Polymet Mining (PLM) 6-K Form NI 43-101F1 Technical Report

Exhibit 99.1

Form NI 43-101F1 Technical Report Minnesota, USA Zachary J. Black, Jennifer J. Brown, Nicholas Dempers, Thomas L. Drielick, Art S. Ibrado, Erin L. Patterson, Thomas J. Radue, Jeff S. Ubl and Herbert E. Welhener Prepared For:

NorthMet Project

Form NI 43-101F1 Technical Report

DATE AND SIGNATURES PAGE

This report is effective as of March 26, 2018. The Technical Report Contributors’ Professional Qualifications and certificates are attached as Appendix A. These certificates are considered the date and signature of this report in accordance with Form NI 43-101F1.

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Note:  This Report contains “forward-looking statements”.  Within the meaning of applicable Canadian securities legislation and Section 27A of the United States Securities Act of 1933 and Section 21E of the United States Securities Exchange Act of 1934, forward-looking statements are not, and cannot be, a guarantee of future results or events. Forward looking statements are based on, among other things, opinions, assumptions, estimates and analyses that are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking statement. All statements in this Report that address events or developments that PolyMet expects to occur in the future are forward-looking statements and are generally, although not always, identified by words such as “expect”, “plan”, “anticipate”, “project”, “target”, “potential”, “schedule”, “forecast”, “budget”, “estimate”, “intend” or “believe” and similar expressions or their negative connotations, or that events or conditions “will”, “would”, “may”, “could”, “should” or “might” occur. These forward-looking statements include, but are not limited to, PolyMet’s objectives, strategies, intentions, expectations, production, costs, capital and exploration expenditures, including an estimated economics of future financial and operating performance and prospects for the possible expansion of the operation based on a PEA-level study and a ramp-up evaluation representing production growth and improved margins mine, life projections, recovery rate and concentrate grade projections, ability to obtain all necessary environmental and government approvals to completion and if undertaking an expansion case, ability to obtain at all, the viability and all information with respect to the ability to develop the Project to additional potential by mining additional resources beyond the permit design at a higher production rate. Prior to any decision to apply for permits to develop the project further, PolyMet would need to complete preliminary and definitive feasibility studies, as well as an analysis of the environmental impact and alternatives of any proposal.  In addition, any future proposal would be subject to environmental review and permits, public notice and comment, and approval by appropriate federal and state Agencies. All forward-looking statements in this Report are qualified by this cautionary note.

The material factors or assumptions that PolyMet has identified and were applied by PolyMet in drawing the conclusions or making forecasts or projections set in the forward-looking statements include, but are not limited to:

· various economic assumptions, in particular, metal price estimates, set out in Section 22 of this Report and elsewhere;

· certain operational assumptions set out in the Report, including mill recovery, operating scenarios;

· construction schedules and timing issues; and

· assumptions concerning timing and certainty regarding the environmental review and permitting process.

The risks, uncertainties, contingencies and other factors that may cause actual results and events to differ materially from those expressed or implied by the forward-looking statement may include, but are not limited to, risks generally associated with the mining industry, such as: economic factors (including future commodity prices, currency fluctuations, inflation rates, energy prices and general cost escalation); uncertainties related to the development of the NorthMet Project; dependence on key personnel and employee relations; risks relating to political and social unrest or change, operational risk and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks; failure of plant, equipment, processes, transposition and other infrastructure to operate as anticipated; compliance with governmental and environmental regulations, including permitting requirements; etc., as well as other factors identified and as described in more detail under the heading “Risk Factors” in PolyMet’s most recent Annual Information Form, which may be viewed on www.sedar.com and sec.gov. The list is not exhaustive of the factors that may affect the forward-looking statements. There can be no assurance that such statements will prove to be accurate, and actual results, performance or achievements could differ materially from those expressed in, or implied by, these forward-looking statements. Accordingly, no assurance can be given that any events anticipated by the forward-looking statements will transpire or occur, or if any of them do, what benefits or liabilities PolyMet will derive therefrom. The forward-looking statements reflect the current expectations regarding future events and operating performance and speak only as of the date hereof and PolyMet does not assume any obligation to update the forward-looking statements if circumstances or management’s beliefs, expectations or opinions should change other than as required by applicable law. For the reasons set forth above, undue reliance should not be placed on forward-looking statements.

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Cautionary Note to U.S. Investors – Information Concerning Preparation of Resource Estimates

This Report has been prepared in accordance with the requirements of the securities laws in effect in Canada, which differ from the requirements of United States Securities laws. The terms “mineral reserve”, “proven mineral reserve” and “probable mineral reserve” are Canadian mining terms as defined in accordance with Canadian National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and the Canadian Institute of Mining Metallurgy and Petroleum (the “CIM”) – CIM Definition Standards on Mineral Resources and Mineral Reserves, adopted by the CIM Council, as amended. These definitions differ materially from the definitions in the United States Securities and Exchange Commission’s (“SEC”) Industry Guide 7 under the United States Securities Act of 1933, as amended. Under SEC Industry Guide 7 standards, mineralization cannot be classified as a “reserve” unless the determination has been made that the mineralization could be economically and legally extracted at the time the reserve determination is made. As applied under SEC Industry Guide 7, a “final” or “bankable” feasibility study is required to report reserves, the three-year historical average price is used in any reserve or cash flow analysis to designate reserves, and the primary environmental analysis or report must be filed with the appropriate governmental authority.

In addition, the terms “mineral resource”, “measured mineral resource”, “indicated mineral resource” and “inferred mineral resource” are defined in and required to be disclosed by NI 43-101; however, these terms are not defined terms under SEC Industry Guide 7 and are normally not permitted to be used in reports and registration statements filed with the SEC. Investors are cautioned not to assume that all or any part of a mineral deposit in these categories will ever be converted into SEC Industry Guide 7 reserves. “Inferred mineral resources” have a great amount of uncertainty as to their existence, and great uncertainty as to their economic and legal feasibility. It cannot be assumed that all or any part of an inferred mineral resource will ever be upgraded to a higher category. Under Canadian rules, estimates of inferred mineral resources may not form the basis of feasibility or pre-feasibility studies, except in rare cases. Investors are cautioned not to assume that all or any part of an inferred mineral resource exists or is economically or legally mineable. Disclosure of “contained metal” in a resource is permitted disclosure under Canadian regulations; however, the SEC normally only permits issuers to report mineralization that does not constitute “reserves” by SEC Industry Guide 7 standards as in place tonnage and grade without reference to unit measures.

Accordingly, information concerning mineral deposits contained in this Report may not be comparable to similar information made by public U.S. companies subject to the reporting and disclosure requirements under the United States federal securities laws and the rules and regulations thereunder.

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NORTHMET PROJECT
FORM NI 43-101F1 TECHNICAL REPORT

TABLE OF CONTENTS

SECTION PAGE

DATE AND SIGNATURES PAGE I

TABLE OF CONTENTS IV

LIST OF FIGURES AND ILLUSTRATIONS XIII

LIST OF TABLES XV

1 EXECUTIVE SUMMARY 19

1.1 Key Results 19

1.1.1 Project Phases 19

1.1.2	Key Results for Both Phases	20
1.1.3	Phase I Key Results at 32,000 STPD	20
1.1.4	Phase II Key Results at 32,000 STPD	21

1.2 Location AND Ownership 21

1.3 Geology And Mineralization 21

1.4 Status of Exploration 21

1.5 Mineral Reserve Statement 22

1.6 Mineral Resource Estimate 22

1.7 Mining And Processing 23

1.8 Environmental 24

1.9 Economics 24

1.10 Potential Expansion Opportunities – Basis of 59,000 STPD and 118,000 STPD Scenarios 26

1.11 Conclusions And Recommendations 27

2 INTRODUCTION 28

2.1 Purpose 28

2.2 Sources of Information 28

2.3 Terms of Reference 29

2.4 Units of Measure 31

3 RELIANCE ON OTHER EXPERTS 32

4 PROPERTY DESCRIPTION AND LOCATION 33

4.1 Project Location 33

4.2 Project Ownership 34

4.3 Mineral Tenure 34

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4.4 Surface Rights 35

4.5 Royalties And Encumbrances 35

4.6 Environmental Liabilities 35

4.7 Permits 36

4.8 Social License 36

4.9 Significant Risk Factors 36

4.9.1 Permitting 36

4.9.2	Project Financing	36
4.9.3	Commodity Prices	37

4.10 Comments on Section 4 37

5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 38

5.1 Accessibility 38

5.2 Climate 38

5.3 Local Resources and Infrastructure 39

5.4 Physiography 39

5.5 Sufficiency of Surface Rights 39

6 HISTORY 40

6.1 Ownership 40

6.2 Exploration and Sampling 40

6.3 Historical Mineral Resource and Reserve Estimates 40

6.4 Historical Production 40

7 GEOLOGICAL SETTING AND MINERALIZATION 41

7.1 Regional Geology 41

7.1.1 Felsic Series 41

7.1.2	Early Gabbro Series	41
7.1.3	Anorthositic Series	41
7.1.4	Layered Series	41

7.2 LOCAL AND PROPERTY GEOLOGY 44

7.2.1 Local Lithology 45

7.2.2 Unit Definitions and Descriptions 47

8 DEPOSIT TYPES 51

9 EXPLORATION 52

9.1 Geophysical Sounding 52

9.2 U.S. Steel Bulk Sampling 52

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9.3 Down-Hole Geophysical Testing 53

10 DRILLING 54

10.1 Introduction 54

10.2 Historic Drilling 56

10.2.1 U.S. Steel Drilling, 1969 – 1974 56

10.2.2 NERCO Drilling 1991 56

10.3 PolyMet Drilling 56

10.3.1	PolyMet Drilling, 1999 – 2000, Reverse Circulation Holes	56
10.3.2	PolyMet Drilling, 1999-2000, Diamond Core Holes	56
10.3.3	PolyMet Drilling, 2005, Diamond Core Holes	57
10.3.4	PolyMet Drilling, 2007, Diamond Core Holes	57
10.3.5	PolyMet Drilling, 2010, Diamond Core Holes	57

10.4 Relevant Results And Interpretation 58

11 SAMPLE PREPARATION, ANALYSES AND SECURITY 59

11.1 Historic Sample Preparation, Analysis and Security 59

11.1.1	U.S. Steel and NERCO	59
11.1.2	PolyMet Sample Preparation, Analysis and Security	60
11.1.3	Sample Preparation	61

11.2 Analytical History 61

11.2.1	Base Metals	61
11.2.2	Platinum Group Elements	62
11.2.3	Total Sulfur	62

11.3 Quality Assurance/Quality Control Procedures 62

11.3.1	Blanks	63
11.3.2	Standards	64
11.3.3	Duplicates	66

11.4 Data Entry Validation Controls 69

11.5 Core Storage And Sample Security 69

11.6 Opinion On Adequacy 69

12 DATA VERIFICATION 70

12.1 PolyMet Data Compilation And Verification 2004 70

12.2 Database Audit 70

12.3 Certificates 71

12.4 Adequacy Of Data 71

13 MINERAL PROCESSING AND METALLURGICAL TESTING 72

13.1 Introduction 72

13.2 Comminution Circuit Test Work And Process Development 74

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13.3 Flotation Circuit Test Work And Process Development 75

13.4 Flotation Circuit Design 80

13.4.1 Flotation Circuit Simulation 82

13.5 Metallurgical Modelling For Recovery And Concentrate Quality 84

13.5.1 Cobalt 85

13.6 Hydrometallurgical Test Work 89

13.6.1	PLATSOL™ Leaching Pilot Plant Testing	90
13.6.2	Precipitation of PGMs by Copper Sulfide	92
13.6.3	Copper Concentrate Enrichment	93
13.6.4	Residual Copper Precipitation	95
13.6.5	Bulk Iron/Aluminum Removal	96
13.6.6	Mixed Hydroxide Precipitation (MHP)	97
13.6.7	Magnesium Removal	98

14 MINERAL RESOURCE ESTIMATES 99

14.1 Data 99

14.2 Block Model Physical Limits 99

14.3 Geological Models 99

14.3.1 Density 101

14.4 Exploratory Data Analysis 101

14.4.1	Sample Statistics	101
14.4.2	Correlation Analysis	104
14.4.3	Contact Plot Analysis	106

14.5 Estimation Methodology 110

14.5.1	Capping	110
14.5.2	Composite Study	113
14.5.3	Variograms	116
14.5.4	Estimation Strategy	118
14.5.5	Mineral Resource Classification	119
14.5.6	Model Validation	120

14.6 Mineral Resources 127

14.6.1	Net Smelter Return (NSR) and Cutoff	127
14.6.2	Test for Reasonable Prospect for Eventual Economic Extraction	127
14.6.3	Resource Statement	128

15 MINERAL RESERVE ESTIMATES 129

15.1 Calculation Parameters 129

15.1.1	Pit Slopes	129
15.1.2	Dilution and Mining Losses	129
15.1.3	Cutoff and NSR Calculation	130

15.2 Mineral Reserve Estimate 132

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15.3 Factors that may Affect the Mineral Reserve Estimate 132

16 MINING METHODS 133

16.1 Open Pit Mine Plan 133

16.2 Resource Model Review 135

16.3 Definition Of Material Types 135

16.3.1	Ore Classification	135
16.3.2	Waste Rock Classification	135
16.3.3	Waste Rock Stockpile Liners	136
16.3.4	Overburden Classification	136

16.4 Geotechnical 137

16.5 Pit Design 137

16.6 Preproduction Development 138

16.7 Production Schedule 138

16.7.1	Yearly Production	139
16.7.2	Pit and Stockpile Progression Maps	142

16.8 Water Management System 145

16.9 Mining Equipment 147

16.9.1	Production Schedule Parameters	147
16.9.2	Drill Equipment and Blast Parameters	148
16.9.3	Loading Equipment Requirements	149
16.9.4	Hauling Equipment Requirements	150
16.9.5	Auxiliary Equipment Requirements	150

16.10 Railroad 151

16.11 Mine Personnel 152

17 RECOVERY METHODS 155

17.1 Plant Design 155

17.1.1	Introduction	155
17.1.2	Crushing and Material Handling	156
17.1.3	Flotation	159
17.1.4	Concentrate Handling	161
17.1.5	Reagent Services	161
17.1.6	Piping Systems	162
17.1.7	Air Systems	163
17.1.8	Plant Electrical Distribution	164
17.1.9	Plant Instrumentation	164

17.2 Process Plant Flowsheet Development 164

17.2.1	Primary and Secondary Crushing	164
17.2.2	Milling	165
17.2.3	Flotation	168
17.2.4	Tailings Disposal	171

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17.2.5	Concentrate Thickening and Filtration	172
17.2.6	Concentrate Storage	174
17.2.7	Reagents	174
17.2.8	Air Services	176
17.2.9	Water Circuits	176
17.2.10	Sampling and Metal Accounting	176

17.3 HYDROMETALLURGICAL PROCESSING 177

17.4 PHASE II – OPTIONAL HYDROMETALLURGICAL PLANT 177

17.4.1	Autoclave	180
17.4.2	Gold and Platinum Group Metals Recovery	180
17.4.3	Concentrate Enrichment	181
17.4.4	Copper Sulfide Precipitation	181
17.4.5	Iron, Aluminum and Acid Removal	181
17.4.6	Mixed Hydroxide Precipitation Recovery	182
17.4.7	Magnesium Removal	182
17.4.8	Process Consumables	182
17.4.9	Hydrometallurgical Plant Water	183
17.4.10	Metal Recoveries	183

17.5 PLANT SITE AIR QUALITY MANAGEMENT 183

17.5.1	Hydrometallurgical Residue Management	184
17.5.2	Hydrometallurgical Residue Cell Design and Operations	184

17.6 WATER MANAGEMENT  184

17.6.1 Hydrometallurgical Plant 184

18 PROJECT INFRASTRUCTURE 185

18.1 PLANT AND ADMINISTRATIVE INFRASTRUCTURE 185

18.1.1	Asset Preservation	185
18.1.2	Plant Workshops	185
18.1.3	Plant Warehouses	185
18.1.4	Administration Offices	186
18.1.5	Site First Aid Station	186
18.1.6	Laboratory	186

18.2 Mine Infrastructure 186

18.2.1	Mine Workshops, Warehouses and Offices	186
18.2.2	Mine Site Service and Refueling Facility	186
18.2.3	Rail Loadout	186

18.3 Haul and Access Roads 187

18.4 Rail Facilities 187

18.5 Water Supply 187

18.5.1	Raw Water Supply	187
18.5.2	Potable Water Distribution	187
18.5.3	Fire Water Distribution	187
18.5.4	Sewage Collection and Treatment	187

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18.6 Flotation Tailings Basin (FTB) 187

18.7 Waste Water Treatment 188

18.7.1	Mine Site Waste Water Collection and Distribution	188
18.7.2	Waste Water Treatment System	188

18.8 Power Supply 189

18.8.1	Plant Power Supply	189
18.8.2	Mine Site Power Supply	189
18.8.3	Emergency Power Plant	189

18.9 Natural Gas Supply 190

18.10 Accommodations 190

19 MARKET STUDIES AND CONTRACTS 191

19.1 Commodity Price Projections 191

19.2 Contracts 191

20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT 192

20.1 Environmental Review and Permitting 192

20.2 Baseline Studies 193

20.3 Environmental Considerations 193

20.3.1	Waste Management	194
20.3.2	Water Management	194
20.3.3	Air Management	195
20.3.4	Land Management	195

20.4 Social Issues 195

20.4.1	Labor and Employment Support	195
20.4.2	Economic Impact	195
20.4.3	Treaties and Indigenous Groups	196

20.5 Closure Plan and Financial Assurance 196

20.6 Discussion on Permitting Risks to Mineral Resources and Mineral Reserves 197

20.7 Comments on Section 20 197

21 CAPITAL AND OPERATING COSTS 198

21.1 Capital Cost Estimate 198

21.1.1	Basis of Phase I Capital Cost Estimate	199
21.1.2	Hydrometallurgical Plant Cost Estimate	206
21.1.3	Indirect Costs	207

21.2 Operating Cost Estimate 210

21.2.1	Mine Operating Cost	210
21.2.2	Process Plant and Assay Operating Cost Estimate Summary	213
21.2.3	Basis of Process Plant Operating Cost Estimate	213
21.2.4	Hydrometallurgical Plant (Phase II) Operating Cost Estimate Summary	222

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21.2.5 Basis of Hydrometallurgical Plant (Phase II) Operating Cost 222

22 ECONOMIC ANALYSIS 224

22.1 FEASIBILITY STUDY ECONOMIC ANALYSIS 224

22.1.1	Economic Assumptions	225
22.1.2	Key Data and Economic Analysis	225
22.1.3	Capital Costs	228
22.1.4	Operating Plans and Costs	228
22.1.5	Sustaining Capital	228

22.2 FINANCIAL MODEL 229

23	ADJACENT PROPERTIES	231
24	OTHER RELEVANT DATA AND INFORMATION	232

24.1 PROJECT IMPLEMENTATION 232

24.1.1	Engineering	232
24.1.2	Demolition	234
24.1.3	Execution and Construction	234
24.1.4	Ramp-up Evaluation	235

24.2 POTENTIAL OPPORTUNITIES 239

24.2.1	Summary	239
24.2.2	Introduction	239
24.2.3	Mill Throughput Tonnages	240
24.2.4	Mine Site Modifications	241
24.2.5	Plant Site Modifications	242
24.2.6	Financial Outlook	243

24.3 ECONOMIC ANALYSIS 244

25 INTERPRETATION AND CONCLUSIONS 249

25.1	INTRODUCTION	249
25.2	INTERPRETATION	249

25.2.1	Surface Rights, Royalties, and Mineral Tenure	249
25.2.2	Geology and Mineralization	249
25.2.3	Exploration	249
25.2.4	Drilling and Sampling	249
25.2.5	Data Verification	249
25.2.6	Metallurgy	250
25.2.7	Mineral Resources	250
25.2.8	Mineral Reserves	250
25.2.9	Mine Plan and Schedule	250
25.2.10	Metallurgical Recovery	250
25.2.11	Infrastructure	250
25.2.12	Market Studies and Contracts	250
25.2.13	Environment, Permits, and Social and Community Impacts	251
25.2.14	Capital and Operating Costs	251
25.2.15	Economic Analysis	251

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25.3	Conclusions	251
25.4	Risks	251
25.5	Opportunities	254

26	RECOMMENDATIONS	256
27	REFERENCES	257
APPENDIX A – PROFESSIONAL QUALIFICATIONS AND CERTIFICATES OF QUALIFIED PERSONS		261

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LIST OF FIGURES AND ILLUSTRATIONS

FIGURE	DESCRIPTION	PAGE
Figure 4-1: Property Layout Map		33
Figure 7-1: Regional Geology		42
Figure 7-2: Copper-Nickel Deposits in the Duluth Complex (after Severson)		43
Figure 7-3: NorthMet Stratigraphic Column (after Geerts, 1994)		45
Figure 7-4: NorthMet Property Bedrock Geology		46
Figure 7-5: NorthMet “Magenta Zone” in Cross Section		50
Figure 10-1: Drill-hole Collar Location by Campaign		55
Figure 11-1: Copper Blank Analysis		64
Figure 11-2: Nickel Blank Analysis		64
Figure 11-3: Copper Results for Standard 4-1		65
Figure 11-4: Nickel Results for Standard 4-1		65
Figure 11-5: Copper and Nickel ¼ Core Duplicate Analysis		66
Figure 11-6: Copper and Nickel 1/8 Core Duplicate Analysis		67
Figure 11-7: Copper Pulp Duplicate Analysis		68
Figure 11-8: Nickel Pulp Duplicate Analysis		68
Figure 13-1: NorthMet Process Block Flow		73
Figure 13-2: Pilot Plant Flowsheet		77
Figure 13-3: Comparative Recoveries between C9 Pilot Work and Previous Pilot Work		78
Figure 13-4: General Block Flow – Rate Tested and Kinetic-Derived Process Streams from Report NM 1-2015 NorthMet Feb 2015		81
Figure 13-5: Total Cu Recovery vs. Cu Head		86
Figure 13-6: Total Ni Recovery vs. Ni Head		86
Figure 13-7: Total Co Recovery vs. Co Head		86
Figure 13-8: Final Tail S Assay vs. S Head		86
Figure 13-9: Total Pt Recovery vs. Pt Head		87
Figure 13-10: Total Pd Recovery vs. Pd Head		87
Figure 13-11: Total Au Recovery vs. Au Head		87
Figure 13-12: Total Ag Recovery vs. Ag Head		87
Figure 13-13: Hydrometallurgical Pilot Plant Flowsheet		89
Figure 13-14: ACD Liquor Ni, Cu, Mg PLS Trends		91
Figure 13-15: ACD Residue Trends		91
Figure 13-16: Correlation between Cu and ORP Observed for Copper Enrichment Trials		94

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Figure 13-17: Correlation Between ORP and Soluble Copper Concentration	96
Figure 14-1: Estimation Domains	100
Figure 14-2: Copper Correlation Plots	105
Figure 14-3: Contact Plot Virginia Formation and Unit 1	106
Figure 14-4: Contact Plot Unit 1 and Unit 3	107
Figure 14-5: Contact Plot Unit 3 and Unit 5	108
Figure 14-6: Contact Plot Unit 5 and Unit 6	109
Figure 14-7: Contact Plot Unit 6 and Unit 7	110
Figure 14-8: Tukey Box Plots for Unit 1	112
Figure 14-9: Copper Composite Study	113
Figure 14-10: North – South Section Looking East Displaying the Dynamic Search Ellipses	119
Figure 14-11: Model Comparison Cumulative Frequency Plot (NN red, ID blue, Composites Black, OK Green)	123
Figure 14-12: Domain 1 Copper Swath Plot Along Rotated Easting	124
Figure 14-13: Copper Cross Section Along Rotated Easting	125
Figure 14-14: Copper Long Section Along Rotated Northing	126
Figure 14-15: Copper Plan Section	126
Figure 16-1: Mine Site Layout	134
Figure 16-2: Pit Shell Map – End of Year 1	143
Figure 16-3: Pit Shell Map – End of Year 20	144
Figure 16-4: Process Plant Water Balance	146
Figure 17-1: Plant Aerial View	156
Figure 17-2: Current Concentrator Arrangement	158
Figure 17-3: Milling Circuit	159
Figure 17-4: Flotation Circuit	160
Figure 17-5: Overall Plant Process Flow Diagram	167
Figure 17-6: Phase I & II - Overall Plant Process Flow Diagram, Highlighting the Hydrometallurgical Plant Section	179
Figure 24-1: Project Execution Schedule Summary	236
Figure 24-2: Project Execution Schedule Summary Continued	237
Figure 24-3: Project Critical Path	238

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LIST OF TABLES

TABLE	DESCRIPTION	PAGE
Table 1-1: Mineral Reserve Statement – January 2018		22
Table 1-2: Summary Mineral Resource Statement for the NorthMet Project Inclusive of Mineral Reserves		23
Table 1-3: Capital Expenditure & Operating Costs – Phase I and Phase I & II		25
Table 1-4: Price Assumptions in the Financial Analysis		25
Table 1-5: Financial Summary – 32,000 STPD		26
Table 2-1: List of Qualified Persons		28
Table 2-2: Units, Terms and Abbreviations		29
Table 5-1: Current Iron Ore Mines on the Mesabi Iron Range		38
Table 10-1: NorthMet Project Drill Hole Summary		54
Table 10-2: Summary of Core Recoveries and RQD Measurements (includes all drilling through 2010)		58
Table 11-1: Detection Limits of Elements		62
Table 11-2: Detection Limits		62
Table 11-3: Details of Sampling of U.S. Steel Core by PolyMet		63
Table 12-1: Summary of the Analytical Data Used in the Estimation of Mineral Resources		71
Table 13-1: Summary of Comminution Test Work Results		74
Table 13-2: Summary of SMC Test Work Results Conducted by Hazen Research		75
Table 13-3: Milling Circuit Design		75
Table 13-4: Summary of Pilot Plant Test Work Results on Sample C9		77
Table 13-5: Projected Metallurgy of Cu-Ni Separation LCT of C9 Pilot Cleaner Concentrate		79
Table 13-6: Summary of Laboratory Test Work Results on Sample C10		79
Table 13-7: Flotation Stage Design Parameters		80
Table 13-8: Flotation Plant Simulation and Design Parameters		82
Table 13-9: Summary of Flotation Circuit Simulation		83
Table 13-10: NorthMet Tank Cell Sizing and Selection		84
Table 13-11: Summary of C-9 and C-10 Metallurgy Compared to Model		88
Table 13-12: Flotation Concentrate Head Assays Used in the Test Campaigns (C1 & C2)		90
Table 13-13: Average Autoclave Feed Flowrates		90
Table 13-14: Base Metal and PGM Recoveries		92
Table 13-15: Summary of PGM Precipitation Operating Parameters		92
Table 13-16: Comparison between PGM Precipitation Circuit Feed and Filtrate Concentrations		93
Table 13-17: Operating Conditions and Feed Parameter for Copper Concentrate Enrichment		94

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Table 13-18: Head and Copper Enrichment Solids and Filtrate Composite Assays	95
Table 13-19: Soluble Copper Precipitation Parameters	95
Table 13-20: NaHS Product Filter Cake Assays	96
Table 13-21: Test Conditions, Target Ni Concentrations and Ni and Co Feed Concentrations for MHP Tests	97
Table 13-22: MHP Stage 1 Final Product Analysis and Distribution for Campaigns C1 & C2	97
Table 13-23: Test Conditions for Bulk Magnesium Removal	98
Table 14-1: Specific Gravity Average per Unit (October 15 Dataset)	101
Table 14-2: Copper Sample Statistics	101
Table 14-3:Nickel Sample Statistics	102
Table 14-4: Platinum Sample Statistics	102
Table 14-5: Palladium Sample Statistics	102
Table 14-6: Gold Sample Statistics	102
Table 14-7: Silver Sample Statistics	103
Table 14-8:Cobalt Sample Statistics	103
Table 14-9: Sulfur Sample Statistics	103
Table 14-10: Spearman Rank Correlation Matrix	104
Table 14-11: Summary of Capped Values for Each Metal	113
Table 14-12: Copper Capped Composite Descriptive Statistics	114
Table 14-13: Nickel Capped Composite Descriptive Statistics	114
Table 14-14: Platinum Capped Composite Descriptive Statistics	114
Table 14-15: Palladium Capped Composite Descriptive Statistics	114
Table 14-16: Gold Capped Composite Descriptive Statistics	115
Table 14-17: Silver Capped Composite Descriptive Statistics	115
Table 14-18: Cobalt Capped Composite Descriptive Statistics	115
Table 14-19: Sulfur Capped Composite Descriptive Statistics	115
Table 14-20: Unit Variogram Parameters	116
Table 14-21: Units 20 and 3, 5, 6, and 7 Variogram Parameters	117
Table 14-22: Magenta Zone Variogram Parameters	118
Table 14-23: Search Volume Parameters for all Domains	119
Table 14-24: Mineral Resource Classification Criteria	120
Table 14-25: Copper Model Statistics	120
Table 14-26: Nickel Model Statistics	120
Table 14-27: Platinum Model Statistics	121
Table 14-28: Palladium Model Statistics	121

