tax years beginning on or after January 01, 2018.

For tax years beginning on or after January 01, 2018, the US Federal income tax corporate rate is 21% of taxable income, as opposed to a 35% rate which was applicable to prior tax years. Taxable income is calculated as revenues less allowable costs. In addition to other allowable costs, Alaska State Income Tax, AMLT, tax depreciation and the greater of the cost depletion or percentage depletion can be deducted. Cost depletion is the ratable recovery of cost basis as units are produced and sold. IRC §613(a) governs percentage depletion and provides that the deduction for depletion shall be a statutorily prescribed percentage of the taxpayer’s gross income from the mineral property during the taxable year. Such allowance shall not exceed 50% of the taxpayer’s taxable income from the property that is mining related. Relevant statutorily prescribed percentages are 15% for gold, silver and copper, and 22% for lead and zinc. Losses incurred during tax years prior to January 01, 2018 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 to the extent certain requirements are satisfied. As a result of the TCJA, losses incurred for tax years beginning on or after January 01, 2018 are not eligible to be carried back but may be carried forward indefinitely. However, losses generated under the TCJA are only eligible to offset 80% of taxable income in future years.

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.

22.6.2 Alaska State Tax

AST is determined on a basis similar to US federal tax. AST is calculated using a graduated rate table times taxable income with 9.4% being the highest applicable rate, where taxable income is calculated on the same basis as US federal tax (except that State tax is not deductible). The Alaskan AMT statutes are tied to the federal AMT statutes; therefore, the repeal of federal corporate AMT has effectively repealed Alaskan State AMT for tax years beginning on or after January 01, 2018.

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22.6.3 Alaska Mining License Tax

AMLT is an income-based tax imposed on the mining income calculated for 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 evaluated at the project level, it was assumed that the initial EIC balance is zero even though the Trilogy Metals 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.

22.6.4 Post-Tax Financial Results

At the base case metal prices used for the Report, the total estimated taxes payable on the Arctic Project profits are $1,162.2 million over the 12-year mine life.

The post-tax financial results are summarized in Table 22-3.

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Table 22-3    Summary of Post-Tax Financial Results

Description	Unit	LOM Value
Recovered Metal Value
Copper	US$ million	5,725.5
Lead	US$ million	405.7
Zinc	US$ million	2,639.0
Gold	US$ million	477.8
Silver	US$ million	724.3
Total Recovered Metal Value	US$ million	9,972.3
Off-Site Operating Costs
Royalties, Refining and Treatment Charges, Penalties, Insurance, Marketing and Representation, and Concentrate Transportation	US$ million	2,526.8
On-Site Operating Costs
Mining	US$/t milled	20.47
Processing	US$/t milled	15.09
G&A	US$/t milled	5.60
Surface Service	US$/t milled	0.95
Road Toll	US$/t milled	4.70
Total Operating Cost	US$/t milled	46.81
Total Operating Cost	US$ million	2,014.7
Capital Expenditure
Initial Capital	US$ million	779.6
Sustaining Capital	US$ million	65.9
Mine Closure & Reclamation	US$ million	65.3
Total Capital Expenditure	US$ million	910.8
Financial Summary
Pre-tax Undiscounted Cash Flow	US$ million	4,520.1
Income Tax	US$ million	1,162.2
Undiscounted Post-Tax Cash Flow	US$ million	3,357.8
Post-tax NPV at 8%	US$ million	1412.7
IRR	%	33.4
Payback Period	years	2.0

22.7 Cash Flow

The annual production schedule and estimated cash flow forecast for the Arctic Project can be found in Table 22-4.

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Table 22-4    Pre and Post-Tax Arctic Project Production and Cash Flow Forecast

Arctic Project	Units	LOM	Year -3	Year -2	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Year 10	Year 11	Year 12
Mine Production
Total Tonnes Mined	ktonnes	340,109	-	23,090	32,002	32,684	31,976	32,098	32,483	29,880	28,297	24,598	20,578	17,463	15,877	11,738	7,344
Mill Feed	ktonnes	43,038	-	-	-	3,096	3,649	3,651	3,650	3,650	3,651	3,651	3,650	3,651	3,650	3,650	3,438
Cu	%	2.32	-	-	-	2.61	2.30	2.18	2.19	2.05	2.21	2.35	2.32	2.66	2.45	2.34	2.17
Zn	%	3.24	-	-	-	3.19	3.32	3.47	3.24	2.57	2.65	2.81	3.58	3.96	3.29	3.49	3.37
Pb	%	0.57	-	-	-	0.53	0.62	0.63	0.62	0.40	0.44	0.47	0.60	0.65	0.54	0.58	0.70
Au	g/t	0.49	-	-	-	0.43	0.53	0.49	0.50	0.31	0.37	0.48	0.52	0.57	0.56	0.53	0.55
Ag	g/t	35.98	-	-	-	37.31	36.98	33.60	32.30	25.38	29.07	34.24	38.95	45.28	39.62	38.68	40.88
Concentrate
Cu Concentrate	ktonnes	2,961	-	-	-	239.7	249.3	236.8	237.0	222.3	239.2	254.9	251.8	288.7	265.7	253.7	221.6
Cu Recovery	%	90.0	-	-	-	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0
Cu Concentrate Grade	%	30.3	-	-	-	30.3	30.3	30.3	30.3	30.3	30.3	30.3	30.3	30.3	30.3	30.3	30.3
Cu Payable	%	96.5	-	-	-	96.5	96.5	96.5	96.5	96.5	96.5	96.5	96.5	96.5	96.5	96.5	96.5
Ag Concentrate Grade	g/t	169	-	-	-	169.0	169.0	169.0	169.0	169.0	169.0	169.0	169.0	169.0	169.0	169.0	169.0
Ag Payable	%	90.0	-	-	-	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0	90.0
Zn Concentrate	ktonnes	2,163	-	-	-	152.9	187.5	196.3	183.5	145.4	149.7	158.6	202.1	224.1	185.8	197.4	179.3
Zn Recovery	%	91.7	-	-	-	91.7	91.7	91.7	91.7	91.7	91.7	91.7	91.7	91.7	91.7	91.7	91.7
Zn Concentrate Grade	%	59.2	-	-	-	59.2	59.2	59.2	59.2	59.2	59.2	59.2	59.2	59.2	59.2	59.2	59.2
Zn Payable in Concentrate	%	85.0	-	-	-	85.0	85.0	85.0	85.0	85.0	85.0	85.0	85.0	85.0	85.0	85.0	85.0
Pb Concentrate	ktonnes	354	-	-	-	24.1	33.0	33.3	32.8	21.4	23.3	24.8	31.9	34.7	28.7	30.8	35.2
Pb Recovery	%	80.0	-	-	-	80.0	80.0	80.0	80.0	80.0	80.0	80.0	80.0	80.0	80.0	80.0	80.0
Pb Concentrate Grade	%	55.0	-	-	-	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0
Pb Payable	%	55.0	-	-	-	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0	55.0
Ag Concentrate Grade	g/t	2,383	-	-	-	2,383	2,383	2,383	2,383	2,383	2,383	2,383	2,383	2,383	2,383	2,383	2,383
Ag Payable	%	95.0	-	-	-	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00
Au Concentrate Grade	g/t	34	-	-	-	34.0	34.0	34.0	34.0	34.0	34.0	34.0	34.0	34.0	34.0	34.0	34.0
Au Payable	%	95.0	-	-	-	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00	95.00
Payable Metal
Payable Cu	('000'lb)	1,908,493	-	-	-	154,546	160,727	152,632	152,777	143,278	154,194	164,282	162,307	186,108	171,268	163,547	142,827
Payable Zn	('000'lb)	2,399,128	-	-	-	169,614	208,053	217,759	203,512	161,269	166,085	175,976	224,246	248,643	206,072	218,949	198,950