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Table 14-29: Gold Model Statistics	121
Table 14-30: Silver Model Statistics	121
Table 14-31: Cobalt Model Statistics	122
Table 14-32: Sulfur Model Statistics	122
Table 14-33: Resource Metal Prices and Estimated Recoveries	127
Table 14-34: Estimated Process Operating Costs	127
Table 14-35: Mineral Resource Statement for the NorthMet Project Inclusive of Mineral Reserves, Hard Rock Consulting, LLC, January 1, 2018	128
Table 15-1: Mineral Reserve NSR Cutoff	130
Table 15-2: Mineral Reserve Metal Prices	130
Table 15-3: Plant Recovery to Concentrates of Reserve Blocks	131
Table 15-4: Summary of Concentrate Treatment Terms	131
Table 15-5: Mineral Reserve Statement – December 2017	132
Table 16-1: Waste Rock Classification	135
Table 16-2: Pit Design Criteria	138
Table 16-3: Yearly Mine Production Schedule	140
Table 16-4: Yearly Mill Feed Schedule	141
Table 16-5: Mine Schedule Parameters	147
Table 16-6: Major Mine Equipment Mechanical Availability, Utilization and Fleet Size	148
Table 16-7: Drill Productivity	148
Table 16-8: Loading Equipment Productivity	149
Table 16-9: Truck Fleet Requirements	150
Table 16-10: Mine Operations and Maintenance Personnel	153
Table 16-11: Mine Operations and Maintenance Salary Personnel	154
Table 17-1: List of Major Equipment in the Hydrometallurgical Plant	178
Table 17-2: Materials Consumed by the Hydrometallurgical Plant Process	182
Table 17-3: Hydrometallurgical Plant Metal Recoveries	183
Table 20-1: Permits Under Application	193
Table 21-1: Phase I Direct Costs	199
Table 21-2: Summary of Mine Capital Cost ($USx1000)	201
Table 21-3: Mine Capital Cost by Year	202
Table 21-4: Phase II Direct Costs (Hydrometallurgical Plant)	207
Table 21-5: Direct and Indirect Costs (Phase I & II)	209
Table 21-6: Mine Operating Costs by Process	210

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Table 21-7: Mine Operating Costs Per Ton Moved ($000) by Cost Centers	212
Table 21-8: Phase I Operating Cost Estimate Summary	213
Table 21-9: Labor Schedule and Rates	215
Table 21-10: Summary of Electric Power Costs	217
Table 21-11: Process Equipment Power Draw Summary	218
Table 21-12: HVAC and Dust Collection Electric Power Summary	219
Table 21-13: Summary of Natural Gas Costs (Heating)	219
Table 21-14: HVAC Natural Gas Demand	219
Table 21-15: Process Plant Reagent and Consumable Consumption and Costs	220
Table 21-16: Maintenance, Parts and Supplies Factors	221
Table 21-17: Breakdown of Laboratory Assay Costs	222
Table 21-18: Phase II Operating Cost Estimate Summary	222
Table 22-1: LOM Operating Cost Highlights – Phase I and Phase I & II Combined	225
Table 22-2: 32,000 STPD Base Case (Phase I) Price and Operating Assumptions and Key Production Numbers	226
Table 22-3: Base Case (Phase I & II) Price and Operating Assumptions and Key Production Numbers	226
Table 22-4: Phase I Economic Projections on a Range of Metal Price Assumptions	227
Table 22-5: Phase I & II Economic Projections on a Range of Metal Price Assumptions	227
Table 22-6: Initial and Expansion Capital Summary	228
Table 22-7: Phase I and Phase I & II Operating Cost Summary	228
Table 22-8: Sustaining Capital Summary	229
Table 22-9: NorthMet Financial Model – 32,000 STPD with Hydrometallurgical Plant (Phase I and Phase II Combined	230
Table 24-1: A Comparison of the Mill Feed Tonnages between 59,000 and 118,000 STPD Throughputs	240
Table 24-2: Comparison of Cost Inputs to NSR Cutoff Grade for Various Throughputs (STPD)	241
Table 24-3: LOM Operating Highlights for 59,000 STPD & 118,000 STPD	244
Table 24-4: 59,000 STPD Economic Highlights	245
Table 24-5: 118,000 STPD Economic Highlights	246
Table 24-6: Metal Price Sensitivity Analysis for 59,000 STPD Phase I	247
Table 24-7: Metal Price Sensitivity Analysis for 59,000 STPD Phase I and II	247
Table 24-8: Metal Price Sensitivity Analysis for 118,000 STPD Phase I	248
Table 24-9: Metal Price Sensitivity Analysis for 118,000 STPD Phase I and II	248
Table 25-1: Project Risks Identified for the Feasibility Study	252
Table 25-2: Project Opportunities Identified for the Feasibility Study	255

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1 EXECUTIVE SUMMARY

Poly Met Mining, Inc. (PolyMet US), a Minnesota company and a wholly owned subsidiary of PolyMet Mining Corp. (PolyMet), contracted M3 Engineering & Technology Corporation (M3) to complete an updated Technical Report (the “Study”), at a Feasibility Study level, for the NorthMet Copper and Nickel Project (the “Project” or “NorthMet”) located near Hoyt Lakes, Minnesota, US.  PolyMet US also retained Independent Mining Consultants (IMC), Senet, (Pty) Ltd. (Senet), Hard Rock Consulting, LLC (HRC) and Barr Engineering Company (Barr) to contribute to this Study.  The update is based on feasibility-study-level engineering as well as the Final Environmental Impact Statement (FEIS, Nov 2015) and recently released draft environmental permits (Jan 2018) for the development of a 32,000-short ton per day (STPD) 225 million short ton production schedule.

PolyMet US also requested that M3 investigate potential project economic valuations using scoping or preliminary economic assessment (PEA) level mine designs at higher throughputs (59,000 and 118,000 STPD). The estimates for these two scenarios are preliminary in nature and both scenarios include Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. There is no certainty that the results for these two cases will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability and there is no certainty that Mineral Resources will become Mineral Reserves.

The purpose of the additional investigations is to quantify the potential viability of identified resources at higher throughputs that are not currently permitted for development. Development of those additional resources would require additional engineering, environmental review and permitting and would require changes in infrastructure that would require significant capital investment. The economic viability of these additional resources has not been demonstrated to date. These scenarios are only being presented in Section 1.10 and Section 24 of this report and, for clarity, they have not been included with the economic analysis presented in Section 22 of this report. In no way do these scenarios demonstrate economic viability.

Based on these results, M3 recommends that additional engineering and environmental studies be performed to further refine the costs, valuations and environmental requirements of the potential production scenarios which may have the opportunity to create additional value.

1.1                   Key Results

1.1.1               Project Phases

This Study details the construction and operation of the Project in two distinct phases.  These phases are:

· Phase I: Involves development of the NorthMet 225-million-ton orebody into an operating mine producing 32,000 STPD of ore over a 20-year life and rehabilitating an existing taconite processing plant, tailings storage facility and infrastructure (also referred to as the “Erie Plant”) located approximately eight miles to the west. Phase I would produce commercial grade copper and nickel concentrates for which Glencore AG (“Glencore”) currently holds offtake agreements payable at market terms.

· Phase II: Involves construction and operation of a hydrometallurgical plant to treat nickel sulfide concentrates into upgraded nickel-cobalt hydroxide and recover additional copper and Platinum Group Metals (“PGM”).

Execution of Phase II would be at the company’s discretion. However, both Phase I and Phase II are currently being permitted, having been included in the FEIS and draft permits.

For the purposes of this Study, all monetary values are in United States Dollars ($).  All references to “ton” or “tons” in this Study refer to US short tons except as noted otherwise.  Life of Mine (LOM capital and operating costs are reported in Table 1‑3. Metal pricing used for the financial analysis is shown in Table 1‑4. Key financial metrics and production figures are shown in Table 1‑5.

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1.1.2 Key Results for Both Phases

Both Phase I and Phase II were developed as Class 3 estimates as defined by AACE International (AACEI), which corresponds to estimates performed at a feasibility level.  Key results common to both phases are as follows:

· Total proven and probable mineral reserves for the Project are estimated to be 254.669 million tons within the pit footprints evaluated in the FEIS and draft permits.  Head grades for the reserve are shown in Table 1‑1.

· Of the mineral reserve tonnage, 225 million tons (Proven and Probable) are included in the 32,000 STPD draft permit mine plan based on metal prices shown in Table 1‑4.  For reference, the mill copper equivalent is 0.586%.

· The mine plan at 32,000 STPD yields a mine life of approximately 20 years.

· Measured and indicated resources total 649.3 million tons at a copper equivalent grade of 0.496%, inclusive of mineral reserves, and using the price assumptions reported in Table 14‑33.

· Inferred resources are estimated at 508.9 million tons at a copper equivalent grade of 0.489% (See Table 1‑2).

· Refurbishing the existing Erie Plant and associated infrastructure with a modern semi-autogenous grinding (SAG) mill and flotation plant is technically viable and will produce saleable copper and nickel concentrate products for the 32,000 STPD design used in this Study. PolyMet US plans to process 11.6 million tons of ore per year, or an average of 32,000 STPD, representing approximately one third of the historic capacity of the plant.

· PolyMet US has secured offtake agreements at market terms for copper, nickel, cobalt and PGM products from Glencore.

1.1.3               Phase I Key Results at 32,000 STPD

Under this phase, PolyMet US plans to refurbish the primary crushing circuit and replace the existing rod and ball mill circuits with a new modern semi autogenous grinding (SAG) mill, a new large ball mill and a new flotation circuit.  Once upgraded, the Erie Plant will produce copper and nickel concentrates that will be transported by rail to third-party smelting facilities.  For Phase I, the 32,000 STPD case for this Study shows:

· Initial Capital Cost Estimate (CAPEX) of $945 million,

· After-tax Net Present Value at a 7% discount rate (NPV@7%) of $173.3 million, and

· Internal Rate of Return (IRR) of 9.6%.

Under Phase I, which only includes revenues based on concentrate sales, payable metals in the concentrate are estimated as 1,096 million lbs of copper, 133 million lbs of nickel, a combined 1.05 million oz of platinum, palladium and gold, 0.96 million oz of silver and 5.6 million lbs of cobalt.  Palladium is the predominant PGM product, totalling 0.836 million oz.

Total life-of-mine (LOM) copper recovered in concentrates is expected to be 91.8%, with 63.5% recovery of nickel in concentrates under this phase.

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1.1.4 Phase II Key Results at 32,000 STPD

Phase II of the Project involves constructing a Hydrometallurgical Process that includes a 1,000 STPD autoclave to solubilize the nickel concentrates to produce a nickel-cobalt hydroxide and a precious metals precipitate.  Copper precipitates from the process will be combined with the copper concentrate.  Timing of Phase II will depend on the nickel concentrate market.  For Phase II, the 32,000 STPD case for this Study shows improved economics as follows:

· Initial CAPEX of $1,204 million (inclusive of Phase I costs),

· After-tax NPV@7% of $271 million, and

· IRR of 10.3%.

Under Phase II, payable metals in enriched copper concentrates and products from the hydrometallurgical plant are 1,155 million lbs of copper, 174 million lbs of nickel, 1.56 million combined oz of platinum, palladium and gold, 0.958 million oz of silver and 6.2 million lbs of cobalt. Palladium is the predominant PGM product, totalling 1.19 million oz.

1.2 Location And Ownership

The NorthMet Deposit is situated on a private mineral lease located in St Louis County in northeastern Minnesota, US, at approximately Latitude 47° 36’ north, Longitude 91° 58’ west, 90 road miles north of Duluth and 6.5 miles south of the town of Babbitt.

The NorthMet Project comprises two elements: The NorthMet Deposit and the nearby Erie Plant. PolyMet US leases the mineral rights to the NorthMet Deposit under a perpetually renewable lease and is acquiring the Erie Plant through contracts for deed with Cliffs Erie, L.L.C. (Cliffs Erie) a subsidiary of Cleveland Cliffs (Cliffs), which will be satisfied once the State of Minnesota issues the NorthMet permits to PolyMet US and assigns certain existing operating permits held by Cliffs Erie to PolyMet US or otherwise terminates those existing Cliffs Erie permits.

1.3 Geology And Mineralization

The NorthMet Deposit is one of twelve known copper-nickel-platinum group metal deposits along the northern margin of the Duluth Complex. The Duluth Complex is a large, composite, layered, mafic intrusion that was emplaced into comagmatic flood basalts along a portion of the Mesoproterozoic Midcontinent Rift System. The NorthMet deposit is hosted by the Partridge River Intrusion (PRI), which consists of troctolitic, anorthositic and minor gabbroic rock types that have been subdivided into seven igneous stratigraphic units. The ore-bearing units are primarily found in the basal unit of the Duluth Complex, which contains disseminated sulfides and minor massive sulfides hosted in troctolitic rocks. The Duluth Complex dips shallowly to the southeast in the western end of the deposit but steepens moving to the east.

The metals of interest at NorthMet are copper, nickel, cobalt, platinum, palladium, silver and gold. Minor amounts of rhodium, osmium, iridium and ruthenium are also present though these are considered to have no economic significance. The majority of the metals are concentrated in, or associated with, four sulfide minerals: chalcopyrite, cubanite, pentlandite, and pyrrhotite.  Platinum, palladium and gold are found in bismuthides, tellurides, and alloys. In general, the metals have strong positive correlations with copper sulfide mineralization. Cobalt has a strong correlation with nickel.  At the NorthMet Deposit, Duluth Complex rocks are overlain by up to 50 feet of overburden. Average overburden depth from all drill holes is 13 feet.

1.4 Status of Exploration

The NorthMet Deposit was formally discovered during drilling exploration carried out by U.S. Steel based on an anomaly identified during airborne survey work completed in 1966. Between 1969 and 1974, U.S. Steel drilled 112 holes for a total of 113,716 feet, producing 9,475 assay intervals, which are included in the Project database. U.S. Steel also collected three bulk surface samples for metallurgical testing from two discrete locations within the NorthMet Project area. In total, eight major exploration programs carried out at NorthMet (U.S. Steel, NERCO, and PolyMet US) have produced 436 boreholes, providing over 300,000 feet of stratigraphic control and extensive assay results.

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All exploration data have been collected in a drill-hole database used for geologic modeling, resource estimation, and mine planning. PolyMet US has verified and validated all drill-hole collar locations, down-hole surveys, lithologies, geotechnical properties, and assay data, organized all related records, and established procedures for ongoing database maintenance.

1.5 Mineral Reserve Statement

Proven and Probable Mineral Reserves of 255million tons are reported within the final pit design used for the mine production schedule and shown in Table 1‑1. All inferred material was classified as waste and scheduled to the appropriate waste stockpile.  The final mineral reserves are reported using a $7.98NSR cut-off inside the pit design using the diluted grades. Both the mineral resource and mineral reserve estimates take into consideration metallurgical recoveries, concentrate grades, transportation costs, smelter treatment charges and royalties in determining NSR values.  Table 1‑1 also shows the mineral reserves by classification category and grade. The Qualified Person responsible for the Mineral Reserve estimate is Herb Welhener, Vice President of IMC.

Table 1‑1: Mineral Reserve Statement – January 2018

Class	Tonnage (x 1,000)		Grades (Diluted)
			Copper		Nickel		Platinum		Palladium		Gold		Cobalt		Silver		NSR		Cu-Eq
			(%)		(%)		(ppb)		(ppb)		(ppb)		(ppm)		(ppm)		$/ton		(%)
Proven		121,849		0.308		0.087		82		282		41		74.81		1.11		19.87		0.612
Probable		132,820		0.281		0.081		78		256		37		74.06		1.02		18.02		0.559
Total		254,669		0.294		0.084		80		268		39		74.42		1.06		18.90		0.584

Notes:
(1)	Mineral reserve tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbers may not add due to rounding
(2)	All reserves are stated above a $7.98 NSR cutoff and bound within the final pit design.
(3)	Tonnage and grade estimates are in Imperial units
(4)	Total Tonnage within the pit is 628,499 ktons; average waste: ore ratio = 1.47
(5)	Copper Equivalent (CuEq) values are based on the metal prices in Table 15‑2 and total mill recoveries in Table 15‑3 and diluted mill feed.
(6)	Copper Equivalent (CuEq) = ((Cu head grade x recovery x Cu Price) + (Ni head grade x recovery x Ni Price) + (Pt head grade x recovery x Pt Price) + (Pd head grade x recovery x Pd Price) + (Au head grade x recovery x Au Price) + (Co head grade x recovery x Co Price) (Ag head grade x recovery x Ag Price)) / (Cu recovery x Cu Price)
(7)	NSR values include post property concentrate transportation, smelting and refining costs and payable metal calculations.

1.6 Mineral Resource Estimate

Zachary J. Black, RM-SME, of Hard Rock Consulting, LLC (HRC) is a Qualified Person as defined by NI 43-101 for mineral resource estimation and classification. HRC estimated the mineral resource for the NorthMet Project from drill-hole data constrained by geologic boundaries using an Ordinary Kriging (“OK”) algorithm.

The NorthMet Deposit was divided into eight units for geological modeling: the Biwabik Iron Formation including banded iron formation, sedimentary marine rocks of the Virginia Formation that overlie the Biwabik Formation, and five distinct units within the Duluth Complex and overburden.

The Magenta Zone, a smaller mineralized zone that cuts through Units 3 through 7 but resides primarily within Units 5 and 6, was modeled from select intercepts provided by PolyMet US.

Grades that were estimated include copper, nickel, cobalt, platinum, palladium, gold, silver and total sulfur.

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HRC created a rotated three-dimensional (3D) block model in Datamine Studio 3® mining software. The block resource model was estimated using the lithologic boundaries of the Duluth Complex as the basis for an estimation domain. Units 1, 3, 5, 6, 7, the Magenta Zone, and Virginia Formation were all estimated using only samples that resided inside of the defined boundaries. Grades were estimated from 10-foot (ft) down-hole composites using Ordinary Kriging. Composites were coded according to their domain. Each metal was estimated using variogram parameters established by AGP Mining Consultants Inc. (AGP) in 2013, which were re-evaluated by HRC and deemed acceptable for use in the current mineral resource estimation.

The mineral resources reported herein are classified as Measured, Indicated and Inferred in accordance with standards defined by the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) “CIM Definition Standards – For Mineral Resources and Mineral Reserves,” prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council on May 10, 2014. Each individual mineral resource classification reflects an associated relative confidence of the grade estimates.

The mineral resources estimated for the NorthMet Project includes 649.3 million tons of Measured and Indicated resources and 508.9 million tons Inferred resources. The resource has been limited to the material that resides above the optimized pit shell. All mineralization below the optimized pit shell has been excluded from any resource classification and is not considered to be part of the mineral resource.

The mineral resource estimate for the NorthMet Project is summarized in Table 1‑2. This mineral resource estimate includes all drill data obtained as of January 31, 2016 and has been independently verified by HRC. The Measured and Indicated mineral resources are inclusive of the mineral reserves. Inferred mineral resources are, by definition, always additional to mineral reserves. Encouraging results have prompted recommendations for additional exploration drilling to better define the Inferred mineral resources (see Note 1 in Table 1‑2).

Table 1‑2: Summary Mineral Resource Statement for the NorthMet Project Inclusive of Mineral Reserves

Class	Tonnage (Mt)		Grades (UnDiluted)
			Copper		Nickel		Platinum		Palladium		Gold		Cobalt		Silver		NSR		Cu-EQ
			(%)		(%)		(ppb)		(ppb)		(ppb)		(ppm)		(ppm)		$/ton		(%)
Measured		237.2		0.270		0.080		69		241		35		72		0.97		19.67		0.541
Indicated		412.2		0.230		0.070		63		210		32		70		0.87		16.95		0.470
M&I		649.3		0.245		0.074		65		221		33		71		0.91		17.94		0.496
Inferred		508.9		0.240		0.070		72		234		37		66		0.93		17.66		0.489

Source: Hard Rock Consulting, LLC, January 2018

Notes:

(1) Mineral resources are not mineral reserves and do not have demonstrated economic viability.

(2) All resources are stated above a $7.35 NSR cut-off.  Cut-off is based on estimated processing and G&A costs. Metal Prices and metallurgical recoveries used for the development of cut-off grade are presented in Table 14‑33.

(3) Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbers may not add due to rounding.

(4) Cu-Eq (copper equivalent grade) is based on the mill recovery to concentrates and metal prices (Table 14‑33).

(5) Copper Equivalent (Cu Eq) = ((Cu head grade x recovery x Cu Price)) + (Ni head grade x recovery x Ni Price) + (Pt head grade x recovery x Pt Price) + (Pd head grade x recovery x Pd Price) +(Au head grade x recovery x Au Price) + (Co head grade x recovery x Co Price) + (Ag head grade x recovery x Ag Price)) / (Cu recovery x Cu Price).

1.7 Mining And Processing

The NorthMet Deposit will be mined from three pits: The East Pit, the Central Pit, and the West Pit. After mining in each pit is completed, waste from the West Pit will be backfilled into the East and Central Pits, along with waste rock from the temporary waste rock stockpiles.

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Run of Mine (ROM) ore will be loaded onto rail cars at the Mine Site and transported eight miles to the Erie Plant by private railroad.

The Erie Plant processed Taconite from 1957 to 2001, processing up to 100,000 tons per day.

PolyMet US plans to refurbish the plant and reuse the existing primary crusher, and replace the downstream mill circuit with a new 40’ diameter x 22.5’ Effective Grinding Length (EGL) SAG mill and one new 24’ diameter x 37’ ball mill.

Primary ground ore will be processed through a rougher flotation circuit to produce a bulk copper and nickel concentrate. The bulk concentrate will be reground and separated in cleaner flotation. The rougher tailing will be sent to the pyrrhotite flotation circuit so that PGM-rich iron sulfide can be captured as a pyrrhotite nickel concentrate.

Tailing from the flotation circuit will be disposed of in the existing tailings basin, which is partially filled with taconite tailings exclusively, but has more than sufficient capacity for the planned operations.  The waste stream from the Hydrometallurgical Process Plant will be permanently stored in the Hydromet Residue Facility (HRF).

1.8 Environmental

Minnesota has stringent environmental standards and environmental review and permitting processes. The NorthMet environmental review process involved the Minnesota Department of Natural Resources (MDNR), the United States Army Corps of Engineers (USACE), and the United States Forest Service (USFS) as "Co-Lead Agencies." The United States Environmental Protection Agency (EPA) and tribal authorities served as cooperating agencies and the Minnesota Pollution Control Agency (MPCA) took part in the process as a permitting agency.

The most significant area of attention is water quality – the NorthMet Project is in the headwaters of the St Louis River, which flows into Lake Superior and is therefore governed by Great Lakes standards. It is important to note that NorthMet is south of the Laurentian Divide and in a separate watershed from the Boundary Waters Canoe Area Wilderness and Voyagers National Park located to the northeast.

Mineral and property tenure is secure. Permitting risks for the Project were reduced with the completion of a Final Environmental Impact Statement (FEIS) (Nov 2015) and Record of Decision (ROD) from the State of Minnesota (March 2016) indicating that the Project, as reviewed, can meet federal and state environmental standards.  The State of Minnesota has also issued all major state environmental permits in draft form for public comment. See Section 20 for a listing of required permits.

The NorthMet Project is located within an established mining district of existing open pit iron ore mines that have been mined over the last 100 years. The Peter Mitchell pit of the Northshore operations of Cleveland Cliffs lies immediately north of the NorthMet Deposit. Major impacts from the Project are limited to tailings storage in a permitted Flotation Tailings Basin (FTB), hydromet residue facility, and waste rock stockpiles and mine pits in low-lying areas.

1.9 Economics

Phase I of the NorthMet Project involves development of the 225-million-ton orebody into an operating mine producing 32,000 tons per day of ore and rehabilitating an existing taconite processing plant, tailings storage facility and infrastructure located approximately eight miles to the west. Phase I would produce commercial grade copper and nickel concentrates for which Glencore currently holds offtake agreements payable at market terms. Phase II of the Project involves construction and operation of hydrometallurgical plant to process nickel sulfide concentrates into upgraded nickel-cobalt hydroxide and recover additional copper and PGMs.   An estimate of Project capital expenditure and annual operating costs over the life of the mine for Phase I and the combined Phase I and Phase II are summarized in Table 1‑3.

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Table 1‑3: Capital Expenditure & Operating Costs – Phase I and Phase I & II

Cost Category	UOM	Phase I		Phase I & II
Capital Costs
Initial Project Capital	$M		945		1,204
LOM Sustaining Capital	$M		221		221
Operating Costs		LOM
Mining & Delivery to Plant	$/t processed		4.02		4.02
Processing	$/t processed		6.55		8.66
G&A	$/t processed		0.48		0.48
Total	$/t processed		11.05		13.16

To evaluate the economic potential of the capital investment, Phase I was structured to independently assess the overall economics both with and without Phase II (hydrometallurgical plant).  The company compiled, with the aid of its financial partners, a commodity price forecast based on consensus estimates from an extensive list of financial and industry analysts.  These prices are the basis for the financial analysis and are summarized in Table 1‑4.

Table 1‑4: Price Assumptions in the Financial Analysis

	Units	LOM
Copper	US$/lb	3.22
Nickel	US$/lb	7.95
Cobalt	US$/lb	20.68
Platinum	US$/oz	1,128
Palladium	US$/oz	973
Gold	US$/oz	1,308
Silver	US$/oz	18.92

The economic summary and financial analysis reflects processing 225 million tons of ore over a twenty-year mine life, at an average processing rate of 32,000 STPD.  Key financial results for Phase I and combined Phase I and II are presented in Table 1‑5.

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Table 1‑5: Financial Summary – 32,000 STPD

		Phase I				Phase I & II
	Units	First 5 Yrs 1		LOM		LOM 2
Life of Mine	Yrs				20		20
Material Mined	Mt		197		574		574
Ore Mined	Mt		58		225		225
Waste: Ore Ratio			2.4		1.6		1.6
Ore Grade
Copper	%		0.343		0.295		0.295
Nickel	%		0.092		0.085		0.085
Cobalt	ppm		76		75		75
Palladium	ppm		0.327		0.269		0.269
Platinum	ppm		0.099		0.079		0.079
Gold	ppm		0.048		0.039		0.039
Annual Payable Metal Produced
Copper	mlb		66.7		54.8		57.8
Nickel	mlb		7.9		6.6		8.7
Cobalt	mlb		0.33		0.28		0.31
Palladium	koz		57.6		41.8		59.4
Platinum	koz		12.4		8.5		14.3
Gold	koz		3.4		2.2		4.3
Copper Equivalent3	mlb		112.4		90.6		106.4
Cash Costs: by-product	$/lb Cu		0.67		1.06		0.59
Cash Costs: Cu equivalent	$/lb CuEq		1.71		1.91		1.79
Development Capital	$M		945		945		1,204
Sustaining Capital	$M		99		221		221
Annual Revenue	$M		362		292		343
Annual EBITDA	$M		170		118		152
NPV7 (After Taxes)	$M				173		271
IRR (After Taxes)	%				9.6		10.3
Payback (after taxes, from first production)	Years				7.3		7.5

   ¹ Represents first five years at full concentrator production.
² Phase II production is projected to commence in Year 3 of operations.
³ Cu Eq recovered payable metal, is based on prices shown in Table 1‑4, mill recovery assumptions shown in Table 15‑3 and Hydromet Phase II recoveries shown in Table 13‑14.

Financial returns for the Project are highly sensitive to changes in metal prices.  A +/-10% change in prices results in a corresponding $265 million change in NPV@7% for Phase I.  Inclusive of Phase II, the NPV@7% sensitivity is estimated to be +/-$300 million at an accuracy level of ±10%.

1.10 Potential Expansion Opportunities – Basis of 59,000 STPD and 118,000 STPD Scenarios

Metals prices for the financial analysis of both the 59,000 STPD and 118,000 STPD scenarios are based on prices shown in Table 1‑4.  The 59,000 STPD and 118,000 STPD throughput values represent Class 5 estimates as defined by AACE International, corresponding to an Order of Magnitude, Scoping or Preliminary Economic Assessment.  As such, further engineering, environmental studies and permitting would be required to prove the economic viability of these potential scenarios and to improve the economic uncertainties associated with these estimates. Further delineation drilling to move inferred resources into measured and indicated resources is also required in the 59,000 and 118,000 STPD cases.  Overall, the expansion scenarios require significant capital investment.

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The results of this exercise were as follows:

· 59,000 STPD throughput

o 293 million tons of mineralized material grading at 0.576% Cu-Eq, and

o 14-year mine life.

· 118,000 STPD throughput

o 730 million tons of mineralized material grading at 0.533% Cu-Eq, and

o 18-year mine life.

See Section 24 of this report for further evaluation of these cases.  Again, note that the estimates for these two scenarios are preliminary in nature and include Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. There can be no certainty that the results for these two cases will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability and there is no certainty that Mineral Resources will become Mineral Reserves.

1.11 Conclusions And Recommendations

M3 offers the following recommendations:

· M3 recommends that PolyMet proceed with final design, construction and operation of the 32,000 STPD design that is discussed in this Technical Report, and

· Review and update the scope of the Project design to reflect any changes resulting from the environmental review and permitting process.

Recommendations for further work are presented below:

· Based on the initial results of the additional scoping level and PEA level estimates in Section 24 of this study M3 recommends that additional engineering and environmental studies be performed at a pre-feasibility study level to further refine the costs, valuations and environmental requirements for the potential 59,000 STPD and 118,000 STPD production scenarios.  The estimated costs of these studies are expected to be $500,000. An estimated $2.5 million is required to move currently classified inferred material into measured and indicated categories.