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Arctic Project	Units	LOM	Year -3	Year -2	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Year 10	Year 11	Year 12
Payable Pb	('000'lb)	405,727	-	-	-	27,618	37,819	38,171	37,568	24,518	26,756	28,453	36,565	39,770	32,875	35,277	40,336
Payable Au	('000'oz)	368	-	-	-	25	34	35	34	22	24	26	33	36	30	32	37
Payable Ag	('000'oz)	40,238	-	-	-	2,926	3,620	3,581	3,544	2,644	2,868	3,053	3,553	3,937	3,386	3,480	3,644
Contained Metal Value
Contained Metal Value	US$ M	9,972.3	-	-	-	763.0	858.6	845.0	827.8	708.2	755.2	803.3	877.2	989.3	873.0	871.0	800.8
Off-Site Charges
Off-Site Charges*	US$ M	2,526.8	-	-	-	191.3	217.2	216.2	209.8	178.5	189.1	201.0	224.7	252.9	221.0	222.6	202.5
Capital Costs
Mining	US$ M	326.2	51.7	142.1	87.3	10.3	0.0	2.2	3.6	0.1	2.8	8.6	13.5	4.0	-	-	-
Crushing	US$ M	18.3	-	6.1	12.2	-	-	-	-	-	-	-	-	-	-	-	-
Processing	US$ M	114.7	-	56.7	57.1	-	-	0.3	-	-	0.3	-	-	0.3	-	-	-
Tailings	US$ M	50.2	-	14.9	15.4	9.0	-	-	-	7.9	-	-	-	-	3.0	-	-
On-Site Infrastructure	US$ M	84.5	13.2	41.6	29.8	-	-	-	-	-	-	-	-	-	-	-	-
Off-Site Infrastructure	US$ M	15.6	10.2	4.7	0.7	-	-	-	-	-	-	-	-	-	-	-	-
Indirect Costs	US$ M	121.9	40.3	53.4	28.2	-	-	-	-	-	-	-	-	-	-	-	-
Owner's Cost	US$ M	21.9	6.6	6.6	8.7	-	-	-	-	-	-	-	-	-	-	-	-
Mine Closure	US$ M	65.3	-	-	-	-	-	-	-	-	-	-	-	-	-	-	65.3
Contingency	US$ M	92.0	12.7	45.0	34.3	-	-	-	-	-	-	-	-	-	-	-	-
Total Capital Costs	US$ M	910.8	134.7	371.2	273.6	19.3	0.0	2.5	3.6	8.0	3.1	8.6	13.5	4.3	3.0	-	65.3
Operating Costs
Mining	US$ M	881.0	-	-	-	93.2	93.9	95.1	100.7	98.0	86.1	67.1	61.8	56.0	54.6	41.3	33.3
Processing	US$ M	649.7	-	-	-	46.7	55.1	55.1	55.1	55.1	55.1	55.1	55.1	55.1	55.1	55.1	51.9
Surface Operations	US$ M	40.7	-	-	-	3.1	3.4	3.4	3.4	3.4	3.4	3.4	3.4	3.4	3.4	3.4	3.3
G&A	US$ M	240.8	-	-	-	17.3	20.4	20.4	20.4	20.4	20.4	20.4	20.4	20.4	20.4	20.4	19.2
Road Toll & Maintenance	US$ M	202.5	-	-	-	15.9	17.0	17.0	17.0	17.0	17.0	17.0	17.0	17.0	17.0	17.0	16.6
Total Operating Costs	US$ M	2,014.7	-	-	-	176.3	189.8	191.1	196.7	193.9	182.1	163.0	157.8	151.9	150.6	137.2	124.3
Undiscounted Pre-Tax Cash Flow	US$ M	4,520.1	(134.7)	(371.2)	(273.6)	376.1	451.6	435.2	417.8	327.7	380.9	430.7	481.2	580.2	498.5	511.1	408.6
Income Tax	US$ M	1,162.2	-	-	-	12.6	39.9	79.2	85.3	71.3	91.6	114.1	131.7	159.3	135.1	139.2	103.0
Undiscounted Post-Tax Cash Flow	US$ M	3,357.8	(134.7)	(371.2)	(273.6)	363.6	411.6	356.0	332.5	256.4	289.3	316.7	349.5	420.9	363.4	371.9	305.6

*Costs include Royalties, Insurance, Marketing and Representation Fees, Refining, Treatment, Penalties, and Concentrate Transport

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22.8 Sensitivity Analysis

Ausenco investigated the sensitivity of the Project’s pre-tax NPV, and IRR to several project variables. The following variables were elected for this analysis:

· copper price

· zinc price

· lead price

· gold price

· silver price

· capital costs

· on-site operating costs

· off-site operating costs (royalties, refining and treatment charges, penalties, insurance, marketing and representation fees, and concentrate transportation)

Each variable was changed in increments of 10% between -30% to +30% while holding all other variables constant. Figure 22-1 and Figure 22-2 show the results of the pre-tax sensitivity analysis.

The metal grade is not presented in theses sensitivity graphs because the impacts of changes in the metal grade mirror the impact of changes in metal price.

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Figure 22-1    Pre-tax NPV Sensitivity Analysis (Ausenco, 2018)

As shown in Figure 22-1, the Project’s NPV at an 8% discount rate is most sensitive to changes in copper price, followed by zinc price, off-site operating costs, on-site operating costs, capital costs, silver price, gold price, and lead price.

Figure 22-2    Pre-tax IRR Sensitivity Analysis (Ausenco, 2018)

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As shown in Figure 22-2, the Project’s IRR is most sensitive to changes in copper price and capital cost, followed by zinc price and off site operating costs, and in then decreasing order, on-site operating costs, silver price, gold price, and lead price.

22.9 Copper and Zinc Metal Price Scenarios

Metal price scenarios were completed to determine the effects of copper and zinc price on the Project’s IRR, payback period and NPV at an 8% discount rate. In the first scenario the copper price was varied from $2.00/lb to $4.00/lb, while holding all other variables constant, the results of this scenario can be found in Table 22-5. The base case scenario is bolded in the table.

Table 22-5    Pre-tax Copper Price Scenarios

Variable	Unit	Copper Price ($/lb)
		2.00	2.50	3.00	3.50	4.00
NPV at 8%	US$ million	993.4	1,464.3	1,935.2	2,406.1	2,877
IRR	%	25.3	31.9	38.0	43.6	48.9
Payback Period	Years	2.9	2.3	1.9	1.6	1.4

In the second scenario the zinc price was varied from $0.90/lb to $1.30/lb, while holding all other variables constant, the results of this scenario can be found in Table 22-6. The base case scenario is bolded in the table.

Table 22-6    Pre-tax Zinc Price Scenarios

Variable	Unit	Zinc Price ($/lb)
		0.90	1.00	1.10	1.20	1.30
NPV at 8%	US$ million	1699.6	1817.4	1935.2	2053.0	2170.8
IRR	%	35.1	36.5	38	39.4	40.8
Payback Period	years	2.1	2	1.9	1.8	1.8

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23.0 Adjacent Properties

This section is not relevant to this Report.

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24.0 Other Relevant Data and Information

This section is not relevant to this Report.

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25.0 Interpretation and Conclusions

25.1 Introduction

The QPs note the following interpretations and conclusions in their respective areas of expertise, based on the review of data available for this Report.

25.2 Mineral Tenure, Surface Rights, Royalties and Agreements

Information from legal experts supports that the mining tenure held is valid and is sufficient to support declaration of Mineral Resources and Mineral Reserves

Kennecott holds a 1% NSR royalty that is purchasable at any time for a one-time payment of $10 million. This royalty covers the patented federal mining claims and many, but not all, of the State mining claims.

The NANA Agreement granted to NovaCopper US certain rights, for consideration, which include an exclusive right to explore for minerals and a nonexclusive right to enter upon and use certain NANA lands for various purposes including access to NovaCopper’s Ambler mining properties. The NANA Agreement further provides that if NovaCopper completes a Feasibility Study and a Draft Environmental Impact Statement for a project to develop the Ambler properties, NANA will become entitled elect to purchase an interest in the project by exercising a back-in right to acquire an undivided interest of between 16% and 25% (at NANA’s option) of the mining project. If this back-in right were to be exercised by NANA the two parties would form a Joint Venture to proceed with the project. The Joint Venture would lease the mining properties from NovaCopper US Inc. in exchange for a 1% NSR royalty and NANA would enter into a surface use agreement with the Joint Venture in exchange for a 1% NSR royalty. In the alternative, should NANA elect not to exercise this back-in right, the Joint Venture would not be formed and NANA would instead become entitled to a 15% net proceeds royalty. .

South32 may become entitled to exercise an option to require that Trilogy Metals Inc. form a new limited liability company, 50% of the Membership Interest of which would be held by South32, and Trilogy Metals would be obligated to transfer all of its assets to the new entity. The South32 Agreement addresses the coordination of the rights and obligations it creates with those arising under the NANA Agreement.

The owner of a State mining claim or lease will be obligated to pay a production royalty to the State of Alaska in the amount of 3% of net income received from minerals produced from the State mining claims

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25.3 Geology and Mineralization

The Arctic Deposit is considered to be an example of a VMS system.

Knowledge of the deposit settings, lithologies, mineralization style and setting, and structural and alteration controls on mineralization is sufficient to support Mineral Resource and Mineral Reserve estimation

25.4 Exploration, Drilling and Analytical Data Collection in Support of Mineral Resource Estimation

The quantity and quality of the lithological, geotechnical, collar and downhole survey data collected in the exploration and infill drill programs conducted is sufficient to support Mineral Resource and Mineral Reserve estimation.

Analytical and density data are suitable to support Mineral Resource and Mineral Reserve estimation.

NovaGold, NovaCopper and Trilogy Metals sample security procedures met industry standards at the time the samples were collected. Current sample storage procedures and storage areas are consistent with industry standards.

Data collected have been sufficiently verified that they can support Mineral Resource and Mineral Reserve estimation and be used for mine planning purposes.

25.5 Metallurgical Testwork

Metallurgical studies have spanned over 30 years.

Test work conducted prior to 2012 is considered relevant to the project, but predictive metallurgical results are considered to be best estimated from test work conducted on sample materials obtained from exploration work under the direction of Trilogy Metals conducted in 2012 and 2017.

To the extent known, the metallurgical samples are representative of the styles and types of mineralization and the mineral deposit as a whole.

The LOM average mill feed is expected to contain 2.32% Cu, 3.24% Zn, 0.57% Pb, 0.49 g/t Au, and 35.98 g/t Ag.

Concentrate quality test results indicated that key penalty elements, as well as precious metals are typically concentrated into a lead concentrate, leaving the copper concentrate of higher than expected quality given the levels of impurities seen in the test samples. The lead concentrate may have penalties for the high arsenic and antimony concentrations seen in the results of this test work. Precious metal deportment into a lead concentrate is very high and should benefit the payable levels of precious metals at a smelter. Silicon dioxide and fluoride assays should be conducted on the concentrates to determine whether or not they are higher than the penalty thresholds.

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Talc will be managed through a pre-float step.

In general, the flowsheet developed in the 2012 test program and further tested in the 2017 test work program at ALS Metallurgy is feasible for the Arctic Deposit mineralization. Further metallurgical test work is recommended on representative samples to confirm and optimize the flowsheet and better understand the impact of talc levels in the process feed samples.  Lead concentrate quality can be impacted by the level of talc in the process feed and a better understanding of the level of talc in an expected process feed is critical in maximizing the value of a lead concentrate.  There are no outstanding metallurgical issues related to the production of a copper or zinc concentrate from all of the materials tested. 