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2 INTRODUCTION

2.1 Purpose

This report has been prepared specifically for PolyMet by the Qualified Persons (QPs) listed in Table 2‑1 to provide ‘Expert Study’ on the NorthMet Project. The findings and conclusions are based on information available at the time of preparation and data supplied by other consultants as indicated.

This report has been prepared in accordance with the guidelines provided in Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101) dated 24 June 2011 and updated on 10 May 2014.  This Technical Report has been prepared to the level of a Feasibility Study. The effective date of this report is March 26, 2018.

2.2 Sources of Information

Table 2‑1 shows the list of Qualified Persons with their associated responsibilities.

Table 2‑1: List of Qualified Persons

Name of Qualified Person	Certification	Company	Last Site Visit	Section Responsibilities
Zachary J. Black	SME-RM	Hard Rock Consulting	16 May 2017	Sections 1.3, 1.4, 1.6, 7, 8, 9, 10, 11, 12, 14, 23, 25.2.2, 25.2.3, 25.2.4, 25.2.5 and 25.2.7.
Jennifer J. Brown	SME-RM	Hard Rock Consulting	N/A	Section 6.
Nicholas Dempers	Pr. Eng., SAIMM	Senet	1 March 2018	Section 13.1-13.5, 17.1, 17.2, 18.7, 18.9, 21, 21.2.2, 21.2.3, 24.2, 25.2.6 and 25.2.10.
Thomas L. Drielick	P.E.	M3 Engineering	N/A	Sections 1.9, 19, 21, 21.1, 21.2.4, 21.2.5, 22, 25.2.12, 25.2.14, and 25.2.15.
Art S. Ibrado	P.E.	M3 Engineering	N/A	Sections 13.6, 17.3, 17.4, 17.5.1, 17.6 and 25.2.6.
Erin L. Patterson	P.E.	M3 Engineering	11 October 2017	Sections 1.1, 1.2, 1.9, 1.10, 1.11, 2, 3, 4.1- 4.5, 4.7- 4.10, 5, 18.1-18.5, 18.8, 18.10, 24, 24.2 25.1- 25.2.1, 25.2.11, 25.3- 25.5, 26 and 27.
Thomas J. Radue	P.E.	Barr Engineering Co.	11 October 2017	Section 1.7, 1.8, 4.6, 16.3.3, 17.2.4, 17.5.2, 18.6, 20.1-20.6, 20.7, 21, 21.1, 25.2.13.
Jeff S. Ubl	P.E.	Barr Engineering Co.	N/A	18.7
Herbert E. Welhener	SME -RM	Independent Mining Consultants	11 December 2000	Sections 1.5, 15, 16, 21, 21.2.1, 24.2, 25.2.8, and 25.2.9.

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2.3 Terms of Reference

Table 2‑2: Units, Terms and Abbreviation

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2.4 Units of Measure

This report uses U.S. Customary Units expressed in short tons (ton, t, 2,000 lbs), feet, and gallons consistent with U.S. Standards – unless stated otherwise. The monetary units are expressed in United States Dollars.

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3 RELIANCE ON OTHER EXPERTS

M3 relied upon contributions from a range of technical and engineering consultants as well as PolyMet. Data used in this report has been verified where possible and this report is based upon information believed to be accurate at the time of completion.  M3 is not aware of any reason why the information provided by these contributors cannot be relied upon.

Owner’s environmental and permitting costs were supplied by PolyMet. In addition, PolyMet provided all Owner’s costs in the capital cost estimate. Owners Costs are defined in section 21.1.3.3

An independent verification of land title and tenure was not performed.  M3 has not verified the legality of any underlying agreement(s) that may exist concerning the licenses or other agreement(s) between third parties.  Likewise, PolyMet has provided data for land ownership, and claim ownership.  All mineral and surface title work on the project and land exchange is managed by the law firm Hanft Fride, a Professional Association, out of Duluth, Minnesota, USA.

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4 PROPERTY DESCRIPTION AND LOCATION

4.1 Project Location

The NorthMet Project comprises two key elements: the NorthMet Deposit (or Mine Site) and the Erie Plant. The NorthMet Deposit is situated on mineral leases located in St. Louis County in northeastern Minnesota at Latitude 47° 36’ north, Longitude 91° 58’ west, about 70 miles north of the City of Duluth and 6.5 miles south of the town of Babbitt, as shown in Figure 4‑1. The Erie Plant is approximately eight miles west of the NorthMet Deposit.

The NorthMet Deposit site totals approximately 4,300 acres and the Erie Plant site, including the existing tailings basin, covers approximately 12,400 acres.

The NorthMet Project is located immediately south of the eastern end of the historic Mesabi Iron Range and is in proximity to a number of existing iron ore mines including the Peter Mitchell open pit mine located approximately two miles to the north of the NorthMet Deposit. NorthMet is one of several known mineral deposits that have been identified within the 30-mile length of the Duluth Complex, a well-known geological formation containing copper, nickel, cobalt, platinum group metals, silver, gold and titanium.

The NorthMet Deposit is connected to the Erie Plant by a transportation and utility corridor that is comprised of an existing private railroad that will primarily be used to transport ore, a segment of the existing private Dunka Road that will be upgraded to provide vehicle access, and new water pipelines and electrical power network for the NorthMet Mine Site.

Figure 4‑1: Property Layout Map

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4.2 Project Ownership

PolyMet Mining Corporation (PolyMet) owns 100% of Poly Met Mining, Inc. (PolyMet US), a Minnesota corporation.  PolyMet US controls 100% of the NorthMet Project.  As PolyMet is the owner of PolyMet US, for the sake of simplicity this Study will for the most part refer to both entities as PolyMet, except when specific differentiation is required for legal clarity.  The mineral rights covering 4,282 acres or 6.5 square miles at the NorthMet orebody are held through two mineral leases:

· The U.S. Steel Lease dated January 4, 1989, subsequently amended and assigned, covers 4,162 acres originally leased from U.S. Steel Corporation (U.S. Steel), which subsequently sold the underlying mineral rights to RGGS Land & Minerals Ltd., L.P. (RGGS). PolyMet has extended the lease indefinitely by making $150,000 annual lease payments on each successive anniversary date. The lease payments are advance royalty payments and will be deducted from future production royalties payable to RGGS, which range from 3% to 5% based on the net smelter return, subject to minimum payments of $150,000 per annum.

· On December 1, 2008, PolyMet entered into an agreement with LMC Minerals ("LMC") whereby PolyMet leases 120 acres that are encircled by the RGGS property. The initial term of the renewable lease is 20 years with minimum annual lease payments of $3,000 on each successive anniversary date until the earlier of NorthMet commencing commercial production or for the first four years, after which the minimum annual lease payment increases to $30,000. The initial term may be extended for up to four additional five-year periods on the same terms. The lease payments are advance royalty payments and will be deducted from future production royalties payable to LMC, which range from 3% to 5% based on the net smelter return, subject to a minimum payment of $30,000 per annum.

The surface rights are held by the USFS and are currently subject to a land exchange initiative with PolyMet– see Section 4.4.

PolyMet US holds various rights of ownership and use, and other property rights that currently give it control of 100% of the Erie Plant, which covers approximately 12,400 acres, or 19.4 square miles, through contracts for deed with Cliffs Erie, L.L.C. (Cliffs Erie). Further details on the arrangements with Cliffs Erie can be found in Section 4.6.

4.3 Mineral Tenure

In the 1940s, copper and nickel were discovered near Ely, Minnesota, following which, in the 1960s, U.S. Steel drilled what is now the NorthMet Deposit. U.S. Steel investigated the NorthMet Deposit as a high-grade, underground copper-nickel resource, but considered it to be uneconomic based on its inability to produce separate, clean nickel and copper concentrates with the metallurgical processes available at that time. In addition, prior to the development of the automobile-catalyst market in the 1970s, there was little market for platinum group metals (PGM) and there was no economic and reliable method to assay for low grades of these metals.

In 1987, the Minnesota Natural Resources Research Institute (NRRI) published data suggesting the possibility of a large resource of PGMs in the base of the Duluth Complex.

PolyMet, as Fleck Resources, acquired a 20-year perpetually renewable mineral rights lease to the NorthMet Deposit in 1989 from U.S. Steel. The lease is subject to yearly lease payments before production and then to a sliding scale Net Smelter Return (NSR) royalty ranging from 3% to 5%, with lease payments made before production considered as advance royalties and credited to the production royalty. PolyMet leases an additional 120 acres of mineral rights from LMC.

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Mineral and surface rights have been severed, with the USFS owning the surface rights within most of the lease area. U.S. Steel retained the mineral rights and certain rights to explore and mine on the site under the original documents that ceded surface title to the USFS.

4.4 Surface Rights

Surface rights of the NorthMet Deposit are held by the USFS. The United States acquired the surface rights from U.S. Steel in 1938 under provisions of the Weeks Act of 1922. U.S. Steel retained certain mining rights, which PolyMet secured under the U.S. Steel Lease, along with the mineral rights.

PolyMet and the USFS have proposed a land exchange to consolidate their respective land ownerships.

In this land exchange, the USFS will acquire, 6,690 acres of private land in four separate tracts currently held by PolyMet, to become part of the Superior National Forest and managed under the laws relating to the National Forest System. Already located within the Superior National Forest boundaries, these lands will have multiple uses including recreation, research and conservation. The USFS will convey 6,650 acres of federally-owned surface land to PolyMet, which includes the surface rights overlying and surrounding the NorthMet Deposit. These lands are located near an area heavily used for mining and mine infrastructure, are consistent with regional land uses, and will generate economic benefits to the region through employment and tax revenues.

Following the Final NorthMet Environmental Impact Statement (FEIS), the Superior National Forest of USFS issued a Final Record of Decision (ROD) to proceed with the administrative land exchange in January 2017. The ROD stated, among other things, that the proposed exchange will be beneficial to the USFS and is in the public’s interest.  On November 28, 2017, H.R. 3115, the Superior National Forest Land Exchange Act of 2017, passed by voice vote in the House of Representatives.  If enacted into law, H.R. 3115 will legislatively accomplish the same land exchange approved in the January 2017 USFS ROD. The administrative land exchange process is ongoing as of the date of this report.

4.5 Royalties And Encumbrances

The NorthMet Deposit mineral rights carry variable royalties of 3% to 5% based on the NSR per ton of ore mined. For a NMV of under $30 per ton, the royalty is 3%, for $30-35 per ton it is 4%, and above $35 per ton it is 5%. Both the U.S. Steel Lease (RGGS) and the LMC Lease carry advance royalties which can be recouped from future royalty payments, subject to minimum payments in any year.  The US Steel leases were transferred through sale to RGGS though the underlying agreement terms remain the same.

4.6 Environmental Liabilities

Federal, state and local laws and regulations concerning environmental protection affect the PolyMet operation.  As part of the consideration for the purchase of the Erie Plant and associated infrastructure, the Company indemnified Cliffs for reclamation and remediation obligations of the acquired property.  Completion of that purchase remains subject to certain contingencies, including, among other things, issuance of final permits for the NorthMet Project under applicable environmental laws and release of Cliffs, and its subsidiary Cliffs Erie, from its obligations under existing state permits with respect to the Erie Plant and other assets acquired by PolyMet.

According to PolyMet US, the Company’s estimate of the environmental rehabilitation provision under International Financial Reporting Standards (IFRS) on October 31, 2017 was $72.772 million based on estimated cash flows required to settle this obligation in present day costs of $78.729 million, a projected inflation rate of 2.00%, a market risk-free interest rate of 2.66% and expenditures expected to occur over a period of approximately 30 years. This estimate includes but is not limited to water treatment and infrastructure closure and removals, with costs estimated by PolyMet and its consultants and construction contractors. This estimate has been reviewed and accepted by auditors for PolyMet’s financial statement.

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4.7 Permits

Prior to construction and operation of the NorthMet Project, PolyMet will require several permits from federal and state agencies – see Section 20.4.

4.8 Social License

The environmental review process is described in Section 20. The federal, state and local government permits needed for PolyMet to construct and operate the NorthMet Project are described in Section 20.4.

PolyMet has maintained an active community outreach program for many years. The focus of the program has been to provide information about the Project, its likely impact on the environment, and the socioeconomic benefits. The local communities are supportive of the Project. PolyMet continues to receive outstanding community and political support for the Project.  The local mayors, U.S. Senators, Congressmen and elected state officials continue to express public support for both the process and the Project.

The Bois Forte Band of Chippewa (Bois Forte), Grand Portage Band of Chippewa (Grand Portage), and the Fond du Lac Band of Lake Superior Chippewa (Fond du Lac) have been cooperating agencies in preparation of the FEIS. Fond du Lac has expressed the strongest opposition, primarily related to cultural heritage issues and seeking to ensure that water quality is protected.

The most active environmental groups in the area are focused on protecting the Boundary Waters Canoe Area Wilderness, which is located approximately 25 miles northeast of the NorthMet site, in a different watershed.

4.9 Significant Risk Factors

4.9.1 Permitting

Permitting is the most significant risk factor for the Project. The NorthMet Project is the first copper-nickel project in Minnesota to seek permits for construction and operation. Environmental review and permitting is, perhaps, the biggest challenge facing any mining project in the United States.

Permitting risk falls into two primary categories:

1. Permits may be denied or legally challenged, or

2. Operating requirements imposed by the permits could be so financially burdensome that the Project is unable to proceed.

While all final permits remain to be issued and are potentially subject to legal challenge, permitting risk has decreased due to completion and acceptance of the FEIS, the associated state and USFS ROD issuance, and the issuance of the draft state permits.

4.9.2 Project Financing

PolyMet will require successful project financing in order to complete development and construction of the NorthMet Project.  If PolyMet cannot raise the money necessary to fund the Project, development will be suspended.  Sources of such external financing may include future equity and debt offerings. This risk is partially mitigated through the company’s ongoing relationship with Glencore.

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Phase II of the Project includes construction of a hydrometallurgical facility after Phase I operations have commenced.  Financing risk associated with this phase of the Project is mitigated by Phase I financials.

4.9.3 Commodity Prices

If the price of metals in the PolyMet ore body decrease below a specified level it may no longer be profitable to develop the NorthMet Project.  Once developed, if metal prices are, for a substantial period, below foreseeable costs of production PolyMet operations could be negatively affected.

See Section 25.4 of this Study for a discussion of additional risks.

4.10 Comments on Section 4

Mineral and property tenure is secure, pending completion of the land exchange with the USFS and the contracts for deed with Cliffs Erie as referenced in Sections 4.4 and 4.6, respectively.  Acquisition of surface rights is the subject of both the USFS Final ROD, issued in January 2017, and the administrative land exchange or HR 3115, which the US House of Representatives approved on November 28, 2017. Completion of the acquisition of the Erie Plant from Cliffs Erie is subject to, among other requirements, finalization of the draft permits issued by the State of Minnesota for the NorthMet Project and release of Cliffs Erie from certain existing state permits under processes anticipated and described in draft NorthMet permits issued by MDNR and MPCA.  Permitting risks for the Project have been reduced with the completion of the FEIS (Nov 2015) and ROD from the State of Minnesota (March 2016) indicating that the Project, as reviewed, can meet federal and state environmental standards.

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The project site is located just south of the eastern end of the historically significant Mesabi Iron Range, a world-class mining district that has the capacity to produce, annually, approximately 44 million gross tons of iron ore pellets and concentrate from iron bearing ore named taconite.  There are currently six iron ore mines on the Mesabi Iron Range, see Table 5‑1.

Table 5‑1: Current Iron Ore Mines on the Mesabi Iron Range

Operation Name	Ownership	Annual Capacity	Location	Status as of June 1, 2016
Minntac	100% United States Steel	16 million net tons	Mt. Iron, Minnesota	Operating
Keetac	100% United States Steel	6 million net tons	Keewatin, Minnesota	Idle
ArcelorMittal Minorca Mines	100% ArcelorMittal	2.9 million tons	Virginia, Minnesota	Operating
United Taconite	100% Cleveland Cliffs	5.4 million gross tons	The mine is located near Eveleth, Minnesota, the plant is located approximately 10 miles away in Forbes, Minnesota	Idle
Northshore Mining	100% Cleaveland Cliffs	6 million gross tons of pellets and concentrate	The mine is located near Babbitt, Minnesota, the plant is located approximately 47 miles away in Silver Bay, Minnesota	Operating
Hibtac	62.3% ArcelorMittal 23% Cleaveland Cliffs 14.7% United States Steel Note:  This operation is managed by Cleaveland Cliffs	8 million gross tons	Hibbing, Minnesota	Operating

The Northshore Mining Peter Mitchell Pit is located approximately two miles north of the NorthMet Deposit.

5.1 Accessibility

Access to the NorthMet Project is by a combination of good quality asphalt and gravel roads via the Erie Plant site. The nearest center of population is the town of Hoyt Lakes, which has a population of about 2,500 people. There are a number of similarly sized communities in the vicinity, all of which are well serviced, provide ready accommodations, and have been, or still are, directly associated with the region’s extensive taconite mining industry. The road network in the area is well developed, though not heavily trafficked, and there is an extensive railroad network which serves the taconite mining industry across the entire Range. There is access to ocean shipping via the ports at Taconite Harbor and Duluth/Superior (on the western end of Lake Superior) and the St. Lawrence Seaway.

5.2 Climate

Climate is continental and characterized by wide temperature variations and significant precipitation. The temperature in the town of Babbitt, about 6.5 miles north of the NorthMet Deposit, averages four degrees Fahrenheit (°F) in January and 66°F in July.  During short periods in summer, temperatures may reach as high as 90°F with high humidity. Average annual precipitation is about 28 inches with about 30% of this falling mostly as snow between November and April. Annual snowfall is typically about 60 inches with 24 to 36 inches on the ground at any one time. The local taconite mines operate year-round and it is rare for snow or inclement weather to cause production disruption.

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5.3 Local Resources and Infrastructure

The area has been economically dependent on the mining industry for many years and while there is an abundance of skilled labor and local mining expertise, the closure in 2001 of the LTVSMC open pit mines and taconite processing facility has had a significant negative impact on the local economy and population growth. There are, however, several other operating mines in other parts of the Iron Range. Because of this, the mining support industries and industrial infrastructure remains well developed and of a high standard.

The Erie Plant site is connected to the electrical power supply grid and a main HV electrical power line (138 kV) runs parallel to the road and railroad that traverse the southern part of the mining lease area.  PolyMet has a long-term power contract with Minnesota Power.

There are plentiful local sources of fresh water, and electrical power and water is available nearby.  Previous operations at the site processed 100,000 STPD with adequate water supply, which is more than three times the plan for PolyMet.

5.4 Physiography

The Iron Range forms an extensive and prominent regional topographic feature. The Project site is located on the southern flank of the eastern Range where the surrounding countryside is characterized as being gently undulating. Elevation at the Project site is about 1,600 ft asl (1,000 ft above Lake Superior). Much of the region is poorly drained and the predominant vegetation comprises wetlands and boreal forest. Forestry is a major local industry and the Project site and much of the surrounding area has been repeatedly logged. Relief across the site is approximately 100 ft.

5.5 Sufficiency of Surface Rights

Tenure of surface rights is described in some detail in Section 4.4. The surface rights over the ore body are currently owned by the USFS.  PolyMet has proposed a land exchange with the USFS which has been evaluated in the FEIS.  The USFS issued a ROD in January 2017 indicating that the proposed exchange is in the public interest and meets the objectives of the Superior National Forest Plan.

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6 HISTORY

6.1 Ownership

U.S. Steel held mineral and surface rights over much of the region, including the NorthMet lease, until the 1930s when, for political and land management reasons, surface title was ceded to the USFS.  In negotiating the deeds that separated the titles, U.S. Steel retained the mineral rights and the rights to explore and mine any mineral or group of minerals.

U.S. Steel first drilled what is now known as the modern day NorthMet deposit in the 1960s during exploration for a high-grade, underground copper-nickel resource. In 1989, Fleck Resources Ltd. of British Columbia, Canada, acquired a 20-year perpetually renewable mineral rights lease to the NorthMet deposit from U.S. Steel. Fleck Resources developed joint ventures with NERCO Inc. in 1991, and with Argosy Mining Corp. in 1995, in order to advance exploration of the NorthMet deposit.

In June 1998, Fleck Resources changed its name to PolyMet Mining Corp. U.S. Steel sold much of its real estate and mineral rights in the region in 2004, including the NorthMet deposit, to privately held RGGS of Houston Texas. PolyMet’s U.S. Steel lease was transferred to RGGS at that time without any change in conditions.  With the exception of a hiatus between 2001 and 2003, PolyMet has continuously carried out exploration and evaluation of the NorthMet deposit since 1989, and currently holds 100% interest in the NorthMet Project.

6.2 Exploration and Sampling

The NorthMet deposit was formally discovered in 1969 during exploration carried out by U.S. Steel. Between 1969 and 1974, U.S. Steel drilled 112 holes for a total of 113,716 ft, producing 9,475 assay intervals which are included in the modern-day Project database. Assay data from U.S. Steel core samples was not necessarily collected at the time of the original drilling. U.S. Steel also collected three bulk surface samples for metallurgical testing from two discrete locations within the NorthMet Project area. The drill-hole and data accumulated during exploration by U.S. Steel provides important stratigraphic information, and is used to help define the edges of the NorthMet geologic model. U.S. Steel’s exploration efforts, including drilling and sampling procedures and general results, are described in greater detail in Sections 9 and 10 of this report.

6.3 Historical Mineral Resource and Reserve Estimates

A number of historic mineral resource estimates were completed (U.S. Steel, Fleck Resources, NERCO) prior to PolyMet’s acquisition of the NorthMet Project.  These resource estimates predate current NI 43-101 reporting standards and the associated resource models, electronic or otherwise, are not available for verification.  Although it is reasonable to presume that they were completed using industry best practices at the time, these mineral resources are not classified using current CIM definition standards, are not reported according to modern reporting codes, are not considered reliable, and therefore are not presented here.

6.4 Historical Production

There is no historical production data to report for the NorthMet Project.

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7 GEOLOGICAL SETTING AND MINERALIZATION

The information presented in this report section is largely excerpted and/or modified from the Geology and Mineral Potential of the Duluth Complex and Related Rocks of the Northeastern Minnesota (Miller et al., 2002).

7.1 Regional Geology

The NorthMet Deposit is situated on the western edge of the Duluth Complex in northeastern Minnesota (shown in Figure 7‑1). The Duluth complex is a series of distinct intrusions of mafic to felsic tholeiitic magmas that intermittently intruded at the base of a comagmatic volcanic edifice during the formation of the Midcontinental rift system between 1108 and 1098 Ma. The intrusives of the Duluth Complex represent a relatively continuous mass that extends in an arcuate fashion from Duluth to the northeastern border between Minnesota and Canada near the town of Grand Portage.  Footwall rocks are predominantly comprised of Paleoproterozoic and Archean rocks, the hanging wall rocks are made up of mafic volcanic rocks and hypabyssal intrusions, and internally scattered bodies of strongly granoblastic mafic volcanic and sedimentary hornfels can be found.

The Duluth Complex has been subdivided into four general rock series based on age, dominant lithology, internal structure, and structural position within the complex.

7.1.1 Felsic Series

Massive granophyric granite and smaller amounts of intermediate rock that occur as a semi continuous mass of intrusions strung along the eastern and central roof zone of the complex emplaced during early stage magmatism (~1108 Ma).

7.1.2 Early Gabbro Series

Layered sequences of dominantly gabbroic cumulates that occur along the northeastern contact of the Duluth Complex that were also emplaced during early stage magmatism (~1108 Ma).

7.1.3 Anorthositic Series

A structurally complex suite of foliated, but rarely layered, plagioclase-rich gabbroic cumulates that was emplaced throughout the complex during main stage magmatism (~1099 Ma).

7.1.4 Layered Series

A suite of stratiform troctolitic to ferrogabbroic cumulates that comprises at least 11 variably differentiated mafic layered intrusions and occurs mostly along the base of the Duluth Complex.  These intrusions were emplaced during main stage magmatism, but generally after the anorthositic series (~1099 Ma).

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Figure 7‑1: Regional Geology

Intrusive rocks of the layered series typically reside along the western edge of the Duluth Complex, and host the 11-known copper-nickel deposits (some contain platinum group elements) including the NorthMet Deposit (Figure 7‑2). The layered series is comprised of 11 discrete mafic layered intrusions spread throughout the Duluth Complex. The 11 known layered series intrusives are known as; Layered series at Duluth, Boulder Lake intrusion, Western Margin intrusion, Partridge River intrusion, South Kawishiwi intrusion, Lake One troctolite, Tuscarora intrusion, Wilder Lake intrusion, Bald Eagle intrusion, Greenwood Lake intrusion, Osier Lake intrusion.

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Figure 7‑2: Copper-Nickel Deposits in the Duluth Complex (after Severson)

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7.2 Local And Property Geology

The NorthMet Deposit is situated within the Partridge River Intrusion (“PRI”). The PRI has been mapped, drilled, and studied in detail because of its importance as a host for copper-nickel (“Cu-Ni”) and iron-titanium (“Fe-Ti”) deposits. The PRI consists of varied troctolitic and (minor) gabbroic rock types that are exposed in an arcuate shape that extends from the Water Hen (Fe-Ti) deposit in the south to the Babbitt (Cu-Ni) deposit in the North (Figure 7‑2). Miller and Ripley (1996) estimated the PRI to be nearly 8,000 feet thick. The PRI is bound on the west by the Paleoproterozoic Virginia Formation (slate and graywacke), and to a lessor extent, the Biwabik Iron Formation (“BIF”). The upper portion of the PRI forms a complex contact an assemblage of anorthositic, gabbroic, and hornfelsic rocks. This assemblage is also found as large inclusions within the interior of the PRI (Severson and Miller, 1999). The inclusions are thought to represent earlier roof zone screens that were overplated by later emplacement of Partridge River intrusion magmas.

The bottom 3,000 feet of the PRI is well defined from the abundance of exploration drill core.  There are over 1,100 exploration drill holes in this part of the Complex, and nearly 1,000,000 feet of core has been logged or re-logged in the past fifteen years by a small group of company and university research geologists (see Patelke, 2003). This marginal zone, consisting of varied troctolitic and gabbroic rock types, is subdivided into seven stratigraphic units (Severson and Hauck, 1990, 1997; Geerts, 1991; Severson, 1991, 1994) that can be correlated over a strike length of 15 miles. These igneous units generally exhibit shallow dips (10º to 25°) to the southeast.  The stratigraphy shown in Figure 7‑3 is based on the relogging of drill core.

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Figure 7‑3: NorthMet Stratigraphic Column (after Geerts, 1994)

7.2.1 Local Lithology

The following paragraphs describe the principal rock types (and associated map units) within the Project area.

Igneous rock types in the PRI are classified at NorthMet by visually estimating the modal percentages of plagioclase, olivine, and pyroxene. Due to subtle changes in the percentages of these minerals, a variation in the defined rock types within the rock units may be present from interval to interval or hole to hole. This is especially true for Unit 1.

Unit definitions are based on: overall texture of a rock type package; mineralogy; sulfide content; and context with respect to bounding surfaces (i.e., ultramafic horizons, oxide-rich horizons). Unit definitions are not always immediately clear in logging, but usually clarified when drill holes are plotted on cross-sections. In other words, to correctly identify a particular igneous stratigraphic unit, the context of the units directly above and below must also be considered. Figure 7‑4 shows a plan view of the NorthMet geological contacts within the mining lease area.

Based on drill hole logging, the generalized rock type distribution at NorthMet is about 83% troctolitic, 6% anorthositic, 4% ultramafic, 4% sedimentary inclusions, 2% noritic and gabbroic rocks, and the rest as pegmatites, breccia, basalt inclusions and others.

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Figure 7‑4: NorthMet Property Bedrock Geology

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7.2.2 Unit Definitions and Descriptions

The units of the NorthMet deposit are described below starting at the top of the PRI.

7.2.2.1 Unit 7

Unit 7 is the uppermost unit intersected in drill holes at the NorthMet Deposit. It consists predominantly of homogeneous, coarse-grained, anorthositic troctolite and troctolitic anorthosite. The unit is characterized by a continuous basal ultramafic sub-unit that averages 20 ft thick. The ultramafic consists of fine to medium-grained melatroctolite to peridotite and minor dunite. The average thickness of Unit 7 is unknown due to the truncation by erosion on the surface exposure.

7.2.2.2 Unit 6

Similar to Unit 7, Unit 6 is composed of homogeneous, fine to coarse-grained, troctolitic anorthosite and troctolite.  It averages 400 ft thick and has a continuous basal ultramafic sub-unit that averages 15 ft thick.  Sulfide mineralization is generally minimal, although many drill-holes in the southwestern portion of the NorthMet deposit contain significant copper sulfides and associated elevated platinum group elements (Geerts 1991, 1994). Sulfides within Unit 6 generally occur as disseminated chalcopyrite/cubanite with minimal pyrrhotite.

7.2.2.3 Unit 5

Unit 5 exhibits an average thickness of 250 ft and is composed primarily of homogeneous, equigranular-textured, coarse-grained anorthositic troctolite.  Anorthositic troctolite is the predominant rock type, but can locally grade into troctolite and augite troctolite towards the base of the unit.  The lower contact of Unit 5 is gradational and lacks any ultramafic sub-unit; therefore, the contact with Unit 4 is a somewhat arbitrary pick.  Due to the ambiguity of the contact, reported thicknesses of both units vary dramatically.  The combined thickness of Units 4 and 5, however, is fairly consistent across the extent of the deposit.