On-going grinding test work is recommended prior to the commencement of a feasibility study, including detailed SAG mill characterization test work. 

25.6 Mineral Resource Estimates

Mineral Resources have been prepared using industry-standard methods and software.

Mineral Resources have had reasonable prospects of eventual economic extraction considerations applied.

Mineral Resources conform to the 2014 CIM Definition Standards.

25.7 Mineral Reserves and Mine Planning

Amec Foster Wheeler selected conventional open-pit mining because of the deposit’s geometry and proximity to surface. An owner-operated and maintained fleet has been specified, with outside providers supporting mine operations.

The Arctic Project PFS mine plan is based on Probable Mineral Reserves resulting from modifying factors being applied to a subset of the Indicated Mineral Resource estimates. The pit will operate for 14 years, including two years of pre-production.

25.8 Recovery Plan

The recovery plan is conventional.

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.

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25.9 Project Infrastructure

The planned mine will be a greenfields site and require construction of mine and process-related infrastructure. Access roads in and around the Project site will be required. Site access assumes that the AMDIAP road will be available.

Power generation will be provided by six LNG generators.

25.10 Environmental, Permitting and Social

To date, a moderate amount of baseline environmental data collection has occurred in the area including surface water quality sampling, surface hydrology monitoring, wetlands mapping, stream flow monitoring, aquatic life surveys, avian and mammal habitat surveys, cultural resource surveys, hydrogeology studies, meteorological monitoring, and acid base accounting studies.

The Arctic Project will be subject to a mine permitting process typical for a mine of its size in Alaska. In order to support this process, Trilogy Metals will have to broaden their existing baseline environmental program and complete a number of studies that will support the permit applications.

Trilogy Metals has formally started engaging the Arctic Project stakeholders and recognizes the need to earn their trust and support by making the Arctic Project directly beneficial to them throughout the life of the Arctic Project.

Trilogy Metals will be required to develop a mine plan that is protective of the environment during mining operations as well as reclamation and closure plan that ensures the environment is protected after mine closure.

Closure activities will be undertaken at the end of mine life to bring the mine facilities in a state consistent with the stated closure objectives and compliant with the regulations for closure and abandonment.

Closure costs at this stage of Project evaluation are estimated at $65.3 million.

25.11 Markets and Contracts

No contracts are currently in place for any production from the Project.

25.12 Capital Costs

Overall capital costs are estimated at 779.6 million. The estimate accuracy is -20% to +30%.

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25.13 Operating Costs

The average LOM operating cost for the Arctic Project is estimated to be $46.81/ t milled.

There is a risk to the capital and operating cost estimates if the toll road is not built in the time frame required for the Arctic Project, the design basis for the road cost estimate changes (for example from a single lane roadway as assumed in this Report to a dual lane), or if the annual toll charges that will be levied are significantly different from what was assumed.

25.14 Economic Analysis

The financial analysis was conducted on a 100% ownership basis.

The base case pre-tax NPV was $1,935.2 million, calculated at the beginning of the construction period in Year -3, using an 8% discount rate. The base case pre-tax IRR, and payback period on initial capital were 38.0%, and 1.9 years respectively.

The post-tax NPV was $1,412.7 million, calculated at the beginning of the construction period in Year -3 using an 8% discount rate. The post-tax IRR and payback period on initial capital were 33.4%, and 2.0 years respectively.

The Project’s pre-tax NPV at an 8% discount rate is most sensitive to changes in copper price, followed by zinc price, off-site operating costs, on-site operating costs, capital costs, silver price, gold price, and lead price.

The Project’s pre-tax IRR is most sensitive to changes in copper price and capital cost, followed by zinc price and off site operating costs, and in then decreasing order, on-site operating costs, silver price, gold price, and lead price.

The economic analysis does not include an allocation for the 1% NSR payable to Kennecott or the alternative option to purchase payment of $10 million. It is likely that Trilogy Metals would avail itself of the option to purchase the NSR; the impact of such a purchase on the NPV was included within the ranges shown for the capital cost sensitivity in Figure 22-1. It was assumed that no payment of any 3% NSR to the State of Alaska on State claims was required; the impact of such this royalty payment on the NPV was included within the ranges shown for the off-site operating costs sensitivity in Figure 22-1.

25.15 Conclusions

Under the assumptions presented in this Report, the Project demonstrates positive economics.

The financial analysis does not include any future potential impact on Trilogy Metals’ Project interest if NANA elects to purchase an interest of between 16% and 25% in the Arctic Project under the NANA Agreement or, alternatively, the impact on Trilogy Metals’ Project interest if the 15% net proceeds royalty becomes applicable should NANA elect not to exercise this back-in right. The financial analysis also does not include any future impact on the Trilogy Metals’ Project interest if South32 elects to form a joint venture under the South 32 Option Agreement.

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The economic analysis does not include an allocation for the 1% NSR payable to Kennecott or the alternative option to purchase payment of $10 million. It is likely that Trilogy Metals would avail itself of the option to purchase the NSR; the impact of such a purchase on the NPV was included within the ranges shown for the capital cost sensitivity in Figure 22-1.

The Mineral Reserve pit extends onto a portion of the State claims. A 3% net income received from minerals produced from the State mining claims is payable to the State. The economic analysis does not include an allocation for such a royalty. The potential impact of this royalty on the economic analysis was included within the ranges shown for the onsite operating cost sensitivity in Figure 22-1.

The financial analysis excludes consideration of the NANA Agreement, whereby NANA has the right, following a construction decision, to elect to purchase a 16% to 25% direct interest in the Arctic Project or, alternatively, to receive a 15% Net Proceeds Royalty.

The financial analysis excludes consideration of the South32 Option Agreement, whereby South32 has the right to form a 50/50 Joint Venture with Trilogy Metals over Trilogy Metal’s Alaskan interests, including the Arctic Project.

The cost assumptions for the AMDIAP road are estimates provided by Trilogy Metals. There is a risk to the capital and operating cost estimates, the financial analysis, and the Mineral Reserves if the toll road is not built in the time frame required for the Arctic Project, or if the toll charges are significantly different from what was assumed.

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26.0 Recommendations

26.1 Introduction

A single-phase work program is proposed by Ausenco for the Project at a total cost of about $3.3 million.

26.2 Work Program

The following work is recommended:

· Geotechnical studies, including geotechnical investigations of the pit area, plant site, WRD & TMF site, WRCP site, airstrip and other project related locations (estimated cost of $1,000,000), including

o conduct further geotechnical engineering studies for the foundations of the TMF, WRD, and WRCP including borehole drilling and test pit excavations to test all assumptions made in this report and determine the foundation, borrow, and fill placement conditions for design

o complete detailed geotechnical characterization of the materials that will be used to construct the TMF & WRCP

· Structural and Hydrogeological studies required for a feasibility study slope design (estimated cost of $600,000), including:

o The next phase requires a combined geotechnical, hydrogeological and structural field program of around 3000m. The program should include a pump test, airlift testing and some water quality work. Additional outcrop mapping will be required that includes rock sampling for additional testing on the talc unit.

o The structural and lithological model should be updated based on the additional drilling and mapping, and further work should be undertaken to constrain timing relationships between the brittle structures.

o Ongoing hydrogeological evaluations should assess compartmentalization and the potential for confined pressure below the talc beds. Transient pore pressure models should be developed and their impact on interim slopes assessed. Plans need to be developed to monitor pore pressures and contingency plans developed for pore pressure reduction. Additional work is required to predict pit water quality; inflow and water quality results should be integrated into site water management.

o The rock mass model needs to be fully integrated with the updated geological and structural models. Additional geotechnical assessments need to consider the impact of the near surface weathering, potential seismic hazards and the potential for buckling on the footwall. A full evaluation of the interim slopes should be undertaken that considers the influence of the talc and the potential water pressures.

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· Engineering studies, including power supply and optimization of the layout of the process and service related facilities (estimated cost of $500,000).

· Engineering studies, including water management and treatment, WRD and TMF design (estimated cost of $600,000), including

o more detailed analysis of the water and load balance to predict the accumulation of mill reagents and their degradation products in the process water circuit

o evaluating the size of the mixing zone need in the Shungnak River to meet in stream selenium water quality limits

o updating pit groundwater numerical model and inflow estimates, and incorporating seasonal effects to assess high and low water conditions

o updating WRD and TMF design based on additional field investigation results

o developing tailings deposition plan and waste rock placement sequence to match pit development and mill output.

· Additional baseline studies and environmental permitting activities ($100,000), including

o precipitation monitoring, including snowfall, in Sub-Arctic Valley to confirm the revised estimates of mean annual precipitation.

· Metallurgical studies focused on grinding test work and additional floatation test work primarily to provide better information on the internal water balance of the plant and characterization of the process water chemistry owing to the use of cyanide in the lead separation circuit (estimated cost of $500,000), including

o conducting detailed test work to demonstrate the operational requirements of the proposed cyanide destruction plant.

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27.0 References

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ADF&G, 2015, Caribou Management Report - Species Management Report, Alaska Department of Fish and Game, Division of Wildlife Conservation, ADF&G Report 2015-4.

ADF&G, 2011, Caribou Management Report - Species Management Report, Alaska Department of Fish and Game, Division of Wildlife Conservation, Patricia Harper Editor, 2011.

Alaska Industrial Development and Export Authority, AIDEA, and Alaska Energy Authority, 2013, Interior Energy Project Feasibility Report. Proposed Project – North Slope LNG Plant, July.

Albers, D.A., 2004, Arctic Project Status Report: NovaGold Internal report.

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.

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Broman, B.N., 2014, Metamorphism and Element Redistribution: Investigations of Ag-bearing and associated minerals in the Arctic Volcanogenic Massive Sulfide Deposit, SW Brooks Range, NW Alaska.

Brown, W.J., 1985, Pre-AFD Report

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CIM, May 2014, CIM Definition Standards - For Mineral Resources and Mineral Reserves.http://web.cim.org/UserFiles/File/CIM_DEFINITION_STANDARDS_MayNov_20140.pdf.

Clark, L.A. and Sweeney, M.J., 1976, Volcanic Stratigraphy, Petrology, and Trace Metal Geochemistry, Arctic Deposit: Kennecott Exploration Internal report.

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Craig, C., 2017, 2017 UKMP Water Quality Report, May 17th, Internal Report by Trilogy Metals Inc., June.

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Craig, C., 2015, 2015 UKMP Water Quality Report, Third Quarter, Internal Report by NovaCopper US Inc., October.

Craig, C., 2014, 2014 UKMP Water Quality Report, Third Quarter, Internal Report by NovaCopper US Inc., October.

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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.

Dimock, R.R., 1984, Arctic Evaluation Update: Kennecott Exploration Internal report

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DOWL, 2016, Large Mammal Survey, Consultant Report for Trilogy Metals, Inc. December.

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Exploration Agreement and Option to Lease between NovaCopper US Inc. and NANA Regional Corporation, Inc. dated October 19, 2011, as amended.

Franklin, J.M., Gibson, H.L., Jonasson, I.R., and Galley, A.G., 2005, Volcanogenic Massive Sulfide Deposits, in Hedenquist, J.W., Thompson, J.F.H., Goldfarb, R.J., and Richards, J.P., eds., Economic Geology 100th Anniversary Volume: The Economic Geology Publishing Company, p. 523-560.

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Hale, C., 1996, 1995 Annual Ambler District Report: Kennecott Exploration Internal report.

Halls, J.L. 1974, Ambler District Evaluation: Kennecott Exploration Internal report

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Halls, J.L. 1978, Arctic Deposit: Kennecott Exploration Internal report

Hammitt, J.W., 1985, 1985 Annual Progress Report – Ambler District: Kennecott Corporation Internal Report.

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.

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.

Hunt, G.A., 1999, PIMA alteration mapping and structural interpretation of drill core from the Arctic Deposit, Alaska, a report prepared for Rio Tinto Mining and Exploration Ltd, 10 p.

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Kennecott, 1977, Annual Report Arctic Deposit: Unpublished in-house report.

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Kennecott Research Center, August 1972, Amenability Testing of Samples from Bear Creek Mining Company’s Arctic Deposit, TR 72-12.

Kennecott Research Center, January 1997, Process Selection for Arctic Deposit, Technical Report RTR 77-4.

Kennecott Research Center, September 1968, Amenability Testing of Diamond Drill Core Samples from Arctic, Alaska Project, TR 68-20.

Kennecott Research Center, September 1976, Recovery of Mineral Values Arctic Prospect, RTR 76¬22.

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.

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Lindberg, P. A., 2005, A Preliminary Attempt to Unfold the Arctic Volcanogenic Ore Deposit and Determine it’s Original Metal Zonation: Alaska Gold Company Unpublished Report.

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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, Geologic Society of America, The Geology of North America, v. G-1.

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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.