7.2.2.4 Unit 4

Unit 4 is somewhat more mafic than Unit 5, and is characterized by homogeneous, coarse-grained, ophitic augite troctolite with some anorthosite troctolitic.  Unit 4 averages about 250 ft thick.  At its base, Unit 4 may contain a thin (<6 in), discontinuous, local ultramafic layer or oxide-rich zone.  The lower contact with Unit 3 is generally sharp.  With the exception of the Magenta Zone (described further in Section 7.2), sulfides only occur in Unit 4 in trace amounts of finely disseminated grains of chalcopyrite and pyrrhotite.

7.2.2.5 Unit 3

Unit 3 is the primary marker bed used to determine stratigraphic position in drill core.  Unit 3 is composed of fine to medium-grained, poikilitic and/or ophitic, troctolitic anorthosite to anorthositic troctolite.  Characteristic poikilitic olivine gives the rock an overall mottled appearance.  On average, Unit 3 is 300 ft thick.  The lower contact of Unit 3 can be disrupted, with multiple “false starts” into relatively homogeneous rocks typical of Unit 2, only to return to the mottled appearance characteristic of Unit 3 with depth.  This roughly alternating sequence, or transitional zone, is commonly encountered in the southwestern portion of the NorthMet deposit, and can span for many tens of feet of core before the transition into Unit 2 can be confidently identified.  The transitional zone between Units 2 and 3 suggests that Unit 3 is disturbed and intruded by Unit 2 near the base of Unit 3.  As with Units 4 and 5, the independent thicknesses of Units 2 and 3 tend to be highly variable, whereas their combined depth is relatively consistent throughout the deposit (though not as consistent as Units 4 and 5).

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Unit 3 can contain both footwall meta-sedimentary (Virginia Formation) and hanging wall basalt inclusions, which are interpreted as an indication of earliest emplacement within the intrusive sequence of the NorthMet deposit.  This interpretation is exemplified by the fact that few sedimentary inclusions are found above Unit 3, and few basalt inclusions are found below it, which can be attributed to the intrusion of Unit 3 between the two rock types.

7.2.2.6 Unit 2

Unit 2 is characterized by homogeneous, medium to coarse-grained troctolite and pyroxene troctolite with a consistent basal ultramafic sub-unit.  The continuity of the basal ultramafic sub-unit, in addition to the relatively uniform grain size and homogeneity of the troctolite, cause this unit to be distinguishable from Units 1 and 3.  Unit 2 has an average thickness of 100 ft.  The ultramafic sub-unit at the base of Unit 2 is the lowermost continuous basal ultramafic horizon at the NorthMet deposit, averaging 25 ft thick, and is composed of melatroctolite to peridotite and minor dunite.

The boundaries of Unit 2 and its arrangement within the sequence of intrusion are ambiguous; it can be interpreted as the lower part of Unit 3, the upper part of Unit 1, or a separate unit all together.  Based on the continuity of the ultramafic sub-unit, it seems to be a lower, more mafic, counterpart to Unit 3.  The general lack of footwall inclusions in Unit 2 counter the contention that Unit 2 is older than Unit 1, and instead indicate an intrusive sequence of 3, 1 then 2.  Though Unit 2 has historically been described as barren, mineralization which is grossly continuous at the top of Unit 1, has been encountered in Unit 2 in the western portion of the NorthMet deposit.

7.2.2.7 Unit 1

Of the seven igneous rock units represented within the NorthMet Deposit, Unit 1 is the only unit that contains significant, deposit-wide sulfide mineralization.  Sulfides occur primarily as disseminated interstitial grains between a dominant silicate framework and are chalcopyrite > pyrrhotite > cubanite > pentlandite.  Unit 1 is also the most complex unit, with internal ultramafic sub-units, increasing and decreasing quantities of mineralization, complex textural relations and varying grain sizes, and abundant metasedimentary inclusions.  It averages 450 ft thick, but is locally 1,000 ft thick and is characterized lithologically by fine to coarse-grained heterogeneous rock ranging from anorthositic troctolite (more abundant in the upper half of Unit 1) to augite troctolite with lesser amounts of gabbro-norite and norite (becoming increasingly more abundant towards the basal contact) and numerous metasedimentary inclusions.  By far, the dominant rock type in Unit 1 is medium-grained ophitic augite troctolite, though with wildly variable texture.  Two internal ultramafic sub-units with an average thickness of 10 ft are encountered in drill holes in the southwest portion of the deposit.

7.2.2.8 Footwall: Animikie Group and Archean Rocks

The footwall rocks of the NorthMet deposit consist of Paleoproterozoic (meta) sedimentary rocks of the Animikie Group.  These rocks are represented by the following three formations, from youngest to oldest: the Virginia Formation; the Biwabik Iron Formation; and the Pokegama Quartzite.  They are generally underlain by Archean granite of the Giants Range Batholith, but there are Archean basalts and metasediments mapped in an outcrop near the Project area.  The Virginia Formation is the only member of the Animikie Group in contact with the Duluth Complex in the NorthMet Project area.

The Virginia Formation was metamorphosed during emplacement of the Duluth Complex.  Non-metamorphosed Virginia Formation (as found to the north of the site) consists of a thinly-bedded sequence of argillite and greywacke, with lesser amounts of siltstone, carbonaceous-sulfidic argillite/mudstone, cherty-limey layers, and possibly some tuffaceous material.  However, in proximity to the Duluth Complex, the grade of metamorphism (and associated local deformation) progressively increases, and several metamorphic varieties and textures are superimposed on the original sedimentary package at an angle to the original stratigraphy.  At least four distinctive metamorphosed Virginia Formation varieties are present at NorthMet and are informally referred to as the cordieritic metasediments; disrupted unit; recrystallized unit; and graphitic argillite (often with pyrrhotite laminae).  These sub-units are fully described in Severson et al., 2000.

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7.2.2.9 Inclusions in the Duluth Complex

Two broad populations of inclusions occur at NorthMet: hanging wall basalts (Keweenawan) and footwall meta-sedimentary rocks.  The basalts are fine-grained, generally gabbroic, with no apparent relation to any mineralization.  Footwall inclusions may carry substantial sulfide (pyrrhotite) and often appear to contribute to the local sulfur content.  Footwall inclusions are all Virginia Formation; no iron-formation, Pokegama Quartzite, or older granitic rock has been recognized as an inclusion at NorthMet.

7.3 Local Structure

Footwall faults are inferred from bedding dips in the underlying sedimentary rocks, considering the possibility that Keweenawan syn-rift normal faults may affect these underlying units and show less movement, or indeed no effect on the igneous units.  Nonetheless, without faults, the footwall or igneous unit dips do not reconcile perfectly with the overall slope of the footwall.  There are some apparent offsets in the igneous units, but definitive and continuous fault zones have not been identified.  So far, no apparent local relation between the inferred location of faults and mineralization has been delineated.

Outcrop mapping (Severson and Zanko, 1996) shows apparent unit relations that require faults for perfect reconciliation.  However, as with information derived from drill core, neither igneous stratigraphic unit recognition, nor outcrop density, is sufficiently definitive to establish exact fault locations without other evidence.

There is a wealth of regional (and some local) geophysical data available, though the resolution of core logging and field mapping is probably better than that of the geophysics, hence while the geophysical data is interesting, it has not yet been useful at delineating the structural geology of the site nor proved to be a guide to mineralization.

7.4 Mineralization

The metals of interest at NorthMet are copper, nickel, cobalt, platinum, palladium, silver, and gold.  Minor amounts of rhodium and ruthenium are present though these are considered to have no economic significance.  In general, except for cobalt and gold, the metals are positively correlated with copper mineralization.  Cobalt is well correlated with nickel. Most of the metals are concentrated in, or associated with, four sulfide minerals: chalcopyrite, cubanite, pentlandite, and pyrrhotite, with platinum, palladium and gold also found as elements and in bismuthides, tellurides, and alloys.

Mineralization occurs in four broadly defined horizons or zones throughout the NorthMet property.  Three of these horizons are within basal Unit 1, though they likely will not be discriminated in mining.  The upper horizon locally extends upward into the base of Unit 2.  The thickness of each of the three Unit 1 enriched horizons varies from 5 ft to more than 200 ft.  Unit 1 mineralization is found throughout the base of the NorthMet deposit.  A less extensive mineralized zone (the copper-rich, sulfur-poor Magenta Zone, Figure 7‑5) is found in Units 4, 5 and 6 in the western part of the NorthMet deposit.

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Figure 7‑5: NorthMet “Magenta Zone” in Cross Section

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8 DEPOSIT TYPES

Information in this section is largely excerpted and/or modified from the Occurrence Model for Magmatic Sulfide-Rich Nickel-Copper-(Platinum Group Element) Deposits Related to Mafic and Ultramafic Dike-Sill Complexes (Schulz et al., 2014).

The NorthMet deposit is considered a magmatic Copper - Nickel ± platinum group element (PGE) deposit. These are a broad group of deposits containing nickel, copper, and PGEs occurring as sulfide concentrations associated with a variety of mafic and ultramafic magmatic rocks (Zientek, 2012; Eckstrand and Hulbert 2007). Magmatic Cu-Ni sulfide deposits with or without PGEs account for approximately 60 percent of the world’s nickel production. Magmatic Ni-Cu±PGE sulfide deposits are spatially and genetically related to bodies of mafic and/or ultramafic rocks. The sulfide deposits form when the mantle-derived magmas become sulfide-saturated and segregate immiscible sulfide liquid, commonly following interaction with continental crustal rocks.

Deposits of magmatic Ni-Cu sulfides occur with mafic and/or ultramafic bodies in a wide array of geologic settings.  The deposits range in age from Archean to Tertiary, but the largest number of deposits are Archean and Paleoproterozoic, as with the NorthMet deposit. Although deposits occur on most continents, ore deposits (deposits of sufficient size and grade to be economic to mine) are relatively rare; major deposits are present in Russia, China, Australia, Canada, and southern Africa. Ni-Cu sulfide ore deposits can occur as single or multiple sulfide lenses within mafic and/or ultramafic bodies with clusters of such deposits comprising a district. Typically, deposits contain grades of between 0.5 and 3.0 percent Ni and between 0.2 and 2.0 percent Cu. Tonnages of individual deposits range from a few tens of thousands to tens of millions of tons (Mt). Two giant Ni-Cu districts, with ≥10 Mt Ni, dominate world Ni sulfide resources and production. These are the Sudbury district, Ontario, Canada, where sulfide ore deposits are at the lower margins of a meteorite impact-generated igneous complex and contain 19.8 Mt Ni; and the Noril’sk-Talnakh district, Siberia, Russia, where the deposits are in subvolcanic mafic intrusions related to flood basalts and contain 23.1 Mt Ni. In the United States, the Duluth Complex in Minnesota, comprised of a group of mafic intrusions related to the Midcontinent Rift system, represents a major Ni resource of 8 Mt Ni. The Duluth Complex deposits generally exhibit lower grades of nickel and copper (0.2 percent Ni, 0.66 percent Cu).

The sulfides in magmatic Ni-Cu deposits generally constitute a small volume of the host rock(s) and tend to be concentrated in the lower parts of the mafic and/or ultramafic bodies, often in physical depressions or areas marking changes in the geometry of the footwall topography. In most deposits, the sulfide mineralization can be divided into disseminated, matrix, and massive sulfide, depending on a combination of the sulfide content of the rock and the silicate texture. The major Ni-Cu sulfide mineralogy typically consists of an intergrowth of pyrrhotite, pentlandite, and chalcopyrite. Cobalt, PGE, and gold (Au) are extracted from most magmatic Ni-Cu ores as by-products, and such elements can have a significant impact on the economics of the deposits, such as the Noril’sk-Talnakh deposits, which produces much of the world’s palladium. In addition, deposits may contain between 1 and 15 percent magnetite associated with the sulfides.

The NorthMet deposit is a large-tonnage, disseminated accumulation of sulfide in mafic rocks, with rare massive sulfides. Copper to nickel ratios generally range from 3:1 to 4:1. Primary mineralization is probably magmatic, though the possibility of structurally controlled re-mobilization of the mineralization (especially PGE) has not been excluded.  The sulfur source is both local and magmatic (Theriault et al., 2011).  Extensive detailed logging has shown no definitive relation between specific rock type and the quantity or grade quality of sulfide mineralization in the Unit 1 mineralized zone or in other units, though local noritic to gabbronoritic rocks (related to footwall assimilation) tend to be of poorer PGE grade and higher in sulfur.

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9 EXPLORATION

The information presented in this section is largely excerpted and/or modified from the Updated NI 43-101 Technical Report on the NorthMet deposit prepared by AGP Mining Consultants, Inc. (AGP, 2013).

U.S. Steel’s interest in the NorthMet deposit (also known as the Dunka deposit) was triggered by an anomaly identified during airborne survey work conducted in 1966. U.S. Steel mapped and ground surveyed the property the following year, and initiated drilling exploration in 1968. Drilling has been the primary method of exploration at the Project, however, 240 geophysical soundings, numerous test pits, and down-hole geophysical testing have been completed to better understand the depth to bedrock and the lithologic contacts.

9.1 Geophysical Sounding

Ninety-Eight Vertical Electrical Soundings (VES) were completed at the NorthMet project in 2006. The VES geophysical method was selected to determine the depth to bedrock and to characterize the overburden material. The method is based on the estimation of the electrical conductivity or resistivity of the material. The estimation is performed based on the measurement of voltage of electrical field induced by the grounded electrodes (current electrodes).

In general, the measured profiles consisted of three differing resistive layers. A high resistivity layer primarily consisting of the surficial frozen layer. Below the surficial layer a resistivity low represents the till. The resistivities varied widely in this layer, depending on the material properties of the till. The bottom layer is bedrock, either Duluth complex or Virginia formation. In nearly all of the measurements the bottom layer has a higher resistivity than the till above, with the exception of a few locations above the Virginia formation. Portions of the Virginia formation can be enriched in pyrite, pyrrhotite or graphite, making it more conductive than the till above.

9.2 U.S. Steel Bulk Sampling

U.S. Steel took at least three bulk samples from the Dunka Road deposit, labeled in their documentation as Bulk No. 1, Bulk No. 2, and Bulk No. 3. U.S Steel also took a few small trench samples and processed some drill core composites from the site. These are recorded in the sample receiving books at Coleraine Minerals Research Laboratory (Patelke and Severson, 2006).

Bulk No. 1 was collected in 1980 in NW¼ Section 10, T59N, R13W, near the location of U.S. Steel drill-hole DDH 26058. Historic records indicate that a 70 to 85-ton sample was collected from this site, which returned a reported bulk head grade of 0.39% Cu, 0.14% Ni, and 0.50% S, but there is no associated documentation regarding site selection or metallurgical testing (Patelke and Severson, 2006).

Bulk No. 2 was the first of two samples collected from the Project in 1971. This sample consisted of 300 tons of material from a pit located directly north of the up-dip projection of DDH 26105. According to U.S. Steel documents, the sample did not intersect the grades expected, and the low grade was attributed to contamination by barren footwall rock.

Bulk No. 3 was collected at the south edge (stratigraphically higher) Bulk No. 2 pit to move up-section from the footwall rock contamination encountered in Bulk No. 2. A 20-ton sample was collected, which returned a bulk head grade of 0.58% Cu, 0.22% Ni, and 0.98% S (Patelke and Severson, 2006).

Associated U.S. Steel documents only reference DDH 26105 prior to collecting the bulk samples. It is not known whether any blast holes or studies were completed in preparation or during the collection of the samples.

The pilot plant tests on three bulk samples of copper-nickel sulfides from the Project resulted in recoveries of 83 to 89 percent of the total copper and 72 to 85 percent of the sulfide nickel in a cleaned bulk sulfide concentrate containing 20 percent copper and 4.5 percent nickel. Mineral liberation required grinding to 75 percent passing a minus 200 mesh. Crushing and grinding consumed about 23 net kWh per ton.

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Differential flotation of the bulk sulfide concentrate was unsuccessfully attempted to make separate copper and nickel concentrates.  It was determined that a selective flotation scheme maintained good selectivity and high metal recovery in bench scale tests. This was accomplished in two steps; 1) floating the copper sulfides, and 2) and floating the previously depressed nickel sulfides. However, this method was problematic in the pilot plant as it was difficult to control the critical parameters, notably pH of the pulp, during the various stages of flotation.

The historic documents indicate that U.S. Steel was confident that the extraction process would be economically feasible. However, the additional test work required for detailed costing was never completed (Patelke and Severson, 2006).

9.3 Down-Hole Geophysical Testing

In 1970 and 1971, a geophysical company and the United States Geological Survey (USGS) respectively, initiated two separate attempts to determine if down-hole geophysical methods could be used to:

· Determine the distribution of sulfide-mineralized material around a single drill hole,

· Determine the continuity of sulfide-mineralized zones between drill holes,

· Determine if lithologic rock type differences could be detected by geophysical methods,

· Provide background information for surface exploration techniques, and/or

· Test new and modified logging instruments.

Hewitt Enterprises of Draper, UT, conducted two types of down-hole surveys on five U.S. Steel drill-holes in 1970. An in-hole electrical survey was used to make resistivity and induced polarization (IP) measurements at regular intervals in three drill holes, and five drill holes were logged using the potential drop method to measure self potential (SP), IP and electric potential (ΔV). Results from both surveys were judged to be ineffectual in responding to sulfide content or lithology (Severson and Heine, 2007).

In 1971, the USGS made in-hole logging measurements of seven U.S. Steel drill holes.  Due to several unfortunate incidents with the probe becoming stuck in some of the holes, only a minimum of information was obtained. According to Severson and Heine (2007), preliminary results suggested that:

· Continuous in-hole logging is more advantageous than the spot measurements that were made in 1970,

· IP measurements could not be made because of the extremely high resistivity of 20,000 to 30,000-ohm meters and relatively short delay time (12 milliseconds) after cessation of current pulse,

· The gamma ray logs delineated the graphitic hornfels with an associated higher background radioactivity,

· Resistivity and magnetic susceptibility measurements could be used collectively to distinguish between pyrrhotite-rich zones and magnetite-rich zones,

· It appeared that resistivity could not be used to correlate sulfide zone in one hole to a nearby hole, and

· In-hole logging does not appear to show any meaningful results for determining the continuity of mineralized zones between drill holes, and thus, does not appear to be a substitute for drilling.

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10 DRILLING

10.1 Introduction

Exploration drilling was carried out by U.S. Steel between 1969 and 1974. In total, eight drilling programs have been conducted at NorthMet (U.S. Steel, NERCO, and PolyMet) resulting in 439 drill holes, representing over 300,000 feet of stratigraphic control and analytical results.

In addition to the data provided by the drilling exploration programs, stratigraphic data is available from another seventy exploration holes drilled in the area for nearby projects, hydrogeological studies, or water supply wells. All exploration data is maintained by PolyMet in a drill-hole database used for resource evaluation, reserve calculation, and mine planning. PolyMet has verified and validated all drilling locations, down-hole surveys, lithology, rock property, and assay data, organized all related records, and established procedures for ongoing database maintenance.

Prior to PolyMet’s involvement in the Project, 116 core holes were drilled in the main Project area by U.S. Steel and NERCO. Table 10‑1 lists the drill-holes by series, type and company drilled specifically for the NorthMet Project.  Figure 10‑1 shows the drill-hole locations.

Table 10‑1: NorthMet Project Drill Hole Summary

Date	Hole Identification Range	Exploration Company	Drill-hole Type	No. Of Holes Drilled	Reported/Actual Feet
1969 -1974	26010 - 26143	U.S. Steel	Core	112	133,716
1991	26086A, 26101A	NERCO	Core	2(4)	842
1998-2000	"98-," "99-," "00-"	PolyMet	RC	52	24,650
1999-2000	"99-," "00-"	PolyMet	Core	32	22,156
2000	"99-"	PolyMet	Core	3	2,697
2005	"05-"	PolyMet	Core	109	77,167
2007	"07-"	PolyMet	Core	61	24,530
2010	"10-"	PolyMet	Core	66	20,132

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Figure 10‑1: Drill-hole Collar Location by Campaign

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10.2 Historic Drilling

10.2.1 U.S. Steel Drilling, 1969 – 1974

From 1969 to 1974, U.S. Steel contracted Longyear to drill 112 diamond core holes across the property.  Early exploration drilling programs were designed to test geophysical targets. The US Steel drilling was designed to intersect a potential geophysical conductor. The first hole drilled on the NorthMet deposit intersected 4.8% Cu in a 3-ft intersection of massive sulfide, 115 ft from the surface.  Follow up drill results were less impressive, however drilling resulted in the delineation of a broad zone of low-grade copper-nickel sulfide mineralization.  Further drilling indicated that the original geophysical target was graphitic argillite in the footwall, rather than mineralization in the Duluth Complex.

The majority of the core was BQ size.  All but 14 of the holes drilled by US Steel were vertical. Hole depths ranged from 162 ft to 2,647 ft, averaging 1,193 ft.  Five holes were drilled to depths exceeding 2,500 ft.

10.2.2 NERCO Drilling 1991

NERCO conducted a minor drilling campaign in 1991, which consisted of four holes at two sites.  At each site, a BQ sized core hole (1.43 inches) was drilled and the entire drill hole was sampled.  A PQ (3.3 inch) hole twinned each of these holes, and the associated core was sent in its entirety for metallurgical work on the assumption that the assays on the smaller diameter core would represent the larger diameter core.  Both sets of holes twinned existing U.S. Steel holes (Pancoast, 1991). A total of 165 assays from the smaller diameter cores were processed at ACME.

10.3 PolyMet Drilling

PolyMet completed 290 drill holes between 1998 and 2010 totaling 171,332 ft. Of the 290 holes drilled by PolyMet, 52 were drilled using reverse circulation, and 238 are diamond core holes.  Drilling exploration conducted by PolyMet is summarized in Table 10‑1, and drill hole distribution is shown on Figure 10‑1.

10.3.1 PolyMet Drilling, 1999 – 2000, Reverse Circulation Holes

From 1998 to 2000, PolyMet drilled 52 vertical reverse circulation (RC) holes to supply material for a bulk sample.  A portion of these drill-holes twinned U.S. Steel holes, and others served as in-fill over the extent of the NorthMet deposit. The RC holes averaged 474 ft, with a minimum of 65 ft and a maximum depth of 745 ft. The drilling was completed by a contractor from Duluth with extensive RC experience, and was carried out year-round.  The type of bit and extraction system used (cross-over sub or face-sampling) is not known.  Available recorded sample weights indicate a recovery of at least 85%.  Metallurgical core drilling, in approximately February and March of 2005, twinned some of these RC holes.

10.3.2 PolyMet Drilling, 1999-2000, Diamond Core Holes

The first PolyMet core drilling program was carried out during the later parts of the RC program, with three holes drilled late in 1999 and the remainder in early 2000.  There were seventeen BTW (1.65 inch) and fifteen NTW (2.2 inch) diameter holes all of which were vertical.  Three RC holes were re-entered and deepened with AQ core. Core holes averaged 692 ft in depth, with a minimum of 229 ft and a maximum depth of 1,192 ft. (not including RC holes extended with AQ core). These holes were assayed from top to bottom (with minimal exception) on 5-foot intervals.  Samples were split into half core at the PolyMet field office in Aurora, Minnesota. Core logging was completed at the PolyMet office by geologists trained to recognize the stratigraphic units and the subtleties of the mineralogy and textures described by Severson (1988).

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10.3.3 PolyMet Drilling, 2005, Diamond Core Holes

PolyMet’s 2005 drilling program had four distinct goals: collection of metallurgical sample, continued in-fill drilling for resource estimation, resource expansion, and collection of oriented core for geotechnical data.  The program included 109 holes totaling 77,165 ft, including:

· 15 one-inch diameter holes for metallurgical samples (6,974 ft) drilled by Boart-Longyear of Salt Lake City (February - March 2005).

· PQ sized holes (core diameter 3.3 inches) totaling 6,897 ft, to collect bulk sample material, and to improve the confidence in the known resource area (February - March 2005).

· 52 NTW sized holes (2.2 inches) totaling 41,403 ft for resource definition.

· 30 NQ2 sized holes (2.0 inches) totaling 21,892 ft for resource definition and geotechnical purposes.  The NTW and NQ2 size core was drilled in the spring (February-March) and fall (September-December) of 2005.

Roughly 11,650 multi-element assays were collected from the 2005 drilling program.  Another 1,790 assays were performed on previously drilled U.S. Steel and PolyMet core during, as well.  ALS-Chemex completed all the analytical test work for 2005 drilling and re-sampling program.

Of the 109 holes drilled in 2005, 93 were drilled at an angle. The angled holes were aligned on a grid oriented N34W with dips ranging from -60° to -75°.  Sixteen NQ2 sized holes were drilled and marked for oriented core at varying dips, for geotechnical assessment across the Project.  These holes targeted positions of the projected pit walls, as defined by Whittle pit shells (AMDAD mining consultants).  The targeted locations and geotechnical data are continually reviewed as the project advances and are considered to be reasonable for the current iteration of the pit design.

PolyMet analyzed close to 900 core intervals for “whole rock” oxides, 300 samples were analyzed for Rare Earth Elements (REE), and thousands of density measurements were completed.  This data is used to support resource evaluation as well as waste characterization efforts required for permitting.

Separately, about 100 samples from previously drilled and analyzed core were submitted for humidity cell testing.  These samples represented a broad cross-section of units, rock-types, metal content, and sulfur content.  In addition, these humidity cell samples were all re-assayed, analyzed for whole rock and assessed in thin-section and by micro-probe.

10.3.4 PolyMet Drilling, 2007, Diamond Core Holes

In 2007, PolyMet conducted two drilling programs, a winter program of 47 holes totaling 19,102.5 ft and a summer program of 14 holes totaling 5,437.5 ft.  The initial 16 winter holes were NTW sized, the remaining drill holes from both programs were NQ2 core.  Most of these holes were angled to north-northwest (azimuth 326°). The 2007 holes averaged 402 ft in depth, with a minimum of 148 ft and maximum of 768.5 ft.

10.3.5 PolyMet Drilling, 2010, Diamond Core Holes

In 2010, PolyMet conducted a winter drilling program with two objectives:

1. Collect detailed geostatistical data across a grid in the initial mining area, and

2. Develop a geologic and assay framework around the west margin of the deposit.

Secondary to these purposes was the gathering of approximately ten tons of potential bulk sample material.

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The grid area in the planned east pit encompassed 8,720 ft of drilling with 1,664 multi-element assays and the western drilling totaled 11,401 ft with 1,345 samples taken. Grid drilling was sampled by elevations representing bench levels. Data from this was used to establish appropriate sampling protocols during mining.

Assay results in the grid area were consistent with expectations from previous block models. In the west, Unit 1 and Magenta Zone ore grade mineralization continue well outside the planned pit boundaries with the furthest hole in this program 2,600 feet to the west of the planned pit edge.

10.4 Relevant Results And Interpretation

Very little documentation is available on drilling and sampling procedures employed by U.S. Steel and NERCO.  However, the drilling was conducted by companies experienced in exploration and production and is considered reliable.

In all cases, drilling has shown a basal mineralized zone (Unit 1) in heterogeneous troctolitic rocks with the highest values in the upper portion with grades generally diminishing to depth along drill holes.  Grade appears to increase down dip, but less information is available as the depth to the unit intersection increases.  The main ore zone is 200 to 1,000 ft thick, averaging about 450 ft.  The mineralization extends from base of the till at the north edge of the Project and continues to depths greater than 2,500 ft.  Sampling on the deepest holes is sparse, with little in-fill work done since the original U.S. Steel drilling. PolyMet collected 700 samples from the deeper U.S. Steel holes in the spring of 2006, this data is included in the exploration database.

Core recovery is reported by PolyMet to be upwards of 99% (Table 10‑2) with rare zones of poor recovery.  Rock quality designation (RQD) is also very high, averaging 85% for all units, excluding the Iron formation.  Experience in the Duluth Complex indicates that core drilling has no difficulty in producing samples that are representative of the rock mass.  Rock is fresh and competent and the types of alteration (when observed: sausserization, uralization, serpentinization and chloritization) do not affect recovery.

Values exceeding 100 may arise from errors associated with assembling broken core or from core runs that are slightly longer than the core barrel.

Table 10‑2: Summary of Core Recoveries and RQD Measurements (includes all drilling through 2010)

Unit	Recovery Count	Recovery Percentage (%)	RQDCount	RQDPercent
1	8,906	99.9	4,194	91.8
2	1,879	99.5	968	90.3
3	4,374	100	2,632	93.5
4	2,160	100	1,063	96.4
5	1,901	100	838	94.3
6	2,262	100	1,041	94.7
7	951	99.3	396	87.4
Virginia Formation	2,095	99.7	1,069	87.6
Inclusions	62	98.1	57	86.6
Biwabik Iron Formation	381	100	60	79.8
Duluth Complex Average		99.96		92.82

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11 SAMPLE PREPARATION, ANALYSES AND SECURITY

There are multiple generations of sample analyses that contribute to the overall project assay database:

· Original U.S. Steel core sampling, by U.S. Steel, 1969-1974

· Re-analysis of U.S. Steel pulps and rejects, selection by Fleck and NRRI, 1989-1991

· Analysis of previously un-sampled U.S. Steel core, sample selection by Fleck and NRRI in 1989-1991, and 1999-2001

· Analysis of 2 of the 4 NERCO drill-holes, 1991

· PolyMet RC cuttings, 1998-2000

· PolyMet core, 2000, 2005, 2007, and 2010

The laboratories utilized by U.S. Steel were not independent of the company, and no information regarding accreditation is available. All the labs that have provided analytical testing for PolyMet were or currently are fully accredited, independent, commercial labs that are not related to any of the exploration companies or any of its directors or management.