Newberry, R.J., Crafford, T.C., Newkirk, S.R., Young, L.E., Nelson, S.W., and Duke, N.A., 1997, Volcanogenic massive sulfide deposits of Alaska, in Goldfarb, R.J., and Miller, L.D., eds., Mineral deposits of Alaska: Economic Geology Monograph 9, p. 120-150.

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.

Otto,B., 2006, Arctic Progress Report: NovaGold Internal Report

Parker, Bradley, 2017, Aquatic Biomonitoring at the Arctic-Bornite Prospect, 2016, Alaska Department of Fish and Game Technical Report No. 17-06, February.

Plafker, G., Jones, D.L., and Pessagno, E.A., Jr., 1977, A Cretaceous accretionary flysch and mélange terrane along the Gulf of Alaska margin in Blean, K.M. ed., The USGS in Alaska Accomplishments during 1976: USGS Survey Circ 751-B, p. B52-B54.

Proffett, J. M., 1999, Summary of Conclusions on Geology of the Arctic Deposit, AK: Kennecott Minerals Company Unpublished Report.

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Rubin, C.M.,1982, Ambler Project Annual Report: Anaconda Metals Company Internal report.

Russell, R. H., 1977, Annual Report, Arctic Deposit: Bear Creek Mining Company Unpublished Report.

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Schmandt, D., 2009, Mineralogy and origin of Zn-rich horizons within the ArcticvVolcanogenic Massive Sulfide deposit, Ambler District, Alaska. Undergraduate Thesis, Smith College, 59 p.

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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.

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SRK Consulting (Canada) Inc. (2018).  Technical Report on the Arctic Project for PFS level Waste Management, Water Management & Closure Design, report prepared for Trilogy Metals Inc., SRK Project Number 1CT030.001, February.

SRK, 2016, Pre-Feasibility Slope Geotechnical and Hydrogeological Report for the Arctic Deposit (2016), SRK Consulting (Canada), Inc., February 2017.

SRK, 2012, NI 43-101 Preliminary Economic Assessment Ambler Project Kobuk, AK, SRK Consulting (U.S.), Inc., March.

SRK, 1998, Arctic Project Preliminary Scoping Study. Prepared by SRK Consulting, U.S., Inc. for NovaCopper. Report Date March 9, 2012. 276 pages.

SRK, 2012, NI 43-101 Preliminary Economic Assessment, Ambler Project, Kobuk, AK. Prepared by SRK Consulting Inc. for Kennecott Minerals Company. Report Date November, 1998. 69 pages.

SRK Consulting (Canada) Inc..  Arctic Project Pre-feasibility Study: Overburden Site Investigation, report prepared for Trilogy Metals Inc., SRK Project Number 1CT030.001, Report Date October, 2017.

SRK Consulting (Canada) Inc..  Technical Report on the Arctic Project for PFS level Waste Management, Water Management & Closure Design, report prepared for Trilogy Metals Inc., SRK Project Number 1CT030.001, Report Date March 2018.

Trilogy Metals Inc.	27-6
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

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.

Stephens, J.D., and Cameron, J.W., 1970, Arctic Alaska Project – Mineralogic Study of Diamond Drill Core Samples: Technical Report 70-01. Kennecott Research Center – Metal Mining Division, 21 p.

Stevens, M.G., 1982, Ambler District Generalized District Petrology: Bear Creek Mining Company Internal report.

Tetra Tech, 2013, Preliminary Economic Assessment Report on the Arctic Project, Ambler Mining District, Northwest Alaska, September.

Tetra Tech, 2011, Arctic Deposit Access Environmental Baseline Data Collection – Aquatics, Ambler Mining District, Alaska, January 20.

Tetra Tech, 2010a, Arctic Deposit Access Environmental Baseline Data Collection - Hydrology, Ambler Mining District, Alaska, December.

Tetra Tech, 2010b, Arctic Deposit Access Environmental Baseline Data Collection - Wetlands & Vegetation, Ambler Mining District, Alaska, November.

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.

Trilogy Metals Inc., 2018: Arctic Pre-Feasibility Study – AMDIAP Road Toll Costs: memorandum prepared for Ausenco and Amec, dated March 27, 2018, 15 p.

Twelker, E., 2008, Progress Report: Immobile element lithogeochem work at Arctic: NovaGold Internal Report

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.

Trilogy Metals Inc.	27-7
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

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.

Vallat, C., 2015, Quality Assurance and Quality Control Review on Analytical Results on the Arctic Project, unpublished report prepared for NovaCopper Inc.

Vallat, C., 2016, Quality Assurance and Quality Control Review on Analytical Results on the Arctic Project, unpublished report prepared for NovaCopper Inc.

Vogl, J. J., Calvert, A. J., Gans, P. B., 2003, 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.

Website: http://alaskamininghalloffame.org/inductees/berg.php, February 14, 2012.

West, A., 2013, Memo to E. Workman & GeoSpark Consulting Inc. “NovaCopper Arctic Project Database Verification”: Internal NovaCopper Inc. Memo, January 8, 2013

West, A., 2014, Identified 2013 Erroneous SG Measurements, internal memo prepared for NovaCopper.

WHPacific, 2016, Cultural Resources Assessment of the Proposed Arctic Pit and Support Facilities Project in Northwest Arctic Borough, Alaska, Consultant Report for Trilogy Metals Inc., December 19.

Zieg, G. A., et al., 2005, Ambler Project 2005 Progress Report, Unpublished Internal Report, Alaska Gold Company.

Trilogy Metals Inc.	27-8
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.0 Certificates of Qualified Persons

28.1 Paul Staples, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

Paul Staples, P.Eng., Ausenco Engineering Canada Inc.

I, L. Paul Staples, P.Eng. am employed as the Vice President and Global Practice Lead, Minerals and Metals with Ausenco Engineering Canada Inc. with a business address of 855 Homer St., Vancouver, B.C., V6B 2W2.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a registered Professional Engineer of New Brunswick, membership number 4832. I graduated from Queens University in 1993 with a degree in Materials and Metallurgical Engineering.

I have practiced my profession for 25 years. I have been directly involved in process operation, design and management from over 15 similar studies or projects including the 80 Mt/y Grasberg complex in Indonesia (1998–2003), the 20 Mt/y Lumwana Project in Zambia (2005–2007), and the 38 Mt/y Dumont feasibility study in Canada (2010–2016).

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

My most recent personal inspection of the Arctic site was on July 25, 2017, for one day.

I am responsible for Sections 1.1, 1.2, 1.8 to 1.10, 1.12, 1.13, 1.16, 1.17; Section 2; Section 3; Section 4, Section 5, Section 6; Section 17; Sections 18.1 to 18.7, 18.11, 18.13 to 18.17; Section 19; Section 21.1.1 to 21.1.6, 21.1.11, 21.2.1, 21.2.3 to 21.2.5; Section 24; Section 25.1, 25.2, 25.8, 25.9, 25.11 to 25.13, 25.15; Section 26; and Section 27.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Paul Staples, P.Eng.

Vice President and Global Practice Lead, Minerals and Metals

Ausenco Engineering Canada Inc.

Trilogy Metals Inc.	28-1
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.2 Justin Hannon, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

Justin Hannon, P.Eng., Ausenco Engineering Canada Inc.

I, Justin Hannon, P.Eng., am employed as a Senior Mining Engineer and Financial Analyst with Ausenco Engineering Canada Inc. with a business address of 855 Homer St., Vancouver, BC, V6B 2W2.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a Professional Engineer of the Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists, registration number L3207. I graduated from Queen’s University (B.Sc Mining Engineering, 2007; B.A. Economics, 2008).

I have practiced my profession for over 7 years. I have been directly involved in the economic evaluation of mining projects, cost modelling, financial analysis, and mine planning. I have been involved in the technical studies of gold and base metal mining projects in Canada and have completed financial analysis due diligence on base metals, gold, and coal mining projects in Australia.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic site.

I am responsible for Sections 1.14 to 1.15; Sections 2.3, 2.5, 2.6; Section 3; Section 22; and Section 25.14 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Justin Hannon, P.Eng.

Senior Mining Engineer and Financial Analyst

Ausenco Engineering Canada Inc.

Trilogy Metals Inc.	28-2
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.3 Antonio Peralta Romero, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

Antonio Peralta Romero, PhD, P.Eng., Amec Foster Wheeler Americas Ltd.

I, Antonio Peralta Romero, P.Eng., am employed as a Principal Mining Engineer with Amec Foster Wheeler Americas Limited at 400-111 Dunsmuir Street, Vancouver, BC V6B 5W3.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a Professional Engineer of Engineers and Geoscientists of British Columbia; License # 45323. I graduated from the University of Guanajuato in 1984 with a B.S. in Mining Engineering, from Queen’s University in 1991 with a M.Sc. in Mining Engineering, and from Colorado School of Mines in 2007 with a Ph.D. in Mining and Earth Systems Engineering.

I have practiced my profession for 33 years. I have been directly involved in mine planning and design, ore control, production forecasting and management, slope stability monitoring, and mineral reserve estimation, mainly for open-pit precious, base metal mines and iron ore mines.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects (NI 43–101).

I visited the Arctic site on July 25, 2017 for one day.

I am responsible for Section 1.7, Sections 2.3 to 2.6; Section 3; Section 15; Sections 16.1 to 16.7, 16.9; Sections 21.1.7, 21.1.8, 21.2.2; and Section 25.7.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Antonio Peralta Romero, P.Eng.

Principal Mining Engineer

Amec Foster Wheeler Americas Limited

Trilogy Metals Inc.	28-3
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.4 Bruce M. Davis, FAusIMM

CERTIFICATE OF QUALIFIED PERSON

Bruce M. Davis, FAusIMM, BD Resource Consulting, Inc.

I, Bruce M. Davis, FAusIMM, do hereby certify that:

1. I am an independent consultant of:

BD Resource Consulting, Inc.

4253 Cheyenne Drive

Larkspur, Colorado USA 80118

2. I graduated from the University of Wyoming with a Doctor of Philosophy (Geostatistics) in 1978.

3. I am a Fellow of the Australasian Institute of Mining and Metallurgy, Number 211185.

4. I have practiced my profession continuously for 40 years and have been involved in mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia, Africa and Australia.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6. I am the author of sections in the technical report titled, “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”), and accept professional responsibility for Sections 1.3, 1.4, 1.6, 2.3 to 2.6, 11, 12, 14, 25.4, and 27.

7. I visited the Arctic Property on 26-27 July 2011 and again on 25 September 2012 and again on 10-12 August 2015.

8. I was the author of a previous Technical Report dated 9 November 2017.

9. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10. I am independent of Trilogy Metals Inc. applying all of the tests in Section 1.5 of NI 43-101.

11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 6^th day of April, 2018

“original signed and sealed”

Bruce M. Davis, FAusIMM

President

BD Resource Consulting, Inc.

Trilogy Metals Inc.	28-4
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.5 John Joseph DiMarchi, CPG

CERTIFICATE OF QUALIFIED PERSON

John Joseph DiMarchi, CPG, Core Geoscience Inc.

I, John Joseph DiMarchi, CPG, am employed as a Principal Consultant with Core Geoscience LLC. located at 5319 NE 62^nd Street, Seattle, WA 98115.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a Member in Good Standing of the American Institute of Professional Geologists CPG #9217. I graduated from Colorado State University with a BS Geology in 1978.

I have practiced my profession for 30 years. I have been directly involved in mine permitting and environmental work in Alaska as a consultant, industry geologist and as large mine permitting coordinator for the State of Alaska, Department of Natural Resources since 2007. I also served as QP for environmental portions of the Preliminary Economic Assessment of the Arctic Project in 2013.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic site.

I am responsible for Sections 1.11.1 to 1.11.3, 2.3, 2.6, 20.1, 20.1.1 to 20.1.8, 20.1.10, 20.2, 20.3, 20.4.2, 25.10, and 27 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

My only previous involvement with the Arctic project was acting as QP for portions of the Preliminary Economic Assessment Technical Report in 2013.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

John J. DiMarchi, CPG

Principal Consultant

Core Geoscience

Trilogy Metals Inc.	28-5
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.6 Jeffery B. Austin, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

Jeffery B. Austin, P.Eng., International Metallurgical & Environmental Inc.

I, Jeffery B. Austin, P.Eng., am employed as President of International Metallurgical and Environmental Inc. located at #13-2550 Acland Road, Kelowna, B.C. V1X7L4, Canada.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a registered, professional engineer of the Engineers and Geoscientists British Columbia, license number 15708. I graduated from the University of British Columbia in 1984 with a Bachelors of Applied Science specializing in Mineral Process Engineering.

I have practiced my profession for 34 years. I have been directly involved in the development of metallurgical processes, supervision of test work and research programs and mineral processing operations.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic project site.

I am responsible for Section 1.5, Sections 2.3, 2.4, 2.6; Section 13; Section 25.5; and Section 26.2 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have been involved with the Arctic Project as an independent consultant since 2011

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Jeffrey B. Austin, P.Eng. – President

International Metallurgical and Environmental Inc.