PolyMet's drill hole and assay database is administered by company geologic staff from the operational headquarters in Hoyt Lakes. PolyMet uses Excel and Gemcom GEMS to manage the geologic data. Paper logs are available at the operational headquarters.

11.1 Historic Sample Preparation, Analysis and Security

11.1.1 U.S. Steel and NERCO

There is no documentation indicating sample handling protocols at drill sites, and only limited documentation of sample handling between the drill site and assay laboratory for programs conducted by U.S. Steel and NERCO.

U.S. Steel assayed approximately 22,000 ft of the 133,716 ft drilled, on nominal 10-ft intervals. The drill programs were focused was on delineating an underground resource and sampling was restricted to zones of continuous “higher grade” mineralization.  The selected sample intervals targeted the primary zone of mineralization (Unit 1) rather than intermittent mineralized intervals or presumed waste rock.

Core was split by U.S. Steel using a manual core splitter.  Samples submitted for assay were typically half core.

Samples were shipped to Lerch Brothers of Hibbing Minnesota (Lerch) or to the State of Minnesota for preparation prior to analysis. Both laboratories used a jaw crusher to reduce the nominal sample size to minus 1/4 inch. The samples were then reduced to a 250-gram split and a Bico Type Plate grinder pulverized the remaining sample to minus 149 µm. Samples processed by Bondar Clegg were processed in the same manner but were pulverized in a ring mill to minus 106 µm.

U.S. Steel completed approximately 2,200 samples. Each sample was analyzed for copper, nickel, sulfur, and iron.  Assays were completed at one of two U.S. Steel laboratories in Minnesota, the Applied Research Laboratory (ARL) in Coleraine (now the NRRI mineral processing laboratory), or at the Minnesota Ore Operations (MOO) laboratory at the Minntac Mine in Mountain Iron, MN. It is not known what type of certification ARL or MOO may have had between 1969-1974.

The analytical methods utilized at the U.S. Steel laboratories is unknown. While standards were developed and used (as evidenced by documents in PolyMet files), it is not thought the standards were inserted into the sample stream in a blind manner. It is likely that these were used for calibration or spot checks.

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U.S. Steel was cognisant of the potential PGEs from the assaying of concentrates derived from bench scale tests but did not systematically assay for these metals on drill core. Most of the U.S. Steel samples have been replaced in the database by the results of the reanalysis programs that include PGEs. There are less than 200 sample intervals of U.S. Steel copper-nickel values that remain in the database.

Seventeen of the U.S. Steel holes were “skeletonized” after assaying, with only 1 ft retained for each 5 or 10-ft “un-mineralized” and un-sampled run.  Drilling by PolyMet adjacent to the locations of skeletonized core indicate the possibility that some mineralized intervals may have been missed and discarded in the skeletonizing process.

U.S. Steel geologists did not document any interpretation of comprehensive igneous stratigraphy during drill hole logging.  Mark Severson of the Natural Resources Research Institute (NRRI), in Duluth, Minnesota began re-logging the U.S. Steel drill holes in the late 1980s as part of a Partridge River intrusion geochemistry project.  He recognized Unit 3 as a marker horizon, which led to reliable correlations among the other units. Steve Geerts, working for the NRRI with Fleck Resources, refined the geologic model for the NorthMet Deposit considering the igneous stratigraphy.  His interpretation is still considered valid by PolyMet, and currently guides the interpretation of the NorthMet Deposit (Severson 1988, Severson and Hauck 1990, Geerts et al. 1990, Geerts 1991, 1994).

Starting in 1989 Fleck and NRRI began to reanalyze pulp rejects and unsampled intervals from the U.S. Steel drill programs. Fleck, NRRI, and PolyMet continued the reanalysis through 2006. In total 5,032 samples intervals and 229 duplicates were submitted for analysis.

The remaining available core from the U.S. Steel drill programs is stored at the Project and is available for further analysis.

11.1.2 PolyMet Sample Preparation, Analysis and Security

Employees of PolyMet (or Fleck Resources) have been either directly or indirectly involved in all sample selection since the original U.S. Steel sampling.  Sample cutting and preparation of core for shipping has been done by PolyMet employees or contract employees.  Reverse circulation sampling at the rig was done by, or in cooperation with, PolyMet employees and the drilling contractor.

The diamond drillers remove the drill core samples from the rods and place them into covered core boxes. PolyMet representatives collect the trays and transport them to the core storage facility located near the processing plant each day where the core is inventoried prior to processing. Once the geologist is ready to log the hole, the core trays are laid out on core logging tables where all logging takes place prior to sampling.

Drill core samples are placed into plastic sample bags, sealed, and placed into a cardboard box. The cardboard box is sealed shut with tape and couriered to the laboratory. Once the laboratory has accepted delivery of the samples they remain under the control of the laboratory.

The RC holes were assayed on 5-ft intervals.  Six-inch RC drill-holes produced about 135 lb to 150 lb of sample for every 5 feet of drilling.  This material was split using a riffle splitter into two samples and placed in plastic bags and stored underwater in five-gallon plastic buckets.  A 1/16th sample was taken by rotary splitter from each 5-ft interval of chip sample for assay.  The assay values were used to develop a composite pilot plant sample from bucket samples.  Actual compositing was completed after samples had been shipped to Lakefield (Patelke and Severson, 2006).  A second 1/16th sample was sent to the Minnesota Department of Natural Resources for their archive.

There are 5,216 analyses from the RC drilling in the current PolyMet database. RC sample collection involved a 1/16 sample representing each five-foot run. These were sent to Lerch for preparation, and then sent to ACME or Chemex for analysis.

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Chip samples were collected and logged at the PolyMet office and are currently retained at the PolyMet warehouse.  While the chip sample logging is less precise than logging of core samples, the major silicate and sulfide minerals are identifiable, and the location of marker horizons can be derived based on the composition of the individual samples.  The underlying metasedimentary rocks (Virginia Formation) are readily recognized in chip sample, and the base of the NorthMet Deposit is relatively easy to define.  Where rock recognition is difficult, the higher zinc content of the footwall rocks is used to help define the contact.

PolyMet geologists log all drill cores at the core storage facility located near the processing plant. The geologists record information for each drillhole (Supplemental Information, 2018) including the hole number, azimuth, total depth, coordinate datum, drilling company, hole logger, start and end of drilling dates, rock codes, and a written description of stratigraphy, alteration, texture, mineralogy, structure, grain size, ground conditions, and any notable geologic features. The rock quality designation (RQD) and recovery percentage are also recorded.

Sample intervals are determined by the geologist with respect to stratigraphy, mineralization, and sulfide content, otherwise a standard 10-ft interval is sampled. Zones of increased sulfide mineralization >2.5 ft are sampled down to 5-ft intervals. Core within Unit 1 is sampled on 5-ft intervals. Core samples are cut to ¼ or ¹/₈ of the total core with a diamond bladed saw by trained personnel following written procedures. Each sample is placed in a numbered plastic sample bag with the corresponding sample number tag and placed in a cardboard box for transport to the laboratory. All QA/QC samples are inserted into the sample stream prior to shipment.

11.1.3 Sample Preparation

Samples were prepared for analysis at Lerch, Acme, or Chemex facilities. In general, all the facilities followed a similar preparation procedure. Samples were crushed to an approximate -10 mesh, prior to being reduced to a 250-gram split for pulverization (149 to 106 µm range).  Pulps were split again to separate a sample for the following analyses:

· Base metals (Cu, Co, Mo, Ni and Zn) - Four-acid digestion with ICP-AES finish,

· Base metals (Ag, Cu, Co, Mo, Ni and Zn) – Aqua Regia digestion with ICP-AES finish,

· PGEs (Au, Pt and Pd) – 30 gm fire assay with ICP-AES finish, and

· Total Sulphur by LECO furnace.

Select core samples were crushed to -1/2 inch and placed in a poly bottle, purged with nitrogen, and capped and sealed for special metallurgical and environmental analysis

11.2 Analytical History

Information in this section is largely excerpted and/or modified from the Review of the PolyMet 2005-2006 Quality Control Program (Bloom, 2006).

11.2.1 Base Metals

PolyMet samples were analyzed using a 0.250 g Aqua Regia or four-acid digestion with an Inductively Coupled Plasma – Atomic Emission Spectroscopy (ICP-AES) finish. Detection limits for the elements analyzed by these methods are presented in Table 11‑1.

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Table 11‑1: Detection Limits of Elements

Element	Symbol	Detection Limit	Upper Limit	Units
Silver	Ag	2	10	ppm
Cobalt	Co	1	10,000	ppm
Copper	Cu	0.001	1	%
Molybdenum	Mo	1	10,000	ppm
Nickel	Ni	0.001	1	%
Zinc	Zn	2	10,000	ppm

11.2.2 Platinum Group Elements

Samples analyzed for PGEs utilized 30 g Fire Assay (FA) with an ICP-AES finish.  In this method a prepared sample (30 g) is mixed with a fluxing agent. The flux assists in melting, helps fuse the sample at a reasonable temperature and promotes separation of the gangue material from the precious metals. In addition to the flux, lead or nickel is added as a collector. The sample is then heated in a furnace where it fuses and separated from the collector material button, which contains the precious minerals. The button is digested for 2 minutes at high power by microwave in dilute nitric acid. The solution is cooled, and hydrochloric acid is added. The solution is digested for an additional 2 minutes at half power by microwave. The digested solution is then cooled, diluted to 4 ml with 2% hydrochloric acid, homogenized and then analyzed for gold, platinum and palladium by inductively coupled plasma – atomic emission spectrometry emission spectrometry. Detection limits for the elements analyzed by this method is presented in Table 11‑2.

Table 11‑2: Detection Limits

Element	Symbol	Detection Limit	Upper Limit	Units
Gold	Ag	1	10,000	ppb
Platinum	Co	1	10,000	ppb
Palladium	Cu	5	10,000	ppb

11.2.3 Total Sulfur

Total sulfur was analyzed by a LECO Furnace with Infrared Spectroscopy. In this method the sample is analyzed for total sulfur using a Leco analyzer. A stream of oxygen passes through a prepared sample (0.05 to 0.6 g) while it is heated in a furnace to approximately 1350°C. Sulfur dioxide released from the sample is measured by an infrared detection system and the total sulfur result is provided. This technique has a lower detection limit of 0.01% and an upper detection limit of 50%.

11.3 Quality Assurance/Quality Control Procedures

QA/QC samples used by PolyMet include blanks, standards and field duplicates. PolyMet inserts QA/QC samples into the sample stream at the following frequencies:

· Insertion of coarse blank every 40 samples;

· Insertion of Standard Reference Material (SRM) every 40 samples; and

· Submission of duplicate ¼ or ¹/₈ of the drill core every 40 samples.

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A stockpile of crushed Biwabik Iron Formation rock was submitted as a coarse preparation blank.  The blank is uncertified, but analysis has demonstrated that is below detection limit for the metals of interest.

PolyMet contracted CDN Resources Laboratories Ltd. (Vancouver) to prepare three SRMs for the drilling programs. The SRMs were prepared by CDN Resources Laboratories Ltd. (Vancouver) from 63 coarse reject U.S. Steel samples in 2004.  The SRM performance range was determined through a round robin analysis in 2005. The round robin results are shown in Table 11‑3.

Table 11‑3: Details of Sampling of U.S. Steel Core by PolyMet

Element	SM 4-1		SM 4-2		SM 4-3
	Average	Std. Dev	Average	Std. Dev	Average	Std. Dev
Co (ppm)	90.1	10.44	95.10	10.64	110.73	11.11
Cu (%)	0.201	0.008	0.378	0.009	0.589	0.019
Mo (ppm)	13.87	1.78	9.61	1.36	12.25	1.40
Ni (%)	0.109	0.007	0.143	0.009	0.197	0.015
Zn (ppm)	174.15	14.62	116.77	12.18	124.76	12.65
Au (ppb)	57.85	12.70	33.32	6.48	54.18	7.36
Pt (ppb)	36.54	9.50	55.76	11.15	125.52	15.55
Pd (ppb)	117.52	10.66	238.95	14.64	518.05	22.18
S (%)	1.17	0.04	0.91	0.04	1.15	0.005

Averages are based on twenty samples of each standard with 4-acid digestion ICP-AES assays completed in 2005.

PolyMet submitted ¼ or ¹/₈ of the core was submitted as a duplicate interval.  During the drilling programs, PolyMet submitted coarse blanks, core duplicates, and SRMs.

11.3.1 Blanks

Coarse blanks monitor the integrity of sample preparation and are used to detect contamination during crushing and grinding of samples.  Blank failures can also occur during laboratory analysis or as the result of a sample mix-up.  A blank analysis ≥5 times the detection limit is considered a blank failure Table 11‑1 and Table 11‑2.

PolyMet submitted 697 coarse pulp blanks to monitor sample preparation during the drilling programs.  Less than 4% of the samples blank samples submitted to reported values exceeding 5 times the detection limit for a particular element.  In all cases 10 samples either side of the blank were re-submitted, and a new blank was inserted.  Results were acceptable. Copper and nickel blank analyses are presented in graphical form in Figure 11‑1 and Figure 11‑2, respectively.

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Figure 11‑1: Copper Blank Analysis

Figure 11‑2: Nickel Blank Analysis

11.3.2 Standards

Standards are used to monitor laboratory consistency and to identify sample mix-ups. PolyMet inserted standards into the sample stream at a rate of 1:40 for the drill programs conducted between 2005 and 2010. During the drilling programs, acceptable reference standards tolerances were established at ±2 standard deviations (“stdev” or “σ”) from the mean of the standard. In total 762 (301 SM4-1, 287 SM4-2, and 174 SM4-3) standards were submitted for analysis with approximately 5.0% of the samples exceeding the established thresholds. Overall the means of each standard were in line with the reference mean. Standards exceeding the tolerances established by PolyMet were reviewed and, depending on the nature of the failures, samples may be re-run or discarded from the dataset.

HRC reviewed the standards employed by PolyMet to insure reliable assay information throughout the database.  The individual standards were plotted against ±2 and ±3 standard deviations of the expected standard mean (Figure 11‑3 and Figure 11‑4).  The two types of failures can be identified by the red and orange colored symbols on the figures.

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Figure 11‑3: Copper Results for Standard 4-1

Figure 11‑4: Nickel Results for Standard 4-1

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11.3.3 Duplicates

11.3.3.1 Core Duplicates

Duplicates are used to monitor sample batches for sample mix-ups, data variability due to laboratory error and sample homogeneity at each step of preparation.  Sample duplicates should be inserted at every sample split during sample preparation and they should not be placed in sequential order.  When original and duplicates samples are plotted in a scatterplot, perfect analytical precision will plot on x=y (45°) slope.  Core duplicates are expected to perform within ±30% of the x=y slope, coarse preparation duplicates should perform within ±20% of the x=y slope while pulp duplicates are expected to perform within ±10% of the x=y slope on a scatterplot.

PolyMet submitted ¼ and ¹/₈ core duplicates in the drilling programs prior to 2007.  A total of 236 quarter-core duplicate pairs were submitted. The Cu and Ni assays for the original and duplicate samples are compared in Figure 11‑5.

Figure 11‑5: Copper and Nickel ¼ Core Duplicate Analysis

A total of 87 one-eight-core duplicate pairs were submitted. The Cu and Ni assays for the original and duplicate samples are compared in Figure 11‑6.

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Figure 11‑6: Copper and Nickel 1/8 Core Duplicate Analysis

The core duplicate performance suggests that the sample size is adequate for copper and no bias is evident in the comparison.

11.3.3.2 Historic Pulp Re-analysis

The analysis of U.S. Steel pulps, sampling of previously un-sampled core, and two NERCO core holes was completed between 1989-1991 by Fleck Resources in cooperation with the NRRI in Duluth.  Many pulps and coarse rejects from the original U.S. Steel drilling were re-assayed for copper, nickel, PGE, and a full suite of other elements.  The NRRI selected, sampled, and re-logged the unsampled core.  This was the first large-scale testing for PGE done on the Project. Figure 11‑6, Figure 11‑7 and Figure 11‑8 compare the U.S. Steel results with the reanalysis. The copper results generally agree, but the nickel results demonstrated a bias toward the U.S. Steel assays. Most of the U.S. Steel samples have been replaced in the database by the results of the reanalysis programs that include PGEs. There are less than 200 sample intervals of U.S. Steel copper-nickel values that remain in the database.

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Figure 11‑7: Copper Pulp Duplicate Analysis

Figure 11‑8: Nickel Pulp Duplicate Analysis

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11.4 Data Entry Validation Controls

PolyMet manages the drill-hole assay data with a project specific Microsoft Access® database maintained in Gemcom Gems software and various excel spreadsheets. All information has been audited by HRC with limited errors identified. It is HRC’s opinion that PolyMet maintains a complete, well documented, and easily auditable geological and assay database.

11.5 Core Storage And Sample Security

The U.S. Steel core has been stored, either at the original U.S. Steel warehouse in Virginia, Minnesota during drilling, or more recently at the CMRL (now a part of the University of Minnesota).  Core has been secured in locked buildings within a fenced area that is locked at night where a key must be checked out.  The NERCO BQ size core is also stored at this facility.

The PolyMet core and RC reference samples were stored in a PolyMet leased warehouse in Aurora, Minnesota during drilling and pre-feasibility.  Core and samples were then moved in 2002 to a warehouse in Mountain Iron, Minnesota where they remained until 2004.  They were then moved to a warehouse at the Erie Plant site in Hoyt Lakes.  Access to this warehouse is limited to PolyMet employees.

11.6 Opinion On Adequacy

HRC concludes that the sample preparation, security and analytical procedures are correct and adequate for the purpose of this Technical Report.  The sample methods and density are appropriate, and the samples are of sufficient quality to comprise a representative, unbiased database.

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12 DATA VERIFICATION

12.1 PolyMet Data Compilation And Verification 2004

The mineral resource and reserve estimates rely in part on the following information provided to HRC by PolyMet with an effective date of December 31, 2015:

· Discussions with PolyMet personnel,

· An exploration drilling database received as .csv files,

· Modeled solids for the 3 formations present at the Project; the Biwabik Iron Formation, the Duluth Complex, and the Virginia Formation; along with modeled solids for the site overburden and Magenta domain, and

· The most recent Technical Report “Updated NI 43-101 Technical Report on the NorthMet Deposit Minnesota, USA” dated October 12, 2102 and amended January 14, 2013 and authored by AGP Mining Consultants, Inc. (Alsp, 2013).

Topography was provided as 2-ft contours derived from air photo work in 1999.

12.2 Database Audit

The NorthMet mineral resource estimate is based on the exploration drill-hole database available as of April 17, 2014. Drill hole data including collar coordinates, down-hole surveys, sample assay intervals, and geologic logs were provided by PolyMet in Microsoft Excel spreadsheets. The database was reviewed and validated by HRC prior to estimating mineral resources. The NorthMet database includes 114 (116) historic drill holes, 323 PolyMet drill holes, 240 vertical sounding holes, 15 depths to bedrock test pits, and 47 geologic holes from the surrounding area. Of the 739 drill holes, only 437 drill holes were used in the estimation, although many of the 437 holes include only select analytical information. The database was validated using Leapfrog Geo 3D® Version 2.0.0 software. Validation checks performed prior to loading the database into Datamine’s Studio 3 Version 3.24.25.0 mining software included:

· No overlapping intervals,

· Down-hole surveys at drill-hole collar,

· Consistent drill-hole depths for all data tables, and

· Gaps in the “from – to” data tables.

The analytical information used for the resource estimate includes copper, nickel, platinum, palladium, gold, silver, cobalt and sulfur.  All assay values Below Detection Limits (BDL) were assigned a value of one half of the detection limit, and missing or non-sampled intervals were assigned a value of zero (0). Table 12‑1 summarizes the validated analytical information utilized in the estimation of mineral resources.

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Table 12‑1: Summary of the Analytical Data Used in the Estimation of Mineral Resources

Metal	Missing Intervals	Assay Values	BDL Intervals
Cu (%)	1611	37196	791
Ni (%)	1611	37196	153
Pt (ppb)	1805	37002	10245
Pd (ppb)	1805	37002	1480
Au (ppb)	1805	37002	5211
Ag (ppm)	1731	37076	19304
Co (ppm)	1731	37076	1
S (%)	1971	36836	0

12.3 Certificates

HRC received original assay certificates in excel format for the samples collected in 2010 in the current database.  A random manual check of 10% of the database against the original certificates was conducted.  The error rate within the database is considered to be less than 1% based on the number of samples spot checked.

12.4 Adequacy Of Data

HRC reviewed PolyMet’s check assay programs and considers the programs to provide adequate confidence in the data.  Samples that are associated with QA/QC failures were reviewed and reanalyzed as necessary.

Exploration drilling, sampling, security, and analysis procedures were conducted in a manner that meets or exceeds industry standard practice. All drill cores and cuttings from PolyMet’s drilling have been photographed. Drill logs have been digitally entered into an exploration database organized and maintained in Gemcom. The split core and cutting trays have been securely stored and are available for further checks.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

This section was adapted from Senet’s Engineering Report entitled, NorthMet Copper Project: Feasibility Study Technical Report, Revision 2, dated March 2016 and results from the most recent pilot study investigation conducted by SGS on hydrometallurgical processes entitled, An Investigation into PLATSOL^TM Processing of the NorthMet Deposit, Project 12269-001 – Final Report dated April 20th, 2010.

13.1 Introduction

The NorthMet Deposit is hosted in the Duluth Complex in northeastern Minnesota. The Duluth Complex is a large, composite, grossly layered tholeiitic mafic intrusion. The sulfide mineralization of the complex contains metals (copper, nickel, cobalt, titanium and PGMs) that are of economic interest. A significant amount of metallurgical test work has been conducted on the Duluth Complex; therefore, the general metallurgy of the complex is fairly well understood.

Orway Mineral Consultants (OMC) in 2014 studied SAG Mill based comminution circuits for the Project.  This was done to assess if a SAG Mill based circuit would be practical for the Project and capable of rationalizing the existing 4-stage crushing circuit (total of 11 crushers) and 12 lines of Rod Mill + Ball Mill grinding circuits in the existing Erie concentrator.  Comminution test work results from SGS were interpreted by OMC and used to scope out a SAG mill based comminution circuit to process 32,000 STPD.  Further comminution test work was conducted by Hazen Research (Golden, Co.) in 2015 to confirm the comminution parameters.

The development of the current NorthMet flotation process flowsheet was based on test work (SGS, 2015) and includes the following:

· Flotation Test work conducted by SGS Lakefield (SGS) between 1998 and 2014, and

· Supplementary flotation test work conducted by SGS in 2015 and interpreted by Eurus Mineral Consultants (EMC) for circuit modeling and flotation plant design.

SGS conducted extensive flotation test work up until 2010. The work covered by SGS included significant amounts of batch and rate flotation test work on a number of samples provided by PolyMet. A flotation process block flow diagram was developed from the results and observations of the initial batch test work conducted by SGS. The process block flow diagram shown in Figure 13‑1 can be summarized into three main circuits as follows:

1. The Bulk Copper-Nickel Flotation circuit

2. The Copper-Nickel Separation Circuit

3. The Pyrrhotite Flotation Circuit

Pilot scale test work was conducted by SGS to demonstrate the flowsheet developed for the NorthMet process as indicated in Figure 13‑1. The results of the pilot test work are also included in the SGS report.

Additional flotation test work was requested of SGS in 2015 to fill in gaps in the flotation test work.  EMC conducted a flotation circuit simulation of the process flow based on the results obtained from both SGS's batch and pilot scale test work. The work that EMC conducted was initially targeted at simulating the pilot plant, and then to producing full production scale results. EMC's simulations were based on a throughput of 32,000 STPD. The results of the simulations were used to review the previous design and update the current process plant design basis and criteria.

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Source: SGS Flotation Report (2015).

Figure 13‑1: NorthMet Process Block Flow

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A second pilot plant program was carried out by SGS in 2009 to investigate hydrometallurgical processes. This is discussed in more detail starting from Section 13.6 of this report.

13.2 Comminution Circuit Test Work And Process Development

The comminution circuit was designed based on the work done by OMC and vendor information. The comminution circuit was modelled to be capable of processing 32,000 STPD and was based on the historical comminution results available from the test work conducted by SGS. The following comminution test work was conducted on three composite samples:

· SAG milling circuit (SMC) tests

· Abrasion index (Ai) tests

· Rod mill work index (RWi) tests

· Bond ball work index (BWi) tests

An Unconfined Compressive Strength (UCS) test was conducted on a composite of the 3 samples: Comp 1, Comp 2 and Comp 3. The comminution test work results are given in Table 13‑1.

Table 13‑1: Summary of Comminution Test Work Results

Parameter	Unit	Comp 1	Comp 2	Comp 3	UCS
BWi
1	kWh/t	14.8	15.0	16.0	-
2	kWh/t	16.3	15.4	15.1	-
3	kWh/t	15.7	15.2	15.7	-
Average	kWh/t	15.6	15.2	15.6	-
RWi	kWh/t	13.2	13.0	13.9	-
Ai	g	0.39	0.42	0.40	-
UCS
Min.	MPa	-	-	-	41.3
Max.	MPa	-	-	-	234.2
Average	MPa	-	-	-	108.6
JK Drop Weight Test
A		96.5	100	99.0	-
b		0.38	0.38	0.36	-
A × b		36.7	38.0	35.6	-
ta		0.24	0.26	0.22	-
SG		3.02	3.02	2.98	-

Further comminution test work was conducted by Hazen Research in February 2015 to confirm the historical comminution results. A summary of the comminution test work results is given in Table 13‑2.

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Table 13‑2: Summary of SMC Test Work Results Conducted by Hazen Research

Parameter	Units	Value
BWi	kWh/t	13.8
RWI	kWh/t	12.7
Abrasion Index, Ai	g	0.391
JK Drop Weight Test:
A		73.4
b		0.54
A × b		39.6
ta		0.29
Solids SG	lb/ft3	164

Table 13‑3 summarizes the mill specifications when applying parameters obtained from OMC's simulation.

Table 13‑3: Milling Circuit Design

Criteria	Unit	SAG Mill	Ball Mill
Diameter Inside Shell	m	12.19	7.32
Effective Grinding Length (EGL)	m	6.86	11.28
Imperial Mill Dimensions	ft × ft	40.0 × 22.5	24.0× 37.0
L:D Ratio	m/m	0.56	1.54
Discharge Arrangement		Grate	Overflow
Cone Angle	°	15	20
Speed Range	% Nc	60 - 80	Fixed
Speed – Duty	% Nc	67	75
Liner Thickness	mm	120	100
Ball Top Size	mm	125	50
Ball Charge – Duty	% Vol	5	20
Ball Charge – Maximum	% Vol	18	33
Total Load – Duty	% Vol	25	-
Total Load – Maximum	% Vol	35	-
Pinion/Shell Power – Duty	kW	12,900	7,490
Pinion/Shell Power – Maximum at 75% Critical Speed (Nc)	kW	22,830	10,820

13.3 Flotation Circuit Test Work And Process Development

Previous test work reports authored by SGS, and G&T Metallurgical Services, Kamloops, Canada between 2006 and 2014 were received and reviewed by EMC. These reports covered laboratory batch and locked cycle tests (LCTs) as well as pilot scale campaigns for the Bulk Cu-Ni and pyrrhotite circuits. The work also included laboratory scale test work conducted on the Bulk Cu-Ni concentrate. Kinetics were only conducted on selected rougher and cleaner streams as follows:

· Cu-Ni Bulk rougher feed

· Pyrrhotite rougher feed

· Cu-Ni separation rougher feed

· Cu-Ni Bulk rougher concentrate with regrind

· Cu-Ni separation 1st cleaner

· Pyrrhotite 1st cleaner feed with regrind

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The current flotation design is based on all of the test work conducted by SGS. This includes the recent flotation test work carried out by SGS in June 2015 to cover information gaps from previous SGS test work and to confirm the repeatability of the results and generate additional kinetic data for the various flotation stages.

In June 2009, SGS completed a small laboratory scale test work program on an alternative split cleaner circuit for the NorthMet mineralization, shown in Figure 13‑2. The test work program produced encouraging results compared to results from previous test work.  The previous flowsheet had produced a total Bulk sulphide concentrate and had a Cu-Ni separation on the concentrate to produce a salable Cu concentrate.

A decision was therefore made to carry out a small laboratory scale optimization program followed by a pilot plant campaign and a Cu-Ni separation program to demonstrate the suitability of this flowsheet option. The split cleaner flowsheet produces a good quality Bulk Cu+Ni concentrate which allows for easy separation of the Cu minerals from the Ni and Fe minerals to produce a good quality Cu concentrate and a salable Ni concentrate.  The Bulk circuit is then followed by a Pyrrhotite “scavenger” circuit to recover all the remaining sulphides and valuable minerals.  The circuit essentially treats the rougher and scavenger concentrates in separate cleaning circuits, and hence the label of “split cleaner” flowsheet.