Trilogy Metals Inc.	28-6
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.7 Robert Sim, P.Geo.

CERTIFICATE OF QUALIFIED PERSON

Robert Sim, P.Geo., SIM Geological Inc.

I, Robert Sim, P.Geo., am an independent consultant of SIM Geological Inc. 508 – 1950 Robson St. Vancouver, BC, Canada V6G 1E8.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a member, in good standing, of Engineers and Geoscientists British Columbia, License Number 24076. I graduated from Lakehead University with an Honours Bachelor of Science (Geology) in 1984.

I have practiced my profession continuously for 34 years and have been involved in mineral exploration, mine site geology and operations, mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia, Africa and Australia. I have worked on similar VMS-type deposits at Cayeli in Turkey, the Winston Lake mine in Ontario and on several deposits in Rouyn-Noranda, Quebec.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic Property.

I am a co-author of the Technical Report, I am responsible for Sections 1.3, 1.4, 1.6; Sections 2.3, 2.5, 2.6; Section 7; Section 8; Sections 9.0, 9.1, 9.2, 9.3, 9.4, 9.5, 9.6; Section 10; Section 14; Section 23; Sections 25.3, 25.6; and portions of Section 27.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have had prior involvement with the property that is the subject of the Technical Report. I was an author of a previous technical report titled NI 43-101 Technical Report on the Arctic Project, Northwest Alaska, USA, dated November 9, 2017.

I have read NI 43–101 and the sections of the Technical Report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the Technical Report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Robert Sim, P.Geo.

SIM Geological Inc.

Trilogy Metals Inc.	28-7
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.8 Calvin Boese, PEng.

CERTIFICATE OF QUALIFIED PERSON

Calvin Boese, P.Eng., M.Sc., SRK Consulting (Canada) Inc.

I, Calvin Boese, PEng., am employed as a Senior Consultant (Geotechnical Engineering) with the firm of SRK Consulting (Canada) Inc. (SRK) with an office at Suite 205 – 2100 Airport Drive, Saskatoon, Saskatchewan.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a Professional Engineer of the Association of Professional Engineers, Geologists of British Columbia (P.Eng. #29478). I am also a registered Professional Engineer in Alberta and Saskatchewan. I am a Member of the Society for Mining, Metallurgy and Exploration. I am a graduate of the University of Saskatchewan with a B.Sc. in Civil Engineering (1999) and a M.Sc. in Geo-Environmental Engineering (2004).

I have practiced my profession for 18 years. I have been directly involved in geotechnical aspects of mining, including the site selection, design, permitting, operation and closure of mine waste facilities in Canada, the US, Indonesia and Turkey.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

My most recent personal inspection of the Arctic site was from July 24 to July 25, 2017.

I am responsible for Section 1.11.4, Sections 18.10 and 18.12, Section 20.4.1, Section 21.1.9 and co-authored Sections 2.3, 2.4, and 2.5; Section 26 and 27 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Calvin Boese, PEng

Senior Consultant

SRK Consulting (Canada) Inc.

Trilogy Metals Inc.	28-8
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.9 Bruce Murphy, PEng.

CERTIFICATE OF QUALIFIED PERSON

Bruce Murphy, P.Eng., SRK Consulting (Canada) Inc.

I, Bruce Murphy, PEng am employed as a Principal Consultant at SRK Consulting (Canada) Inc, located at 2200 -1066 West Hastings Street, Vancouver, BC.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February 2018 (the “Technical Report”).

I am a professional engineer registered with the Association of Professional Engineers of British Columbia – PEng License No.: 44271; I am a graduate of University of the Witwatersrand, Johannesburg, South Africa with a M.Sc. degree in Mining Engineering. I have practiced my profession continuously since graduation (1989)working in the rock engineering field on operating mines till 2002 and then in the consulting field.

I have practiced my profession for 28 years. I have been directly involved in the geotechnical data acquisition, characterization and slope stability evaluation of the deposit.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic site.

I am responsible for Sections 9.7.0, 9.7.1; Section 16.8 and co-author of Sections 2.3, 2.4, and 2.5; Section 26; and Section 27 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Bruce Murphy, PEng

Principal Consultant

SRK Consulting (Canada) Inc.

Trilogy Metals Inc.	28-9
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

28.10 Tom Sharp, PEng.

CERTIFICATE OF QUALIFIED PERSON

Tom Sharp, PhD, P.Eng., SRK Consulting (Canada) Inc.

I, Tom Sharp, PEng, am employed as a Principal Consultant (Water Management and Treatment Engineering) with the firm of SRK Consulting (Canada) Inc. (SRK) with an office at Suite 2200 – 1066 West Hastings Street, Vancouver, British Columbia.

This certificate applies to the technical report titled “Arctic Project, Northwest Alaska, USA, NI 43-101 Technical Report on Pre-Feasibility Study” that has an effective date of 20 February, 2018 (the “Technical Report”).

I am a Professional Engineer of the Association of Professional Engineers, Geologists of British Columbia (P.Eng. #36988). I am also a registered Professional Engineer in Northwest Territories, Nunavut, Yukon and Montana. I am a Member of the Society for Mining, Metallurgy and Exploration. I graduated from Montana State University and Montana Tech with a B.Sc. and M.Sc. in Biological Sciences (1988 and 1993), M.Sc. in Environmental Engineering (1996) and a Ph.D. in Civil Engineering (1999).

I have practiced my profession for 24 years. I have been directly involved mine water management and treatment on projects in North America, South America, Europe and Asia.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101Standards of Disclosure for Mineral Projects(NI 43–101).

I have not visited the Arctic site.

I am responsible for Sections 9.7.2, 9.7.3; Section 11.7.2; Sections 18.8, 18.9; Section 20.1.9; Section 21.1.10 and co-authored Sections 2.3, 2.4, and 2.5; Section 26 and Section 27 of the technical report.

I am independent of Trilogy Metals Inc. as independence is described by Section 1.5 of NI 43–101.

I have no previous involvement with the Arctic Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the sections of the technical report for which I am responsible contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated this 6^th day of April, 2018

“original signed and sealed”

Tom Sharp, PEng

Principal Consultant

SRK Consulting (Canada) Inc.

Trilogy Metals Inc.	28-10
NI 43-101 Technical Report on the Arctic Project,
Northwest Alaska

 

Appendix A – List of Claims

Trilogy Metals Inc.

NI 43-101 Technical Report on the Arctic Project,

Northwest Alaska