On September 8, 2009, approximately 6.6 tons of a composite sample identified as C9 was delivered to SGS for the optimization test work and pilot program. A series of seven open circuit batch tests and two LCTs were carried out to establish the flotation kinetics of the C9 composite and to optimize process variables such as regrind targets, reagent dosages, and reagent addition points in preparation of the pilot plant campaign.

The pilot plant was only run on the front end of the circuit without the Cu/Ni separation stage. This was due to the fact that there was a very low mass recovery in the Cu-Ni 3rd cleaner concentrate. The pilot plant flowsheet including reagent addition points and dosages is shown in Figure 13‑2.

A total of six surveys were completed and each survey was balanced using the Bilmat mass-balancing software. The results of the pilot run are summarized in Table 13‑4.

Comparisons were made between the performance of the split cleaner flowsheet piloted in 2009 and the previous work conducted on different flowsheets. The performance of the 2009 pilot plant and the previous pilot work are shown in Figure 13‑3.

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Source: SGS Report (2009)

Figure 13‑2: Pilot Plant Flowsheet

Table 13‑4: Summary of Pilot Plant Test Work Results on Sample C9

Product	wt. %	Assays (%, ppm)						Distribution (%)
		Cu	Ni	S	Pt	Pd	Au	Cu	Ni	S	Pt	Pd	Au
Cu-Ni 3rd Cleaner Concentrate	1.48	18.2	3.41	27.7	2.41	10.5	1.33	89.1	58.0	66.1	65.1	69.4	61.3
Po 3rd Cleaner Concentrate	0.53	2.81	0.85	25.5	1.43	4.59	0.89	4.8	5.2	21.8	13.8	10.9	14.3
Combined Concentrate	2.01	14.1	2.74	27.0	2.15	8.97	1.21	93.9	63.2	87.9	78.9	80.3	75.6
Scavenger Tails	98	0.02	0.032	0.08	0.012	0.045	0.008	6.1	36.8	12.1	21.1	19.7	24.1
Feed	100	0.30	0.086	0.61	0.005	0.22	0.003	100	100	100	100	100	100
Source: SGS Report 2009

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Figure 13‑3: Comparative Recoveries between C9 Pilot Work and Previous Pilot Work

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The following conclusions were drawn:

· The split cleaner flowsheet test work produced a combined concentrate grade and recovery that was comparable with the results that were achieved in the 2008 pilot plant campaign and even exceeded the performance of historic pilot plant operations when taking into account the composite head grades.

· The Ni recovery in the final concentrate was the lowest of all the pilot plants. However, it must be noted that the head grade of 0.085% was also amongst the lowest with the exception of the C8 composite.

· Considering the very efficient recovery of the sulfides in the current pilot plant campaign, it is postulated that the C9 composite may have had more Ni units associated with non-sulphide gangue minerals

· The split cleaner flowsheet produced very good PGM recoveries when compared to previous pilot plant results, especially since the PGM head grades of the C9 composite were amongst the lowest of all samples tested.

· The Cu-Ni 3rd cleaner concentrate that was generated in the pilot plant was subject to four small-scale open-circuit Cu/Ni separation tests to establish suitable flotation conditions for a larger scale Cu/Ni separation LCT. The separation was deemed to produce a better Cu concentrate with an easier to conduct separation than from the previous bulk flotation circuit. The projected metallurgy of this LCT combined with the Pilot Plant results is shown in Table 13‑5.

Table 13‑5: Projected Metallurgy of Cu-Ni Separation LCT of C9 Pilot Cleaner Concentrate

Product	wt.%	Assays (%, ppm)			Distribution (%)
		Cu	Ni	S	Cu	Ni	S
Cu 5th Cleaner Concentrate	0.85	26.9	0.56	30.0	80.0	5.6	54.6
Po 3rd Cleaner Concentrate	0.53	2.81	0.85	25.5	4.8	5.2	21.8
Cu 1st Cleaner Scavenger Tail	0.14	7.33	7.50	20.9	3.5	12.1	5.1
Cu Rougher Tail (Ni Concentrate)	0.49	3.87	7.94	25.2	5.6	40.3	15.4
Combined Cu Tail (Ni Concentrate)	0.63	3.81	4.48	24.8	13.9	57.6	42.3
Calculated Head	100	0.30	0.086	0.61	100	100	100

The test work was also conducted on a composite sample identified as C10. The C10 composite was obtained from a shallow part of the NorthMet Deposit.  The EMC review also was to confirm the repeatability of the results and generate kinetic data for the various flotation stages.  A total of fifteen batch tests and a LCTs were conducted on the C10 composite and the results are summarized in Table 13‑6.

Table 13‑6: Summary of Laboratory Test Work Results on Sample C10

	wt.%	Assay (% or ppm)						Distribution (%)
		Cu	Ni	S	Pt	Pd	Au	Cu	Ni	S	Pt	Pd	Au
Cu Sep 4th Cl Concentrate	0.79	28.2	0.66	31.8	1.26	13.7	2.79	76.5	5.5	35.8	13.9	43.2	46.2
Cu Sep Ro Tail	0.48	3.36	6.75	17.8	5.22	8.97	0.41	5.6	34.3	12.2	35.0	17.2	4.1
Cu Sep 1st Cl Scv Tail	0.19	5.27	7.63	21.0	5.27	13.2	0.64	3.5	15.4	5.7	14.1	10.1	2.6
Combined Ni Concentrate	0.67	3.90	7.00	18.7	5.23	10.2	0.48	9.0	49.7	17.9	49.1	27.3	6.7
Po 3rd Cl Concentrate	1.07	1.17	0.67	21.3	0.66	2.36	0.27	4.3	7.5	32.3	9.9	10.0	6.1
Po Ro Tail	97.5	0.03	0.036	0.10	0.02	0.05	0.02	10.2	37.3	13.9	27.2	19.5	40.9
Feed	100	0.30	0.095	0.70	0.07	0.25	0.05	100	100	100	100	100	100

The parameters that were used for the design of the flotation plant are summarized in Table 13‑7.

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Table 13‑7: Flotation Stage Design Parameters

Parameter	Unit	Design
Cu-Ni Rougher Flotation
Grind (P80)	µm	120
pH		8.5 (natural)
Activator		-
Depressant		-
Cu-Ni Cleaner Flotation
Grind (P80)	µm	35
pH		8.5 (natural)
Activator		-
Depressant		CMC
Cu-Ni Separation Flotation
Grind (P80)	µm	15-25
pH		11.5 (lime)
Activator		-
Depressant		CMC
Po Rougher Flotation
Grind (P80)	µm	120
pH		8.5 (natural)
Activator		CuSO4
Depressant		CMC
Po Cleaner Flotation
Grind (P80)	µm	35
pH		8.5 (natural)
Activator		CuSO4
Depressant		CMC

13.4 Flotation Circuit Design

The split cleaner flowsheet test work resulted in increased performance when compared to previous test work, and as such, formed the basis for the flotation circuit design. The simulation and scale-up of the pilot test results to the full- scale plant was carried out by EMC. EMC was requested to review all the existing flotation test work data and use the information available to simulate a full-scale plant design for the NorthMet Deposit using the split cleaner flowsheet. A summary of EMC's work is presented in this section.

EMC's review of the available test work data revealed that sufficient rate tests were performed to kinetically characterize the ore and the various sub-circuits. The flotation performance of the C9 composite was simulated using appropriate kinetics from the C9 and C10 rate tests. C10 kinetics were used, in as-is or modified state, when the C9 kinetics were not representative of the flotation performance in that section of the circuit.

The split circuit flowsheet in Figure 13‑4, shows the streams that were rate tested or where the kinetics were derived.

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Figure 13‑4: General Block Flow – Rate Tested and Kinetic-Derived Process Streams from Report NM 1-2015 NorthMet Feb 2015

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13.4.1 Flotation Circuit Simulation

The simulation and scale-up of the pilot scale results into the production scale plant design were conducted using SUPASIM®, a proprietary flotation simulation program of EMC. SUPASIM® uses the rate data from the two component Kelsall rate equation as the input data and then adjusts the number of cells and cell aeration rate to project along the kinetic curves to determine the optimum time and hence cell volume requirements for each separation stage of the plant.  A total of some 60 case studies have been made using this technology.

EMC simulated the production scale plant design based on a throughput of 32,000 STPD.  The parameters used for the plant simulation and design are shown in Table 13‑8. These are the parameters that were adopted for the process plant design criteria.

Table 13‑8: Flotation Plant Simulation and Design Parameters

Parameter	Unit	Value
Throughput
Throughput	STPD	32,000
Throughput	STPH	1,340
Flotation Feed Solids	% w/w	33.2
Head Grades
Cu	% w/w	0.300
Ni	% w/w	0.086
Co	% w/w	0.010
Fe	% w/w	9.480
S	% w/w	0.610
Au	ppm	0.050
PGM (Rh, Pd, Pt)	% w/w	0.330

The production scale simulations were performed and parameters such as retention time and flotation volume requirements were produced.  EMC produced a mass balance using the results of the simulation. The mass balance analyzed the copper, nickel and sulfur elements. Recoveries and concentrate mass yields were calculated for each stage of the circuit. The simulation for the circuit is summarized in Table 13‑9.

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Table 13‑9: Summary of Flotation Circuit Simulation

Stream	Simulated Plant Mass Balance
	% Mass	% Solids	Pulp	% Cu	Cu % Rec	% Ni	Ni % Rec	% S	S % Rec
			Gpm (m3/h)
New Feed	100.00	33.2	13838 (3143)	0.300	100.0	0.086	100.0	0.61	100.0
Cu-Ni Bulk Rougher Concentrate	11.8	30.7	1810 (411)	2.26	89.0	0.44	60.6	3.67	71.1
Cu-Ni Bulk 1st Cleaner Concentrate	4.11	28.0	705 (160)	6.48	88.7	1.22	58.1	10.4	70.1
Cu-Ni Bulk 2nd Cleaner Concentrate	2.83	24.1	581 (132)	9.52	89.8	1.79	58.9	15.7	72.6
Cu-Ni Bulk 3rd Cleaner Concentrate	1.82	23.5	387 (88)	14.6	88.4	2.74	58.0	23.1	68.9
Cu-Ni Bulk 4th Cleaner Concentrate
Cu-Ni Bulk 1st Cleaner Tail	10.00	32.5	1422 (323)	0.018	0.6	0.022	2.6	0.14	2.2
Feed to Cu-Ni Sep Rougher	1.82	23.5	387 (88)	14.6	88.4	2.74	58.0	23.1	68.9
Cu-Ni Sep Rougher Concentrate	1.56	23.2	335 (76)	16.8	87.1	1.58	28.6	23.2	59.2
Cu-Ni Sep 1st Cleaner Concentrate	1.51	23.1	326 (74)	17.8	89.1	1.12	19.6	23.7	58.5
Cu-Ni Sep 2nd Cleaner Concentrate	1.29	23.0	282 (64)	20.1	86.3	0.81	12.2	27.1	57.3
Cu-Ni Sep 3rd Cleaner Concentrate	1.27	22.9	277 (63)	21.6	91.8	0.65	9.6	30.1	62.8
Cu-Ni Sep 4th Cleaner Concentrate	0.90	22.8	198 (45)	26	77.7	0.45	4.7	34.4	50.6
Cu-Ni Sep 5th Cu Cleaner Concentrate
Cu-Ni Sep 1st Cleaner Tail	0.66	23.8	137 (31)	4.30	9.4	3.13	23.9	7.93	8.6
Cu-Ni Sep Tail (Ni Concentrate)	0.92	24.2	189 (43)	3.49	10.8	4.96	53.3	12.1	18.2
Cu-Ni Bulk Rougher Tail	98.2	33.5	13451 (3055)	0.035	11.6	0.037	42.0	0.19	31.1
Feed to Po Rougher	98.2	33.5	13451(3055)	0.035	11.6	0.037	42.0	0.19	31.1
Po Rougher Concentrate	5.79	29.2	942 (214)	0.35	6.8	0.10	7.0	4.33	41.1
Po 1st Cleaner Concentrate	7.67	29.0	1321 (300)	0.33	8.5	0.10	8.8	13.8	173.8
Po 2nd Cleaner Concentrate	5.65	29.0	945 (215)	0.71	13.4	0.20	13.5	15.6	144.6
Po 3rd Cleaner Concentrate	0.52	28.9	88 (20)	3.08	5.4	0.82	5.0	26.1	22.4
Po 1stCleaner Tail	5.23	29.3	854 (194)	0.079	1.4	0.03	2.0	2.11	18.1
Po Rougher Tail	97.6	33.5	13363 (3035)	0.019	6.2	0.033	37.0	0.050	8.0

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The results of the simulation were used to size the flotation equipment as given in Table 13‑10.

Table 13‑10: NorthMet Tank Cell Sizing and Selection

EMC Tank Cell Sizing and Selection
Flotation Bank	Number of Cells	Cell Volume yd3 (m3)	Total Bank Volume yd3 (m3)	Nominal Residence Time (min)
Cu-Ni Bulk Rougher Bank	4	653 (500)	2612 (2000)	38
Cu-Ni Bulk 1st Cleaner Bank	4	210 (160)	840 (640)	60
Cu-Ni Bulk 2nd Cleaner Bank	3	131 (100)	393 (300)	88
Cu-Ni Bulk 3rd Cleaner Bank	2	131 (100)	262 (200)	83
Cu-Ni Bulk 4th Cleaner Bank	-	-	-	-
Total	13		4107 (3140)	269
Cu-Ni Sep Rougher Bank	3	65 (50)	210 (150)	91
Cu-Ni Sep 1st Cleaner Bank	3	65 (50)	210 (150)	107
Cu-Ni Sep 2nd Cleaner Bank	3	39 (30)	117 (90)	59
Cu-Ni Sep 3rd Cleaner Bank	3	39 (30)	117 (90)	63
Cu-Ni Sep 4th Cleaner Bank	3	39 (30)	117 (90)	69
Cu-Ni Sep 5th Cleaner Bank	3	26 (20)	78 (60)	50
Total	18		849 (630)	439
Po Rougher Bank	5	653 (500)	3265 (2500)	50
Po 1st Cleaner Bank	2	210 (160)	420 (320)	57
Po 2nd Cleaner Bank	2	131 (100)	262 (200)	83
Po 3rd Cleaner Bank	2	65 (50)	131 (100)	57
Po 4th Cleaner Bank			-	-
Total	11		4078 (3120)	247

13.5 Metallurgical Modelling For Recovery And Concentrate Quality

Total metal recovery was adapted from the SGS report “Flotation Grade-Recovery Study Phase II,” Project 11603-004.  This report presented the recovery of all the relevant metals as a function of the Cu head grade.  This data was then augmented with additional data from key laboratory samples and from pilot plant data.  This was done for two primary purposes:

· To further add to the dataset

· Compare pilot performance to the lab performance

The data found that the pilot data fit well with the laboratory data.  The data was then re-presented for all metals’ recovery as a function of their own head grade rather than to Cu head grade.  Although the head grades for all elements generally follow the Cu head grade well, it seemed more appropriate to present each metal as a function of its own head grade.  These plots are given in Figure 13‑5 through Figure 13‑12.

The next step was to build to a full metallurgical model from the total metal recovery curves as a function of the head grade.  The primary data to fill in all the output streams from the flowsheet (3 concentrates and 1 tailings) were taken primarily from the C-9 and C-10 testing.  These are the only two samples which have undergone rigorous “Split Cleaner” flowsheet testing.  Testing prior to this used a different flowsheet (bulk concentrate production which eventually lead to a Cu-Ni separation) and hence this data is not fully relevant for the individual products.  Data from two other lab samples tested were reviewed but were rejected since these samples only underwent simple batch testing and would therefore require data manipulation to reflect an LCT-type of result.

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The following steps were performed:

1. Calculate the total metal recovery.

2. Estimate the Pyrrhotite concentrate recoveries.

a. This was taken as the average recovery from the C-9 and C-10 samples.

b. This then allows calculating the Bulk Cu+Ni concentrate (Cu Separation circuit feed) recovery.

3. Calculate the recovery to the Cu concentrate as a fixed recovery factor for each metal from the Bulk Cu+Ni concentrate (i.e. 90% for Cu, 40% for Pt, etc.).

a. The Cu concentrate has some fixed grade targets of 27% Cu, 0.6% Ni and 31% S.  These are average values from the C-9 and C-10 testing.

b. The above recovery values and concentrate grade targets permit full calculation of the Cu concentrate assays, recoveries and the mass of product.

4. Calculate the Ni concentrate as the difference from Bulk Cu+Ni concentrate and the Cu concentrate.  This is done at a fixed concentrate assay of 20% S, again averaged from the C-9 and C-10 test work.

5. The final tails recovery is calculated as the difference of 100 less the total metal recovery determined in Step 1) above.  The %S in the tail is a function of the S head grade vs. recovery, which is different from the other elements.

6. The next step is a small iterative step (done within EXCEL) which estimates the total concentrate wt.% so that the Pyrrhotite concentrate and tails mass can be estimated.  (Tails mass equates to 100 less the total concentrate mass, and Po concentrate mass equates to tails less Bulk Cu+Ni concentrate).

7. With the mass estimated, then all the assays for the Pyrrhotite concentrate can be determined from the known recoveries and the mass is then iterated for a small adjustment to make the balance whole.

13.5.1 Cobalt

Cobalt is handled differently, mostly since the overall head grade vs. recovery trend is poor.  Cobalt is similar to Ni in that a notable portion of it is tied up in olivine and hence much of the cobalt is non-recoverable as non-sulphide.  Thus, for total recovery we have applied the average recovery for all the samples used for modelling.  The next assumption was that all the sulphide Co was associated with pentlandite; hence, we calculated out the Co assays for the concentrate streams as a simple ratio to the Ni assay.  The ratio was taken from the available mineral chemistry data.  This last assumption is reasonable as most of the sulphide Co is in pentlandite and only a small portion of the Co is as discrete Co minerals.  It is assumed that the discrete Co minerals will likely respond in a fashion similar to pentlandite.

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Table 13‑11 shows the overall mass balance for C-9, C-10 compared to the result of modelling the C-10 heads.

Table 13‑11: Summary of C-9 and C-10 Metallurgy Compared to Model

	C-9		C-10		Model C-10
	Assay	Recovery	Assay	Recovery	Assay	Recovery
Feed
Wt.%	100		100		100
Cu	0.30		0.29		0.29
Ni	0.065		0.095		0.095
Co ppm	86				75
Pt ppb	70		72		72
Pd ppb	220		250		250
Au ppb	30		48		48
Ag ppm	<2		1.3		1.3
S	0.61		0.70		0.70
Cu Concentrate
Wt.%	0.75		0.79		0.84
Cu	26.9	80.0	28.2	76.5	27.0	78.5
Ni	0.56	5.6	0.66	5.5	0.60	5.3
Co ppm	360				300	3.4
Pt ppb	1760	28.8	1260	13.9	2055	24.1
Pd ppb	11600	46.3	13700	43.2	13444	45.4
Au ppb	1280	40.9	2790	46.2	2381	41.9
Ag ppm	60		61.8	38.5	65.6	42.5
S	30	45.6	31.8	35.8	31.0	37.4
Ni Concentrate
Wt.%	0.73		0.67		0.48
Cu	4.16	8.8	3.90	9	5.25	8.7
Ni	7.08	51.7	7.00	49.7	10.39	52.7
Co ppm	3300				5194	33.4
Pt ppb	3767	36.3	5230	49.1	5395	36.1
Pd ppb	11200	23.1	10170	27.3	11588	22.3
Au ppb	3060	20.4	480	6.7	1042	10.5
Ag ppm	33		30.4	16.1	28.7	10.6
S	17.7	20.5	18.7	17.9	20.0	13.8
Po Concentrate
Wt.%	0.58		1.10		1.02
Cu	2.81	4.8	1.17	4.3	1.28	4.5
Ni	0.85	5.2	0.67	7.5	0.74	8.0
Co ppm	630				371	5.1
Pt ppb	1430	13.8	650	9.9	844	12.0
Pd ppb	4590	10.9	2360	10	2443	10.0
Au ppb	890	14.3	270	6.1	469	10.0
Ag ppm	18		8.2	6.9	12.7	10.0
S	25.5	21.8	21.3	32.3	24.0	35.1
Tails
Wt.%	98.0		97.5		97.7
Cu	0.020	6.1	0.030	10.2	0.024	8.2
Ni	0.032	36.8	0.036	37.3	0.033	34.0
Co ppm	57				45	58.2
Pt ppb	12	21.1	20	27.2	20	27.8
Pd ppb	45	19.7	50	19.5	57	22.3
Au ppb	8	24.4	20	40.9	19	37.7
Ag ppm			0.5	38.5	0.5	36.8
S	0.08	12.1	0.10	13.9	0.10	13.8

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13.6 Hydrometallurgical Test Work

The development of the current Phase II process flowsheet (Figure 13‑13) was based on the results of the following test work:

1. PLATSOL™ (autoclave) leaching of nickel and pyrrhotite concentrate,

2. Ferric iron reduction,

3. Copper Sulfide Precipitation of PGM,

4. Copper Concentrate Enrichment,

5. Residual Copper precipitation with NaHS, and

6. Mixed Hydroxide Precipitation (MHP) Recovery.

Figure 13‑13: Hydrometallurgical Pilot Plant Flowsheet

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Bench-scale tests and a pilot plant campaigns yielded promising PLATSOL™ autoclave leaching parameters for extraction of base metals and Au+PGMs from NorthMet concentrates (SGS Lakefield, 2006; SGS Minerals, 2005 and SGS, 2006). Results from the most recent continuous hydrometallurgical pilot plant program conducted by SGS (SGS, 2010) are summarized herein and are the basis for the hydrometallurgical process described in this Study.

13.6.1 PLATSOL™ Leaching Pilot Plant Testing

Nickel Concentrate and Copper Concentrate from 2008 flotation testing (C1) and a pyrrhotite concentrate and copper concentrate from 2009 flotation testing were tested with PLATSOL leach.  Head assays for the concentrates are presented in Table 13‑12.

The single pass autoclave retention time based on a 33-liter autoclave working volume at approximately 225°C was 64 minutes for campaign C1 and 119 mins for campaign C2.  The feed to the autoclave was 9.2-9.5% solid and O₂ over pressure ranged from 100-110 psi.  ACD pulp was filtered on filter pans without thickening or flocculation and residue recycling was initiated as soon as sufficient leach residue cake was available.  Filter cakes were repulped in ACD PLS and adjusted to target pulp density to reach a target of 100% solids recycling.

In this study, two campaigns were conducted for PLATSOL leach and copper enrichment pilot tests, using two copper concentrates: A nickel concentrate from the 2008 flotation testing (C1), and a pyrrhotite concentrate from the 2009 flotation testing (C2).  Each campaign had a runtime of 12-15 hours. Head assays for the concentrates are presented in Table 13‑12.

The PLATSOL continuous tests were conducted in a 33-liter (working volume) autoclave at approximately 225°C with residence times of 64 minutes for Concentrate C1 and 119 minutes for Concentrate C2, and an oxygen overpressure of 100 to 110 psi.  The pulp densities in the autoclave ranged from 9.2 to 9.5% solids after cooling water injection. Part of the autoclave discharge residue was recycled to the autoclave feed such that the residue stream mass is equal to the mass of fresh feed. The autoclave discharge (ACD) was filtered on filter pans without thickening or flocculation and residue recycling was initiated as soon as sufficient leach residue cake was available.  The recycled filter cakes were repulped with ACD pregnant leach solution (PLS) to the target feed pulp before feeding back to the autoclave.

Table 13‑12: Flotation Concentrate Head Assays Used in the Test Campaigns (C1 & C2)

Campaign	Sample Type	Ni (%)	Cu (%)	Fe (%)	Co (%)	Al (%)	Mg (%)	Cr (%)	Ca (%)	Zn (%)	Si (%)	S (%)	S2- (%)	Au (g/t)	Pt (g/t)	Pd (g/t)
C1	NiCon	3.44	5.66	34.7	0.18	1.82	1.91	0.07	1.16	0.06	5.68	24.4	23.3	0.9	3.35	10.3
C2	PoCon	0.8	2.17	32.4	0.04	1.39	2.07	0.04	0.84	0.07	5.21	25.3	23.2	0.62	0.97	3.32

Campaign C2 immediately followed Campaign C1, allowing uninterrupted solids recycling, which meant that campaign C1 leach residue was recycled with the new C2 feed early in the C2 campaign.  PLS from campaign C1 was collected 2 hours into campaign C2 before collection of C2 PLS commenced. The pH of both liquors was adjusted to 2.

Average autoclave feed flowrates are reported in Table 13‑3.

Table 13‑13: Average Autoclave Feed Flowrates

Campaign	Flot Con			ACD Recycling			Dilution Liquor	Total Flow
	% solids	PD, g/L	mL/min	% solids	PD, g/L	mL/min	mL/min	mL/min
C1	57%	1707	63	51%	1764	64	391	518
C2	51%	1676	36	49%	1721	41	201	278

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Average autoclave compartment temperatures over the last 4 hours of each campaign ranged from 220.3°C to 225.3 °C for C1 and 224.9°C to 227.0°C for C2.  Overall oxygen flowrates for both campaigns ranged from 36 to 45 L/min.

Metal recoveries were calculated after correction for mass losses using Si assays as the tie element. While the amounts of silicon that dissolved were minor, they were still corrected for.

ACD liquor and residue trends are shown in Figure 13‑14 and Figure 13‑15 respectively. The change over to C2 happened shortly before 4 Nov 00:00, which caused the Ni content in the liquor to decrease. PLATSOL^TM leaching was successful in both campaigns. Recoveries of base metal and PGMs into the leach liquors are reported in Table 13‑14.

Source: SGS PLATSOL™ Processing Report (2010).

Figure 13‑14: ACD Liquor Ni, Cu, Mg PLS Trends

Source: SGS PLATSOL™ Processing Report (2010).

Figure 13‑15: ACD Residue Trends

PLATSOL™ Leaching was successful in both campaigns leading to the base metal recoveries reported in Table 13‑14.

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Table 13‑14: Base Metal and PGM Recoveries

Campaign	Ni (%)	Cu (%)	Fe (%)	Co (%)	Al (%)	Mg (%)	Cr (%)	Ca (%)	Zn (%)	Si (%)	S2- (%)	Au (%)	Pt (%)	Pd (%)
C1	97.0	99.1	-0.4	98.1	25.5	33.8	10.1	-66.4	97.4	3.1	95.5	91.0	87.6	92.0
C2	95.5	99.0	3.7	96.7	45.0	61.4	-13.2	-12.3	99.1	2.1	97.4	84.0	94.2	95.9

13.6.2 Precipitation of PGMs by Copper Sulfide

The precipitation of platinum group metals (PGM) by CuS is similar to the cementation process based on following reactions:

The CuS is less noble than each of the Au, PdS, PtS, hence the PGMs in solution precipitate in exchange for Cu going into solution. The reaction is conducted at elevated temperatures to accelerate the reactions. The result is a mixed CuS-S-Au-PtS-PdS precipitate for refining.

The PGM Precipitation circuit consisted of a preheat tank, two PGM precipitation tanks and a SO₂ reduction tank. Autoclave filtrates from campaigns were heated to 95°C in the preheat tank, sparged with gaseous SO₂ to reduce ferric iron in the SO₂ reduction tank. The addition of SO₂ was controlled by online ORP measurements.

In the first PGM tank, dissolved PGMs were precipitated onto synthetic CuS beads injected into the tank (target 10 g/L CuS concentration), then filtered onto Buchner filters. Filtered solids were repulped in the second tank filtrate and recycled back to the first tank to reduce the amount of CuS required.  Summarized conditions for the PGM Circuit are presented in Table 13‑15.

Table 13‑15: Summary of PGM Precipitation Operating Parameters

Campaign	Flow rate	RT		Temps			ORP			CuS (dry)
	Feed	PGM 1	PGM 2/3	PGM1	PGM2	PGM3	PGM1	PGM2	PGM3	fresh	rec.	total	conc
	mL/min	Min		°C			(mV)			g/min			g/l
C1	61	73	87	97	96	95	446	452	498	0.2	0.0	0.2	3.9
	64	69	84	98	96	95	401	390	375	0.6	0.1	0.7	10.4
	60	73	81	95	95	95	412	381	357	0.2	0.9	1.1	18.0
	60	73	78	96	96	95	445	382	359	0.2	0.8	1.0	16.2
C2	63	70	83	95	95	95	423	380	361	0.1	0.8	0.9	14.2
	62	71	83	95	95	88	402	366	356	0.1	0.4	0.5	7.5
	63	70	84	95	95	95	417	369	360	0.0	0.7	0.8	12.0
	67	65	85	95	95	95	400	363	358	0.3	0.5	0.8	11.2

Table 13‑16 compares the PGM Precipitation circuit feed liquor composition to the PGM Precipitation filtrate composition.

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Table 13‑16: Comparison between PGM Precipitation Circuit Feed and Filtrate Concentrations

Campaign	Ni mg/L	Cu mg/L	Fe mg/L	Fe(II) mg/L	Co mg/L	Al mg/L	Mg mg/L	Cr mg/L	Ca mg/L	Zn mg/L	Si mg/L	Cl mg/L	Au mg/L	Pt mg/L	Pd mg/L
PGM Feed Liquor
C1	23000	7500	1970	50	1100	820	4800	21	540	480	430	9620	0.05	0.18	0.72
C2	11000	4800	5500	79	540	1900	6600	32	670	520	350	10700	0.04	0.20	0.63
PGM Filtrate
	Ni mg/L	Cu mg/L	Fe mg/L		Co mg/L	Al mg/L	Mg mg/L	Cr mg/L	Ca mg/L	Zn mg/L	Si mg/L		Au mg/L	Pt mg/L	Pd mg/L
C1	18000 20000	6100 6700	2400 2000	--	880 920	430 640	3900 4300	11 16	450 480	490 410	230 350	-- --	<0.01 <0.01	0.01 <0.01	<0.01 0.01
C2	18000 13000 12000	6500 5300 4800	3100 4900 5300	-- -- --	840 580 550	1100 1700 1900	5400 6100 6400	25 27 27	560 640 690	460 520 530	380 380 360	-- -- --	<0.01 <0.01 0.01	<0.01 <0.01 0.01	0.01 <0.01 0.01

Table 13‑16 shows that in both campaigns the precipitation with synthetic CuS beads was successful at clearing all PGM elements in solution to less than 0.01 mg/L.  The final precipitate of the PGM Precipitation Circuit yielded as much as 244 g/t Pd.

13.6.3 Copper Concentrate Enrichment

In the copper enrichment (CuE) stage of the pilot study, soluble copper in the PGM filtrate is mixed with copper concentrate. The following metathesis reactions are thought to occur resulting in an enriched copper grade and Ni & Fe dissolution.

CuFeS₂ + CuSO₄ = 2CuS + FeSO₄

CuFe₂S₃ + 2CuSO₄ = 3CuS + 2FeSO₄

Fe₇S₈ + 7CuSO₄ = 7CuS + 7FeSO₄ + S₀

Nickel Sulfides also react to provide lower Ni in the copper concentrate.

NiS + CuSO₄ = CuS + NiSO₄

Campaign C1 PLS was contacted with the corresponding copper concentrate from the 2008 flotation test program and Campaign C2 PLS was contacted with copper concentrate from the corresponding 2009 flotation program. The process was conducted in three tanks CuE1, CuE2 and CuE3, with only the first tank heated to the reaction temperature and the last two tanks insulated.

Table 13‑17 presents the feed rates and operating conditions employed during copper enrichment of C1 and C2.  Discharge from CuE3 was filtered on filter pans with no washing. The filter cakes were then repulped in CuE3 filtrate and recycled back to CuE1. The target weight ratio of recycled over fresh concentrate was 1. However, Table 13‑17 shows that actual values after commissioning were more in the order of 0.5 to 0.7.

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Table 13‑17: Operating Conditions and Feed Parameter for Copper Concentrate Enrichment

Conc	Feed Rate, mL/min	Fresh Cu Conc, g/min	Recycle Cu Conc, g/min	Ratio, Recycle to Fresh	Temperature, °C			ORP, mV			Pulp Density, g/L
					CuE1	CuE2	CuE3	CuE1	CuE2	CuE3	CuE1	CuE2	CuE3
C1	65 51 55	10.8 11.5 8.1	0 1.5 3.8	0 0.1 0.5	93 95 90	66 74 82	50 53 60	369 304 335	335 257 277	364 346 319	1189 1245 1270	1211 1200 1288	1203 1243 1278
C2	58 63 63 64	9.9 12.6 13.5 9.6	4.6 4.4 6.1 7.0	0.5 0.3 0.5 0.7	89 87 82 81	79 63 66 66	62 54 54 55	319 298 301 308	227 262 250 277	326 309 298 324	1281 1265 1273 1271	1243 1270 1281 1311	1262 1269 1280 1263

Results indicated that the reactions were stable at temperatures as low as 60-70°C and retention times as little as 2-3 hours (data not shown) and that there was a distinct correlation between residual soluble copper and ORP (Figure 13‑16). Hence, ORP can be used to gauge the level of residual copper providing useful opportunities for process control.

Figure 13‑16: Correlation between Cu and ORP Observed for Copper Enrichment Trials

The material was pulped to a target pulp density and head samples were assayed. Composite liquor and residue assays were also obtained and are presented in together with the head assays in Table 13‑18. These data show that no PGM metals were lost to the filtrate (all assays reported <0.01 mg/L).

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Table 13‑18: Head and Copper Enrichment Solids and Filtrate Composite Assays

Campaign	NI %	Cu %	Fe %	Co %	Al %	Mg %	Cr %	Ca %	Zn %	Si %	S %	S2- %	Au g/t	Pt g/t	Pd g/t
Head Assays
Cu Con (C1)	0.38	30.5	33.5	0.018	0.09	0.47	<0.004	0.07	0.038	1.23	32.7	30.5	1.32	1.13	5.76
Cu Con (C2)	0.64	30.5	31.5	0.025	0.15	0.36	<0.006	0.36	0.056	1.21	31.1	29.8	1.6	1.44	9.24
Copper Enrichment Cu3 Solids Assays
Cu Con (C1)	0.33 0.31 0.39	26.5 31.2 30.7	30.4 24.3 30.3	0.02 0.02 0.02	0.21 0.11 0.09	0.66 0.39 0.33	<0.004 <0.004 <0.004	0.1 0.06 <0.04	0.062 0.045 0.043	1.95 1.1 0.9	31.4 31 31.6	29.8 30.9 31.6	nss 1.3 1.7	nss 1.1 1.5	nss 5.2 6.4
Cu Con (C2)	0.39 0.52 0.55	30.7 30.5 29.7	30.3 28.5 29.4	0.02 0.02 0.02	0.09 0.12 0.14	0.33 0.38 0.41	<0.004 <0.004 <0.004	<0.04 <0.04 <0.04	0.043 0.049 0.054	0.9 1.11 1.23	31.6 32 32.7	31.6 32 31.3	1.7 1.6 1.6	1.5 1.3 1.3	6.4 7.7 8.5
Copper Enrichment Cu3 Filtrate Assays
	Ni g/L	Cu g/L	Fe g/L	Co g/L	Al g/L	Mg g/L	Cr mg/L	Ca g/L	Zn g/L	Si mg/L	Cl g/L	Au mg/L	Pt mg/L	Pd mg/L	-
Cu Con (C1)	21 17 17	5.4 1.6 0.29	8.3 8.8 8.9	1.10 0.89 0.86	0.34 0.59 0.89	5.4 4.5 4.1	5 10 19	0.69 0.66 0.67	0.69 0.51 0.48	200 290 390	9.31 7.89 7.90	<0.01 <0.01 <0.01	<0.01 <0.01 <0.01	<0.01 <0.01 <0.01	-
Cu Con (C2)	17 15 11	0.29 0.24 0.25	8.9 9.3 9.8	0.86 0.67 0.48	0.89 1.40 1.80	4.1 4.9 5.8	19 23 25	0.67 0.81 0.88	0.48 0.48 0.51	390 440 390	7.90 9.07 9.12	<0.01 <0.01 <0.01	<0.01 <0.01 <0.01	<0.01 <0.01 <0.01	-

In campaign C1, copper levels decreased from ~6.5 g/L in PGM filtrates to <0.3 g/L Cu, while iron levels increased from ~2.5 g/L Fe to 8.9 g/L. In campaign C2, copper levels decreased from 0.29 g/L to 0.25 g/L Cu, while iron levels increased from 8.9 g/L to 9.8 g/L Fe. Nickel and cobalt dissolution from the copper concentrates was calculated to be 5.6% and 1.8%, respectively in campaign C1, and 29.1% and 20%, respectively in campaign C2.

No PGM losses from the copper flotation stream were observed based on the consistent filtrate assays of <0.01 mg/L for Au, Pt, and Pd compared to PGMs contained in the feed/head assays.

13.6.4 Residual Copper Precipitation

Residual soluble copper recovered in the depleted liquor from the copper enrichment stage was precipitated with NaHS (37.5 g/L) in duplicate titanium tanks.  Table 13‑19 presents the parameters used for this stage in the process.

Table 13‑19: Soluble Copper Precipitation Parameters

Tanks	2
Volume per tank (L)	7.4
Average NaHS Feed Flow (mL/min)	65
RT per tank (min)	114

NaHS tanks were not heated (to minimize corrosion), but the copper enrichment filtrate was preheated in a separate glass vessel.  NaHS addition/flows were governed by monitoring ORP levels; as a direct correlation between ORP measurements and soluble copper concentrations was observed (Figure 13‑17) in test samples and data acquisition.  In general, an ORP level of less than 150 mV was required to achieve a target concentration of 10 mg/L soluble Cu or less.  NaHS consumption was calculated to be 0.027 mol/h with a corresponding copper throughput of 0.015 mol/h for a 2:1 mole ratio of NaHS to copper.  Copper recovered in the NaHS product filter cakes produced a copper grade of approximately 35% (Table 13‑20) for both campaigns, C1 & C2. Table 13‑20 also indicates that some PGMs were precipitated out of solution during this stage.

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Figure 13‑17: Correlation Between ORP and Soluble Copper Concentration

Table 13‑20: NaHS Product Filter Cake Assays

	Ni %	Cu %	Fe %	Co %	Al %	Mg %	Cr %	Ca %	Zn %	Si %	S %	S= %	Au g/t	Pt g/t	Pd g/t
Cake 1	2.04	35.0	1.12	0.16	0.05	0.041	0.005	<0.05	0.029	0.62	30.3	25.3	0.05	0.09	0.19
Cake 2	1.73	34.8	1.51	0.11	0.26	0.11	<0.006	<0.05	0.018	1.27	39.9	20.2	0.09	0.1	0.48

The Cu-NaHS filtrate streams were then subjected to an Fe/Al removal stage followed by two stages of mixed hydroxide precipitation (MHP), ending with a magnesium removal stage.

13.6.5 Bulk Iron/Aluminum Removal

Fresh lime (CaCO₃) was used to precipitate the Fe and Al from the Cu-NaHS filtrate to achieve final soluble Fe and Al concentrations of less than 10 ppm and 30 ppm, respectively. The filtrate was heated to 80°C, agitated and sparged with oxygen. Dry lime was added to achieve a target pH of approximately 4.0. Supernatant samples were analyzed for Fe and Al periodically while maintaining the target pH. Once Fe and Al concentration targets had been achieved, pulps were filtered hot and the products assayed. Analysis of the final supernatant showed that Fe and Al concentrations had both been reduced to <5 mg/L. The amount of limestone used in the Fe/Al removal stage ranged from 61.3 kg limestone per m³ Cu-NaHS filtrate treated in C1 to 74.6 kg limestone per m³ Cu-NaHS filtrate treated in C2. Analysis of the precipitate also showed that some nickel and cobalt precipitated along with Fe and Al as was observed in a previous study (SGS, 2006).

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13.6.6 Mixed Hydroxide Precipitation (MHP)

Filtered Fe/Al precipitated solids were repulped in deionized water and combined with remaining filtrate from the Fe/Al removal stage for each campaign. The resultant solutions were heated and agitated prior to adding a Magnesium Oxide (MgO) pulp (Magchem 30™) to precipitate Ni and Co in Stage 1. Similarly, the filtrate and repulped filtrate produced in Stage 1 MHP was heated and mixed with hydrated lime to further recover more Ni and Co in the precipitate in Stage2. Table 13‑21 shows test conditions employed for both stages of the MHP process for the two campaigns, C1 & C2. ORP and pH were monitored constantly for both stages and samples were taken periodically. When target Ni concentrations were achieved, testing was discontinued.

Table 13‑21: Test Conditions, Target Ni Concentrations and Ni and Co Feed Concentrations for MHP Tests

	Stage 1-C1	Stage 1-C2	Stage 1-C1	Stage 1-C2
Feed Source	Fe/AL removal filtrate		Stage 1 MHP filtrate
Feed Volume (L)	69.6	100	63.6	93.3
Reagent	MgO	MgO	Ca(OH)2	Ca(OH)2
Reagent Pulp Density % (w/w)	20	20	20	20
Target initial pH	-	-	7.3	7.3
Cumulative Reagent Addition (g)	3445	3189	1419	1508
Target Temp. °C	70	70	65	65
Target soluble Ni conc. (mg/L)	20%	20%	10	10

For Stage 1, fresh 20% w/w MgO was added at an initial target dosage of approximately 0.65 kg of MgO per kg of Ni+Co based on previous results (SGS, 2006). Similar results were obtained for both campaigns whereby the Ni concentration in samples taken at the 0.65 dosage rate measured more than 99% of the 80% Ni precipitation anticipated. In the final Stage 1 filtrate for C1, 83% of the Ni was precipitated along with 94% of the Co; whereas, for C2, 78% of the Ni was precipitated and only 89% of the Co was precipitated as shown in Table 13‑22.

Table 13‑22: MHP Stage 1 Final Product Analysis and Distribution for Campaigns C1 & C2

	Vol L, g	Assays					Distribution
		Ni mg/L, %	Co mg/L, %	Zn mg/L, %	Fe Mg/L, %	Mg Mg/L, %	Ni %	Co %	Zn %	Fe %	Mg %
Campaign C1
Feed (Bulk Fe/Al-C1)	69.6	14900	595	350	0.8	4400	-	-	-	-	-
Primary Filtrate	63.6	2580	32.4	<2	<0.2	10000	17.1	4.4	0.6	1.0	99.0
Repulp Wash 1	60.4	282	2.07				1.8	0.3	0.0	0.0	0.0
Repulp Wash 2	56.1	141	1.2				0.8	0.1	0.0	0.0	0.0
Displ. Wash	50.3	128	1.28				0.7	0.1	0.0	0.0	0.0
Residue	1.499	50.9	2.96	1.52	0.081	0.45	79.6	95.0	99.4	99.0	1.0
						Total	100.0	100.0	100.0	100.0	100.0
Campaign C2
Feed (Bulk Fe/Al-C1)	100.0	8760	354	270	0.8	4100	-	-	-	-	-
Primary Filtrate	93.3	1980	37.4	2	<0.2	7600	21.7	8.7	0.7	3.3	98.7
Repulp Wash 1	62.4	207	3.23				1.5	0.5	0.0	0.0	0.0
Repulp Wash 2	59.9	115	1.6				0.8	0.2	0.0	0.0	0.0
Displ. Wash	45.4	76.4	1.34				0.4	0.2	0.0	0.0	0.0
Residue	1.3	50.5	2.86	2.11	0.043	0.73	75.5	90.4	99.3	96.7	1.3
						Total	100.0	100.0	100.0	100.0	100.0

For Stage 2, an initial dosage of 1211 g of 20% (w/w) Ca(OH)₂ was estimated to precipitate the remainder of the Ni to below the 10 mg/L for C1 and 1361 g was estimated for C2 in Stage 2.  Actual cumulative 20% (w/w) Ca(OH)₂ additions in Stage 2 to precipitate Ni to at (or below) the 10 mg/L target concentration were within 20% and 10% for C1 and C2, respectively.  Hydrated lime consumption to achieve a solution pH upwards of 7.5 ranged from 3.2 to 4.5 kg per m³ Stage 1 filtrate tested.  The composition of the precipitate produced in Stage 2 ranged from 20.8% to 21.9% Ni and 0.29% to 0.38% Co. Mg co-precipitation was low (data not shown).

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13.6.7 Magnesium Removal

Bulk magnesium removal was carried out on Stage 2 MHP filtrates including the repulped filtrate. Test conditions for filtrates from both Campaigns (C1 & C2) in agitated heated tanks, are presented in Table 13‑23.

Table 13‑23: Test Conditions for Bulk Magnesium Removal

Campaign	C1	C2
Feed Source	Stage 2 MHP filtrate
Feed Volume (L)	66.7	87.9
Reagent	Ca(OH)2	Ca(OH)2
Reagent Pulp Density % (w/w)	20	20
Target initial pH	8.0	8.0
Estimated Reagent Addition (g)	6220	6787
Cumulative Reagent Addition (g)	6257	6811
Target Temp. °C	50	50
Target Mg precipitation	50%	50%

The amount of hydrated 20% slurry w/w lime required to precipitate 50% of the Mg was calculated based stoichiometrically on the Mg assay obtained for the Stage 2 MHP filtrate. Test results for Mg assay in Stage 2 MHP filtrate for C1 decreased 59% from 9.3 g/L to 4.3 g/L at pH 8.6 and decreased 60% for C2 from 7.7 to 4 g/L at pH 8.3.

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14 MINERAL RESOURCE ESTIMATES

14.1 Data

Zachary J. Black, RM-SME, of HRC is responsible for the resource estimate presented here. Mr. Black is a qualified person as defined by NI 43-101 and is independent of PolyMet. HRC estimated the mineral resource for the NorthMet polymetallic Project from drill-hole data constrained by geologic boundaries with an Ordinary Kriging (“OK”) algorithm. Datamine Studio 3® software was used in combination with Sage 2001 for the variography and Leapfrog Geo® for the geologic model. The metals of interest at NorthMet are copper, nickel, cobalt, platinum, palladium, gold, silver, and sulfur.

The mineral resources reported in this technical report have been classified as Measured, Indicated and Inferred in accordance with standards defined by the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) “CIM Definition Standards - For Mineral Resources and Mineral Reserves,” prepared by the CIM Standing Committee on Reserve Definitions and adopted by the CIM Council in May 2014. Each individual mineral resource classification reflects an associated relative confidence of the grade estimates.

14.2 Block Model Physical Limits

HRC created a rotated three-dimensional (“3D”) block model in Datamine Studio 3® mining software. The block model was created with individual block dimensions of 50x50x50 feet (xyz) rotated 33.94° west of north. The model origin is located at 727,575 northing, 2,896,310 easting, and at an elevation of 1,200 ft below sea level. The block model extends 22,500 ft (450 blocks) in the easting direction, 10,000 ft (200 blocks) in the northing direction, and vertically 3,000 ft (60 blocks) to an elevation of 1,800 ft asl.  All of the block model coordinates are stored as UTM WGS 84, Zone 12 meters. All property and minerals within the block model extents are owned or claimed by PolyMet.

14.3 Geological Models

The NorthMet Project geology is divided into 3 formations consisting of the Biwabik Iron Formation (“BIF”), the Virginia Formation and the Duluth Complex. The Duluth Complex is comprised of 7 main lithological units (1 through 7) and is the primary host of mineralization.  HRC used Leapfrog Geo to model the stratigraphic sequence (bottom to top) consisting of the BIF, Virginia Formation, Unit 1, Unit 2 (Units 2 and 3 combined), Unit 4 (Units 4 and 5 combined), Unit 6, Unit 7, and overburden. The Magenta Zone, a smaller mineralized zone that cuts through Units 3 through 7 but resides primarily within 5 and 6, was modeled from select intercepts provided by PolyMet.  Figure 14‑1 depicts a typical easterly facing geologic cross-section from the geologic model with the Magenta Zone highlighted.

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Figure 14‑1: Estimation Domains

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14.3.1 Density

A total of 6,975 density measurements have been made on core to date using a variety of methods. Typically, measurements have been completed on core samples that have not been oven dried or sealed.  This can result in an overstatement in density due the inclusion of water that would typically be dried out in the oven; although the difference is expected to be less than 1%.

HRC considers that the densities presented in Table 14‑1, including the average specific gravity determinations sorted by unit (October 2007 dataset), are appropriate for use in estimation.

Table 14‑1: Specific Gravity Average per Unit (October 15 Dataset)

Unit	Mean	Count
1	2.98	2,381
3 (2+3)	2.92	1,818
5 (4+5)	2.90	1,266
6	2.90	902
7	2.92	326
20	2.77	273
30	3.17	9
All Units	2.93	6,975

14.4 Exploratory Data Analysis

HRC completed an Exploratory Data Analysis (“EDA”) on the copper, nickel, platinum, palladium, gold, silver, cobalt, and sulfur analytical information contained in the NorthMet exploration database. The purpose of an EDA is to summarize the main characteristics of the data provided using both statistical and visual methods. HRC utilized Leapfrog Geo (“Geo”) and ioGas Software to analyze the assay data.

14.4.1 Sample Statistics

A statistical analysis of each metal within each unit and the Magenta Zone was completed. Descriptive statistics by metal and domain are presented in Table 14‑2 through Table 14‑9.

Table 14‑2: Copper Sample Statistics

Copper Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		%	%	%	%	%
1	22,050	0.001	4.89	0.21	0.13	0.23	1.08
3	9,269	0.001	4.17	0.07	0.02	0.15	2.12
5	3,968	0.001	1.96	0.11	0.03	0.17	1.56
6	2,016	0.001	2.13	0.13	0.03	0.2	1.55
7	573	0.001	1.21	0.03	0.02	0.08	2.57
20	1,342	0.001	1.50	0.06	0.02	0.13	2.24
30	4	0.001	0.013	0.01	0.01	0.01	0.88
2000	2,352	0.001	2.13	0.24	0.17	0.23	0.96

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Table 14‑3:Nickel Sample Statistics

Nickel Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		%	%	%	%	%
1	22,050	0.001	1.170	0.07	0.05	0.06	0.86
3	9,269	0.001	0.460	0.03	0.03	0.03	0.93
5	3,968	0.001	2.359	0.04	0.03	0.05	1.25
6	2,016	0.001	0.294	0.05	0.04	0.03	0.7
7	573	0.011	0.183	0.04	0.03	0.02	0.58
20	1,342	0.001	0.462	0.02	0.01	0.04	1.56
30	4	0.002	0.012	0.01	0.01	0.01	0.75
2000	2,352	0.001	0.410	0.07	0.05	0.04	0.63

Table 14‑4: Platinum Sample Statistics

Platinum Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		ppb	ppb	ppb	ppb	ppb
1	22,050	0.5	1535	45.71	20	65.87	1.44
3	9,269	0.5	4780	25.77	7	70.52	2.74
5	3,968	2.5	638	41.3	11	69.9	1.69
6	2,016	2.5	1430	57.63	19	105.97	1.84
7	573	2.5	1430	20.38	7	71.44	3.5
20	1,342	0.5	305	9.70	2.5	21.86	2.26
30	4	2.5	6	3.38	2.5	1.75	0.52
2000	2,351	2.5	1390	95.63	60	106.05	1.11

Table 14‑5: Palladium Sample Statistics

Palladium Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		ppb	ppb	ppb	ppb	ppb
1	22,050	0.5	10386	175.12	72	263.18	1.5
3	9,269	0.5	6610	78.67	14	211.44	2.69
5	3,968	0.5	2690	106.02	18	205.89	1.94
6	2,016	0.5	3680	144.38	35	286.79	1.99
7	573	0.5	2860	36.60	9	147.23	4.02
20	1,342	0.5	2453	30.90	4	102.6	3.32
30	4	0.5	5	2.13	1.5	2.02	0.95
2000	2,351	0.5	3540	254.86	149	299.19	1.17

Table 14‑6: Gold Sample Statistics

Gold Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		ppb	ppb	ppb	ppb	ppb
1	22,050	0.5	1926	24.12	12	43.06	1.79
3	9,269	0.5	3150	14.11	4	47.94	3.4
5	3,968	0.5	760	20.21	6	36.14	1.79
6	2,016	0.5	545	24.13	8	41.49	1.72
7	573	0.5	388	8.05	3	25.1	3.12
20	1,342	0.5	188	6.28	3	11.42	1.82
30	4	0.5	3	1.25	0.75	1.19	0.95
2000	2,351	0.5	3150	44.85	28	80.07	1.79

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Table 14‑7: Silver Sample Statistics

Silver Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		ppm	ppm	ppm	ppm	ppm
1	22,050	0.05	50.5	0.79	0.5	0.97	1.23
3	9,269	0.05	15.6	0.35	0.25	0.54	1.53
5	3,968	0.05	11.1	0.51	0.25	0.62	1.24
6	2,016	0.05	12.1	0.57	0.25	0.74	1.31
7	573	0.1	4.5	0.3	0.25	0.29	0.95
20	1,342	0.1	3.9	0.45	0.25	0.47	1.04
30	4	0.25	0.7	0.43	0.38	0.22	0.51
2000	2,351	0.05	12.1	0.86	0.5	0.90	1.06

Table 14‑8:Cobalt Sample Statistics

Cobalt Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		ppm	ppm	ppm	ppm	ppm
1	22,050	2	713	68.13	62	31.34	0.46
3	9,269	1	361	53.6	48	22	0.41
5	3,968	0.5	421	54.72	49	18.54	0.34
6	2,016	1	491	65.25	62	20.42	0.31
7	573	21	160	70.66	61	29.45	0.42
20	1,342	2	385	35.31	26	30.65	0.87
30	4	4	23	12.75	12	9.67	0.76
2000	2,351	1	232	66.00	64	19.04	0.29

Table 14‑9: Sulfur Sample Statistics

Sulfur Sample Descriptive Statistics
Unit	Number	Minimum	Maximum	Mean	Median	Std. Dev.	COV
		%	%	%	%	%
1	22,050	0.01	26.1	0.63	0.4	0.81	1.29
3	9,269	0.01	10.8	0.19	0.05	0.5	2.58
5	3,968	0.01	12.22	0.24	0.07	0.43	1.81
6	2,016	0.01	3.62	0.20	0.05	0.31	1.56
7	573	0.01	2.67	0.06	0.03	0.17	2.75
20	1,342	0.01	10.75	1.62	0.89	1.62	1
30	4	0.24	2.29	0.75	0.24	1.02	1.36
2000	2,352	0.01	4.41	0.39	0.29	0.38	0.98

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14.4.2 Correlation Analysis

HRC completed a correlation analysis on each metal within each unit (restricted to the Duluth complex). The correlation matrix shown in Table 14‑10, created using the nonparametric Spearman Rank method, identifies a good overall correlation between the metals, particularly copper. The overall correlation between copper and the other metals is relatively consistent, as illustrated in Figure 14‑2.

Table 14‑10: Spearman Rank Correlation Matrix

Correlation	Cu (%)	Ni (%)	Pt (ppb)	Pd (ppb)	Au (ppb)	Ag (ppm)	Co (ppm)	S (%)
Cu (%)	1	0.85	0.78	0.86	0.86	0.76	0.62	0.86
Ni (%)	0.85	1	0.75	0.81	0.77	0.74	0.83	0.67
Pt (ppb)	0.78	0.75	1	0.9	0.84	0.67	0.52	0.59
Pd (ppb)	0.86	0.81	0.9	1	0.88	0.67	0.55	0.67
Au (ppb)	0.86	0.77	0.84	0.88	1	0.71	0.53	0.72
Ag (ppm)	0.76	0.74	0.67	0.67	0.71	1	0.56	0.67
Co (ppm)	0.62	0.83	0.52	0.55	0.53	0.56	1	0.51
S (%)	0.86	0.67	0.59	0.67	0.72	0.67	0.51	1

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Figure 14‑2: Copper Correlation Plots

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14.4.3 Contact Plot Analysis

HRC examined the relationship of mineralization across the contacts of each unit model. This examination was completed on copper only, assuming that the other metals would behave in a similar manner due to the higher correlation coefficients.

Contact plots are created by averaging the grade of copper over a set distance from the modeled lithologic boundary. The plotted results assist in understanding the relationship of grades as they approach and cross geologic boundaries. This relationship is used in determining whether these boundaries are treated as hard or soft boundaries during the estimation process.

The contact between the Virginia Formation and the base of Unit 1 forms a hard boundary with the mineralized material residing within Unit 1, as shown in Figure 14‑3.

Figure 14‑3: Contact Plot Virginia Formation and Unit 1

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The contact between Unit 1 and Unit 3 is a hard boundary with higher grades found within Unit 1 trending along the contact.  A decrease in average grade across the boundary into Unit 2 suggests two different sample populations in Units 1 and 3. See Figure 14‑4.

Figure 14‑4: Contact Plot Unit 1 and Unit 3

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Figure 14‑5 shows the contact between Units 3 and 5 is mineralized, and grading into lower grade material away from the contact.

Figure 14‑5: Contact Plot Unit 3 and Unit 5

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The contact between Unit 5 and Unit 6 is gradational with a slight increase of grade in Unit 6. See Figure 14‑6.

Figure 14‑6: Contact Plot Unit 5 and Unit 6

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Figure 14‑7 shows that the copper grades across the contact between Unit 6 and Unit 7 are relatively similar. An increase in grade is visible in Unit 6 as the distance from the contact increases.

Figure 14‑7: Contact Plot Unit 6 and Unit 7

14.5 Estimation Methodology

The block model was estimated using the lithologic boundaries of the Duluth Complex as the basis for an estimation domain. Units 1, 3, 5, 6, 7, the Magenta Zone, and Virginia Formation were all estimated using only samples that resided inside of the defined boundaries. See Figure 14‑1. This was done to prevent the smearing of higher grades from the assayed mineralized zones into areas of limited mineralization that were not assayed in the older U.S. Steel drilling campaigns.

14.5.1 Capping

Grade capping assigns statistically high outliers a maximum value in order to arrive at a better estimate of the true mean for the metal being estimated. The cap values were determined by examining Tukey Box Plots (Supplemental Information, 2018) and the sample distribution on log scale cumulative frequency plots (“CFP”) of the assay data.

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Tukey Box Plots divide the ordered values of the data into four equal parts by defining Inter Quartile Range (“IQR”); the median, and 25th and 75th percentiles. The median is defined by a horizontal line within a box that spans the IQR and contains approximately 50% of the data. The mean is represented by a large black circle. The fence is defined here as the central box (IQR) extended by 1.5 times the length of the box towards the maximum and the minimum. The upper and lower whiskers are then drawn from each end of the box to the fence position. Figure 14‑8 is an example of a Tukey Box plot of Unit 1.

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Figure 14‑8: Tukey Box Plots for Unit 1

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Samples outside of the fence are assumed to be Outliers and those that are three times the central box length from the upper or lower quartile boundaries are considered highly anomalous and are called Far Outliers. Table 14‑11 summarizes the capping values established for metals within each domain.

Table 14‑11: Summary of Capped Values for Each Metal

Domain	Cu (%)	Ni (%)	Pt (ppb)	Pd (ppb)	Au (ppb)	Ag (ppm)	Co (ppm)	S (%)
1	2.5	---	---	2250	---	---	330	---
3	1.8	0.4	700	2500	500	3.9	150	8
5	1.6	0.15	600	---	---	3.3	130	4.6
6	1.6	0.15	600	---	---	3.3	130	4.6
7	0.4	0.14	251	305	160	2.8	---	---
20	0.7	0.17	82	400	---	---	160	8.8
2000	---	0.3	900	---	600	8	148	---

14.5.2 Composite Study

HRC completed a composite study comparing the population variance and average grades, see Figure 14‑9. A composite length of 10-ft down-hole was selected for estimation as it is larger in length than the longest sample intervals; long enough to provide a variance reduction relative to using raw assay data, and still short enough to allow the estimate to show local variability of grade consistent with the sample distribution of the deposit. The composite statistics are summarized in Table 14‑12 through Table 14‑19.

Figure 14‑9: Copper Composite Study

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Table 14‑12: Copper Capped Composite Descriptive Statistics

Copper Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		%	%	%	%
1	12,135	0.00	1.57	0.22	0.21	0.96
3	6,275	0.00	1.62	0.06	0.09	1.69
5	2,248	0.00	1.16	0.04	0.08	1.80
6	885	0.00	1.44	0.04	0.09	2.29
7	500	0.00	0.33	0.03	0.04	1.49
20	877	0.00	0.70	0.04	0.08	2.08
2000	1,349	0.00	1.46	0.22	0.20	0.89

Table 14‑13: Nickel Capped Composite Descriptive Statistics

Nickel Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		%	%	%	%
1	12,135	0.00	0.63	0.07	0.05	0.75
3	6,275	0.00	0.26	0.03	0.02	0.76
5	2,248	0.00	0.15	0.03	0.01	0.55
6	885	0.00	0.15	0.03	0.01	0.41
7	500	0.01	0.13	0.04	0.02	0.54
20	877	0.00	0.17	0.02	0.02	1.25
2000	1,349	0.00	0.22	0.06	0.03	0.55

Table 14‑14: Platinum Capped Composite Descriptive Statistics

Platinum Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		ppb	ppb	ppb	ppb
1	12,135	0.0	876.1	47.2	59.8	1.3
3	6,275	0.0	479.3	21.1	34.4	1.6
5	2,248	0.0	525.0	14.8	28.4	1.9
6	885	0.0	537.6	20.8	45.1	2.2
7	500	0.0	248.6	16.1	27.7	1.7
20	877	0.0	82.0	6.5	11.3	1.7
2000	1,349	2.5	595.5	89.3	86.1	1.0

Table 14‑15: Palladium Capped Composite Descriptive Statistics

Palladium Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		ppb	ppb	ppb	ppb
1	12,135	0.0	2250.0	178.5	230.7	1.3
3	6,275	0.0	2228.3	61.7	125.5	2.0
5	2,248	0.0	1568.0	30.0	83.8	2.8
6	885	0.0	2683.7	50.7	171.7	3.4
7	500	0.0	305.0	24.7	42.5	1.7
20	877	0.0	395.4	18.7	49.9	2.7
2000	1,349	0.5	1964.4	236.6	247.7	1.0

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Table 14‑16: Gold Capped Composite Descriptive Statistics

Gold Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		ppb	ppb	ppb	ppb
1	12,135	0.0	916.0	25.0	35.0	1.4
3	2,248	0.0	381.4	7.6	16.5	2.2
5	2,240	0.5	381.4	7.6	16.5	2.2
6	885	0.0	292.9	8.6	18.7	2.2
7	500	0.0	145.4	6.7	14.3	2.1
20	877	0.0	119.4	4.9	8.2	1.7
2000	1,349	0.5	571.5	41.1	41.2	1.0

Table 14‑17: Silver Capped Composite Descriptive Statistics

Silver Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		ppm	ppm	ppm	ppm
1	12,135	0.0	16.5	0.8	0.8	1.0
3	6,275	0.0	3.9	0.3	0.3	1.0
5	2,248	0.0	3.3	0.3	0.2	0.8
6	885	0.0	3.3	0.3	0.3	0.8
7	500	0.0	2.0	0.3	0.2	0.5
20	877	0.1	3.2	0.4	0.4	0.9
2000	1,349	0.1	5.5	0.8	0.7	0.9

Table 14‑18: Cobalt Capped Composite Descriptive Statistics

Cobalt Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		ppb	ppb	ppb	ppb
1	12,135	0.0	309.3	67.0	26.7	0.4
3	6,275	0.0	150.0	51.7	17.7	0.3
5	2,248	0.0	130.0	49.0	11.3	0.2
6	885	0.0	127.5	60.0	13.5	0.2
7	500	0.0	158.6	68.8	28.1	0.4
20	877	9.3	160.0	31.6	21.2	0.7
2000	1,349	1.8	132.5	64.8	16.2	0.3

Table 14‑19: Sulfur Capped Composite Descriptive Statistics

Sulfur Capped and Composited Descriptive Statistics
Domain	Number	Minimum	Maximum	Mean	Std. Dev.	COV
		%	%	%	%
1	12,135	0.00	15.97	0.64	0.71	1.13
3	6,275	0.00	6.03	0.18	0.44	2.45
5	2,248	0.01	3.16	0.13	0.27	2.09
6	885	0.00	1.79	0.06	0.13	2.18
7	500	0.01	1.56	0.05	0.12	2.15
20	877	0.03	8.80	1.65	1.59	0.96
2000	1,349	0.01	2.49	0.36	0.32	0.88

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14.5.3 Variograms

HRC completed a variography analysis on the copper composites in order to evaluate the variography presented in the Updated Technical Report on the NorthMet Deposit dated January 13, 2013. HRC’s analysis of the copper variograms agreed with the structure, weights, and ranges of the variography analysis from the previous report. As such, HRC chose to utilize the parameters as previously stated. Table 14‑20 through Table 14‑22 summarize the variogram parameters utilized in estimation process.

Table 14‑20: Unit Variogram Parameters

Domain	Component	Increment	Cumulative	Rotation	Angle 1	Angle 2	Angle 3	Range 1	Range 2	Range 3
Unit 1 – Au	Nugget C0	0.036	0.036
Code 1	Exponential C1	0.748	0.784	ZYZ	-82.94	-72	45	14.3	60.8	3.4
	Exponential C2	0.216	1	ZYZ	-101.9	-53	11	108.7	466.1	560.8
Unit 1 – Co	Nugget C0	0.044	0.044
Code 1	Exponential C1	0.697	0.741	ZYZ	-99.94	58	4	105.9	221.1	24
	Exponential C2	0.259	1	ZYZ	-135.9	23	93	18	630.2	773.2
Unit 1 – Cu	Nugget C0	0.005	0.005
Code 1	Exponential C1	0.605	0.61	ZYZ	-85.94	-75	-4	26.1	74.9	7.9
	Exponential C2	0.39	1	ZYZ	-202.9	72	36	76.1	611.7	473.7
Unit 1 – Ni	Nugget C0	0.006	0.006
Code 1	Exponential C1	0.6	0.606	ZYZ	-41.94	21	42	58.3	11	33.3
	Exponential C2	0.394	1	ZYZ	-84.94	-46	-5	67.4	488.4	369.3
Unit 1 – Pd	Nugget C0	0.008	0.008
Code 1	Exponential C1	0.671	0.679	ZYZ	-52.94	15	-16	8.2	44.6	22.3
	Exponential C2	0.321	1	ZYZ	-110.9	-51	12	103.9	699.9	441.8
Unit 1 – Pt	Nugget C0	0.014	0.014
Code 1	Exponential C1	0.745	0.759	ZYZ	-108.9	21	21	6.5	33.4	24.1
	Exponential C2	0.241	1	ZYZ	-150.9	-71	31	108.3	494.6	895
Unit 1 – S	Nugget C0	0.015	0.015
Code 1	Exponential C1	0.558	0.573	ZYZ	-92.94	-56	9	19.4	157.1	8.8
	Exponential C2	0.427	1	ZYZ	-100.9	52	51	162.3	357.3	56.2

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Table 14‑21: Units 20 and 3, 5, 6, and 7 Variogram Parameters

Domain	Component	Increment	Cumulative	Rotation	Angle 1	Angle 2	Angle 3	Range 1	Range 2	Range 3
Unit 20 – Au	Nugget C0	0.368	0.368
Code 20	Spherical C1	0.435	0.803	ZYZ	-74.94	90	26	66.6	85.5	6.2
	Spherical C2	0.197	1	ZYZ	-55.94	-12	62	143.8	79.1	546.8
Unit 20 – Co	Nugget C0	0.398	0.398
Code 20	Spherical C1	0.279	0.677	ZYZ	-124.9	-62	81	48.3	215.9	11.4
	Spherical C2	0.323	1	ZYZ	-106.9	50	33	457	1,859.60	223.2
Unit 20 - Cu	Nugget C0	0.45	0.45
Code 20	Spherical C1	0.381	0.831	ZYZ	-94.94	87	-49	163.5	152.2	9
	Spherical C2	0.169	1	ZYZ	-60.94	-5	-54	155.5	500	1,200
Unit 20 – Ni	Nugget C0	0.406	0.406
Code 20	Spherical C1	0.34	0.746	ZYZ	-80.94	90	3	182.4	67.1	7.9
	Spherical C2	0.254	1	ZYZ	-83.94	11	9	78.3	117.5	1,190.40
Unit 20 – Pd	Nugget C0	0.571	0.571
Code 20	Spherical C1	0.198	0.769	ZYZ	-68.94	61	-55	44.1	140.4	163.5
	Spherical C2	0.231	1	ZYZ	-14.94	0	-24	5.4	50.9	609
Unit 20 – Pt	Nugget C0	0.434	0.434
Code 20	Spherical C1	0.402	0.836	ZYZ	-47.94	89	-47	81.3	52.1	4.9
	Spherical C2	0.164	1	ZYZ	-39.94	3	82	179.3	76.5	759.2
Unit 20 – S	Nugget C0	0.227	0.227
Code 20	Spherical C1	0.389	0.616	ZYZ	-150.9	28	3	28.4	60.8	138.8
	Spherical C2	0.384	1	ZYZ	-48.94	0	13	47.9	105.4	1,410.50
Unit 3, 4, 5, 6, 7 – Au	Nugget C0	0.3	0.3
Codes 3,4,5,6,7	Exponential C1	0.7	1	ZYZ	5.06	-22	18	210.6	78.5	20.2
Unit 3, 4, 5, 6, 7 – Co	Nugget C0	0.152	0.152
Codes 3,4,5,6,7	Exponential C1	0.848	1	ZYZ	-5.94	0	7	101.9	17.2	1321.8
Unit 3, 4, 5, 6, 7 – Cu	Nugget C0	0.006	0.006
Codes 3,4,5,6,7	Exponential C1	0.994	1	ZYZ	69.06	20	-55	410	29.7	21
Unit 3, 4, 5, 6, 7 – Ni	Nugget C0	0.142	0.142
Codes 3,4,5,6,7	Exponential C1	0.858	1	ZYZ	12.06	-13	-11	318.9	19.4	58.2
Unit 3, 4, 5, 6, 7 – Pd	Nugget C0	0.4	0.4
Codes 3,4,5,6,7	Exponential C1	0.6	1	ZYZ	-47.94	25	31	216.2	66.1	27.7
Unit 3, 4, 5, 6, 7 – Pt	Nugget C0	0.133	0.133
Codes 3,4,5,6,7	Exponential C1	0.867	1	ZYZ	-11.94	37	-14	133.4	87.8	9.8
Unit 3, 4, 5, 6, 7 – S	Nugget C0	0.011	0.011
Codes 3,4,5,6,7	Exponential C1	0.989	1	ZYZ	79.06	18	-55	176.4	56.9	28.2

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Table 14‑22: Magenta Zone Variogram Parameters

Domain	Component	Increment	Cumulative	Rotation	Angle 1	Angle 2	Angle 3	Range 1	Range 2	Range 3
Magenta Zone – Au	Nugget C0	0.004	0.004
Code 2000	Exponential C1	0.796	0.8	ZYZ	-47.94	41	-57	34.7	77.2	13.1
	Exponential C2	0.2	1	ZYZ	-102.9	-69	3	48.5	1609.1	469.9
Magenta Zone – Co	Nugget C0	0.003	0.003
Code 2000	Exponential C1	0.695	0.698	ZYZ	-68.94	83	-14	16.6	91.5	8.6
	Exponential C2	0.302	1	ZYZ	-91.94	35	48	1415.2	297.2	134.7
Magenta Zone – Cu	Nugget C0	0.004	0.004
Code 2000	Exponential C1	0.81	0.814	ZYZ	-10.94	20	-54	170.1	67.4	19.9
	Exponential C2	0.186	1	ZYZ	-87.94	-53	-4	26.4	1004.3	911.1
Magenta Zone – Ni	Nugget C0	0.006	0.006
Code 2000	Exponential C1	0.816	0.822	ZYZ	-12.96	27	-63	156.4	89	19
	Exponential C2	0.178	1	ZYZ	-88.9	-53	-3	28.7	1396.2	424.5
Magenta Zone – Pd	Nugget C0	0.003	0.003
Code 2000	Exponential C1	0.744	0.747	ZYZ	-63.94	57	11	35.5	79.1	11.5
	Exponential C2	0.253	1	ZYZ	-5.94	-88	-25	60.2	272.8	1068.1
Magenta Zone - Pt	Nugget C0	0.004	0.004
Code 2000	Exponential C1	0.727	0.731	ZYZ	-59.94	59	8	28.3	103.7	1.9
	Exponential C2	0.269	1	ZYZ	-105.9	-74	2	33.1	937.5	246.1
Magenta Zone – S	Nugget C0	0.082	0.082
Code 2000	Exponential C1	0.723	0.805	ZYZ	-4.94	21	-97	149.2	87.1	19
	Exponential C2	0.195	1	ZYZ	-88.94	-68	-2	26.5	551.9	332.2

14.5.4 Estimation Strategy

Because of the subtle changes in direction of the mineralized contacts, the estimation method selected to model the mineralization changes is an Ordinary Kriging (OK) using dynamic search ellipses for Domains 1, 3, 5, 6 and 7, as shown in Figure 14‑10. With this method, the orientation of the search and variogram ellipses changes on a block by block basis utilizing wireframe interpretations of each of the unit boundaries. In this model, five separate surfaces were created and utilized to model the structural fabric of the Duluth Complex in association with the mineral resource. These wireframes were created based on surface geology maps and drill-hole intercepts. The Magenta Zone was estimated using a single search ellipse oriented in the direction of the maximum geologic continuity.

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Figure 14‑10: North – South Section Looking East Displaying the Dynamic Search Ellipses

The grades were estimated from 10-foot down-hole composites using OK. Composites were coded according to their domain. Each metal was estimated using the variogram parameters outlined in Table 14‑20 through Table 14‑22. Table 14‑23 summarizes the search parameters used in the estimation of mineral resources.

Table 14‑23: Search Volume Parameters for all Domains

	Ellipsoid dimension (in ft.)			Number of Samples Used
	X	Y	Z	Min	Max	Max per hole	Comment
Pass 1	300	170	40	6	15	5	Minimum of two holes required
Pass 2	600	340	80	6	15	5	Minimum of two holes required
Pass 3	900	500	115	2	15	5

14.5.5 Mineral Resource Classification

HRC used the anisotropic distance to the nearest composite of each block to classify mineral resources into measured, indicated and inferred. Table 14‑24 summarizes the distances and number of samples used for the mineral classification criteria.

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Table 14‑24: Mineral Resource Classification Criteria

Classification	X	Y	Z	Samples
	Feet	Feet	feet	Min	Max	Max per Hole
Measured	≤300	≤170	≤40	6	15	4
Indicated	≥300 and ≤600	≥170 and ≤340	≥40 and ≤80	6	15	4
Inferred	≥600 and ≤900	≥340 and ≤510	≥80 and ≤120	2	15	2

14.5.6 Model Validation

Overall, HRC utilized several methods to validate the results of the estimation method. The combined evidence from these validation methods verifies the OK estimation model results.

14.5.6.1 Comparison with Inverse Distance and Nearest Neighbor Models

Inverse Distance (ID) and Nearest Neighbor (NN) models were run to serve as comparison with the estimated results from the OK method. Descriptive statistics for the OK method along with those for the ID, NN, and drill-hole composites are shown in Table 14‑25 through Table 14‑32 “N” signifies number of samples in the tables.

Table 14‑25: Copper Model Statistics

Cu (%) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.00	1.62	0.14	0.18	1.30
OK	595,727	0.00	1.29	0.10	0.12	1.20
ID	595,727	0.00	1.32	0.10	0.13	1.26
NN	595,727	0.00	1.62	0.10	0.15	1.51

Table 14‑26: Nickel Model Statistics

Ni (%) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.00	0.63	0.05	0.04	0.86
OK	595,727	0.00	0.30	0.04	0.03	0.67
ID	595,727	0.00	0.40	0.04	0.03	0.70
NN	595,727	0.00	0.63	0.04	0.03	0.86

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Table 14‑27: Platinum Model Statistics

Pt (ppb) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.0	876.1	36.7	55.2	1.5
OK	595,727	0.0	705.1	27.7	38.4	1.4
ID	595,727	0.0	799.4	27.6	40.0	1.5
NN	595,727	0.0	876.1	27.4	51.5	1.9

Table 14‑28: Palladium Model Statistics

Pd (ppb) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.0	2683.7	124.1	201.9	1.6
OK	595,727	0.0	2195.3	86.5	138.0	1.6
ID	595,727	0.0	2176.7	85.8	144.7	1.7
NN	595,727	0.0	2683.7	86.4	181.7	2.1

Table 14‑29: Gold Model Statistics

Au (ppb) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.0	916.0	19.0	30.8	1.6
OK	595,727	0.0	324.0	14.3	19.2	1.3
ID	595,727	0.0	530.4	14.2	20.1	1.4
NN	595,727	0.0	916.0	14.4	28.0	1.9

Table 14‑30: Silver Model Statistics

Ag (ppm) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.0	16.5	0.6	0.7	1.1
OK	595,727	0.0	7.4	0.5	0.4	0.8
ID	595,727	0.0	12.7	0.5	0.4	0.9
NN	595,727	0.0	16.5	0.5	0.5	1.1

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Table 14‑31: Cobalt Model Statistics

Co (ppb) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.0	309.3	59.8	24.2	0.4
OK	595,727	0.0	214.1	55.1	17.7	0.3
ID	595,727	0.0	269.0	55.0	18.3	0.3
NN	595,727	0.0	309.3	55.2	22.4	0.4

Table 14‑32: Sulfur Model Statistics

S (%) Grade Model Comparisons: All Domains
Model	N	Min	Max	Mean	Stan. Dev.	COV
Composites	24,269	0.00	15.97	0.46	0.72	1.57
OK	595,727	0.00	8.25	0.47	0.68	1.46
ID	595,727	0.00	8.25	0.47	0.71	1.53
NN	595,727	0.00	15.97	0.46	0.84	1.80

The overall reduction of the maximum, mean, standard deviation, and coefficient of variation (COV) within the OK and ID models represents an appropriate amount of smoothing to account for the point to block volume variance relationship. This is confirmed in Figure 14‑11, comparing the Unit 1 copper cumulative frequency plots of each of the models and drill-hole composites.

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Figure 14‑11: Model Comparison Cumulative Frequency Plot (NN red, ID blue, Composites Black, OK Green)

14.5.6.2 Swath Plots

Swath plots (Supplemental Information, 2018) were generated to compare average estimated grade from the OK method to the two validation model methods (ID and NN). The results from the OK model, plus those for the validation ID model method are compared using the swath plot to the distribution derived from the NN model. Figure 14‑12 shows average copper grade within Unit 1 along the rotated easting.

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Figure 14‑12: Domain 1 Copper Swath Plot Along Rotated Easting

On a local scale, the nearest neighbor model does not provide a reliable estimate of grade, but on a much larger scale, it represents an unbiased estimation of the grade distribution based on the total dataset. Therefore, if the OK model is unbiased, the grade trends may show local fluctuations on a swath plot, but the overall trend should be similar to the distribution of grade from the nearest neighbor.

Overall, there is good correlation between the grade models, although deviations occur near the edges of the deposit and in areas where the density of drilling is less and material is classified as Inferred resources.

14.5.6.3 Evaluation of Non-Sampled Intervals

U.S. Steel did not assay a number of intervals that did not visually indicate mineralization, particularly in the deeper holes in the southeast area of the deposit. HRC estimated the resources by both replacing the non-sampled intervals with zeros and by ignoring the intercepts to understand the effect on the estimate.  Additionally, hard boundaries were used in the estimate to prevent the smearing of higher grades from the assayed mineralized zones into areas of limited mineralization that were not sampled in the older U.S. Steel drilling campaigns.

Within the optimized pit shell used to determine reasonable prospects for eventual economic extraction there is only a 0.18% difference in material above cutoff between the two different methods for handling the non-sampled intervals. The difference between the models is considered to be within the margin of error of the estimate. HRC selected the model that ignored the non-sampled intervals for the reporting of mineral resources.

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14.5.6.4 Sectional Inspection

Bench plans, cross-sections, and long sections comparing modeled grades to the 10-ft composites were evaluated. Sections displaying copper estimated grades and composite grades are shown in Figure 14‑13 through Figure 14‑15. The figure shows good agreement between modeled grades and the composite grades. In addition, the modeled blocks display continuity of grades along strike and down dip.

Figure 14‑13: Copper Cross Section Along Rotated Easting

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Figure 14‑14: Copper Long Section Along Rotated Northing

Figure 14‑15: Copper Plan Section

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14.6 Mineral Resources

The mineral resources for the NorthMet Project are calculated at 649.3 million tons measured and indicated and 508.9million tons inferred.  The mineral resources and grades are summarized in Table 14‑35 and are reported inclusive of mineral reserves.

14.6.1 Net Smelter Return (NSR) and Cutoff

For each block in the mineral resource model, the net smelter return (NSR) was calculated utilizing the same formulas utilized by IMC in calculating the mineral reserves (see Section 15.1.3).  The NSR calculation takes into account the estimated metal recovery curves for each metal, the treatment charges, payment terms, deducts, penalties, shipping charges and royalties.  HRC reviewed the smelter terms and found them to be within industry norms.  The NSR formula utilized the metal prices as presented in Table 14‑33 and included royalty deducts of 5%if the NSR was $35.00/t or over, 4% if the NSR was under $35.00/t but $30.00 over and 3% if the NSR was under $30.00/t. Table 14‑33 also shows the estimated average metal recoveries for the resources which are calculated from the recovery curves presented in Section 13.6.

Table 14‑33: Resource Metal Prices and Estimated Recoveries

Metal	Price	Recovery
Copper ($/lb)	3.30	91.3
Nickel ($/lb)	8.50	61.4
Cobalt ($/lb)	13.28	30.0
Palladium ($/oz)	734	74.2
Platinum ($/oz)	1286	78.6
Gold ($/oz)	1263	59.9
Silver ($/oz)	19.06	56.5

Table 14‑34 summarizes the operating costs used to develop the $7.35/t NSR cutoff used as the base case for reporting of mineral resources. The estimated operating costs were provided by PolyMet and the cutoff reflects the potential economic, marketing, and other issues relevant to an open pit mining scenario based on a milling recovery process producing copper and nickel concentrates.  HRC has reviewed the cost estimates and finds them to be within industry averages and adequate for reporting of the mineral resources.

Table 14‑34: Estimated Process Operating Costs

Department	Cost
Process Cost ($/t)	6.50
Property G&A Costs ($/t)	0.50
Waste Water Treatment Costs ($/t)	0.35
Total Cost ($/t)	7.35

14.6.2 Test for Reasonable Prospect for Eventual Economic Extraction

In order to identify the mineralization that meets the test for reasonable prospects of eventual economic extraction, and thus be classified as mineral resources, a Lerchs-Grossman pit shell was generated. The optimization parameters utilized the NSR values calculated in each block based on the metal prices presented in Table 14‑33 and the operating costs presented in Table 14‑34. Mining costs for the optimization were estimated at $1.15/t mined at surface and for every 50 feet of depth the mining costs increased $0.02/t.  Pit slope angles were restricted to 48 degrees.

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The mineral resource estimate presented in Table 14‑35 is inclusive of the mineral reserves. The resource has been limited to the material that resides above the optimized pit shell.  All mineralization below the optimized pit shell has been excluded from any resource classification and is not considered to be part of the mineral resource.

14.6.3 Resource Statement

The mineral resource estimate for the NorthMet Project is summarized in Table 14‑35.  This mineral resource estimate includes all drill data obtained as of January 31, 2016 and has been independently verified by HRC.  Mineral resources are not mineral reserves and may be materially affected by environmental, permitting, legal, socio-economic, marketing, political, or other factors. The measured and indicated mineral resources are inclusive of the mineral reserves.  Inferred mineral resources are, by definition, always additional to mineral reserves.

Table 14‑35: Mineral Resource Statement for the NorthMet Project Inclusive of Mineral Reserves, Hard Rock Consulting, LLC, January 1, 2018

	Volume (M ft3)	Density (st/ft3)	Tonnage (M st)	Cu (%)	Ni (%)	S (%)	Pt (ppb)	Pd (ppb)	Au (ppb)	Co (ppm)	Ag (ppm)	NSR (US$/t)	Cu-Eq (%)
Measured	2,564.9	0.092	237.2	0.270	0.080	0.66	69	241	35	72	0.97	19.67	0.541
Indicated	4,468.5	0.092	412.2	0.230	0.070	0.58	63	210	32	70	0.87	16.95	0.470
M+I	7,033.4	0.092	649.3	0.245	0.074	0.61	65	221	33	71	0.91	17.94	0.496
Inferred	5,545.5	0.092	508.9	0.240	0.070	0.54	72	234	37	66	0.93	17.66	0.489

*Notes:

(1) 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 estimated will be converted into mineral reserves.

(2)   Mineral resource grades are reported undiluted.

(3) All resources are stated above a $7.35NSR cutoff.

(4) Cutoff is based on assumed processing and G&A costs of US $7.35 per ton. Metal Prices and metallurgical recoveries are presented in Table 14‑33.

(5) Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbers may not add due to rounding.

(6) CuEq (copper equivalent grade) is based on the mill recovery to concentrates and metal prices (Table 14‑33).

  (7)  Copper Equivalent (CuEq) = ((Cu head grade x recovery x Cu Price) + (Ni head grade x recovery x Ni Price) + (Pt head grade x recovery

          x Pt Price) + (Pd head grade x recovery x Pd Price) + (Au head grade x recovery x Au Price) + (Co head grade x recovery x Co Price) +

          (Ag head grade x recovery x Ag Price)) / (Cu recovery x Cu Price).

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15 MINERAL RESERVE ESTIMATES

The pits were evaluated according to the updated Measured and Indicated Resources and demonstrated to be economically viable; therefore, Measured and Indicated Mineral Resources within the final pit design have been converted to Proven and Probable Reserves. The mineral reserves use the terminology, definitions and guidelines given in the CIM Standards on Mineral Resource and Mineral Reserves (May 2014).  All inferred material was classified as waste and scheduled to the appropriate waste stockpile.

15.1 Calculation Parameters

The pit designs used in this study were compared with pit optimizations run on the updated operating costs and metal prices used in this report and were found to be well within the optimized shells.  The optimized shells were only used to confirm the validity of the pit designs and to report the minable resource.

15.1.1 Pit Slopes

The pit slopes for the pit and internal phase designs followed the recommendations from the June 2006 Golder NorthMet Open Pit Rock Slope Design Report which was reviewed by IMC, and the recommended inter-ramp and overall pit wall recommendations have been incorporated into the designs.

The Golder report indicated inter-ramp angles of 51.4 degrees for all sectors, except one, were possible.  That one sector utilized an inter-ramp angle of 55.1 degrees and was achieved with a bench face angle of 70 degrees versus the other sectors’ 65-degree face angle. The area impacted by the increased bench face angle was minimal.  To simplify the pit design, all areas were designed with a bench face angle of 65 degrees.

The Golder report also included the following design recommendations which are incorporated into the pit wall slopes:

· In cases where the vertical lift is less than 400 ft between haul ramps, a 33.2 ft catch bench is included every 100 ft of vertical lift to achieve an inter-ramp angle of 51.4 degrees.

· In cases where the vertical lift exceeds 400 ft between haul ramps, an additional 27.2 ft is added to one of the normal 33.2 ft catch benches to achieve an overall slope angle of 49.1 degrees.

15.1.2 Dilution and Mining Losses

The mineral resource estimate for NorthMet is considered to be internally diluted by compositing.  HRC also calculated an external diluted grade for all of the grade elements; these diluted grades were used by IMC for the mineral reserve calculation.  To