formation capital corporation, u.s. ecobalt solutions … · suite 900 - 390 bay street, toronto...

SUITE 900 - 390 BAY STREET, TORONTO ONTARIO, CANADA M5H 2Y2 Telephone (1) (416) 362-5135 Fax (1) (416) 362 5763

FORMATION CAPITAL CORPORATION, U.S.

eCOBALT SOLUTIONS INC.

NI 43-101 F1 TECHNICAL REPORT

FEASIBILITY STUDY

FOR THE

IDAHO COBALT PROJECT

IDAHO, USA

Report Date: November 10, 2017

Effective Date: September 27, 2017

Report by

Barnard Foo, P.Eng., MBA

Charley Murahwi, M.Sc., P.Geo., FAusIMM

Christopher Jacobs, CEng, MIMMM

David Makepeace, M.Eng., P.Eng.

Richard Gowans, B.Sc., P.Eng.

Jane Spooner, M.Sc., P.Geo.

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Table of Contents

1.0 SUMMARY ................................................................................................................. 1 1.1 AUTHORIZATION AND PURPOSE ................................................................................. 1 1.2 PROPERTY DESCRIPTION AND OWNERSHIP ................................................................ 1 1.3 GEOLOGY AND MINERALIZATION .............................................................................. 1 1.4 STATUS OF EXPLORATION .......................................................................................... 2

1.5 MINERAL PROCESSING/METALLURGICAL TESTING ................................................... 2 1.6 MINERAL RESOURCE ESTIMATE ................................................................................ 3 1.7 MINERAL RESERVE ESTIMATE ................................................................................... 3 1.8 MINING METHODS ..................................................................................................... 4 1.9 PROCESSING ............................................................................................................... 5

1.9.1 Mill/Concentrator ......................................................................................................... 5 1.9.2 Cobalt Processing Facility (CPF) ................................................................................. 6

1.10 INFRASTRUCTURE ...................................................................................................... 6 1.10.1 Mill/Concentrator ......................................................................................................... 6 1.10.2 CPF Infrastructure ........................................................................................................ 7

1.11 MARKET STUDIES AND CONTRACTS .......................................................................... 7 1.11.1 Market Studies .............................................................................................................. 7 1.11.2 Contracts ....................................................................................................................... 8

1.12 ENVIRONMENT STUDIES, PERMITTING AND SOCIAL/COMMUNITY IMPACT ................ 8

1.13 CAPITAL AND OPERATING COSTS .............................................................................. 9 1.14 ECONOMIC ANALYSIS .............................................................................................. 10

1.14.1 Global Assumptions ................................................................................................... 10 1.14.2 Technical Assumptions .............................................................................................. 11 1.14.3 Discounted Cash Flow Evaluation ............................................................................. 15 1.14.4 Sensitivity ................................................................................................................... 15 1.14.5 Conclusion .................................................................................................................. 16

1.15 CONCLUSION AND RECOMMENDATIONS .................................................................. 16 1.15.1 Geology and Resources .............................................................................................. 16 1.15.2 Mining ........................................................................................................................ 17 1.15.3 Processing – Future Testwork .................................................................................... 18

2.0 INTRODUCTION ..................................................................................................... 20 2.1 AUTHORIZATION AND PURPOSE ............................................................................... 20 2.2 SOURCES OF INFORMATION ...................................................................................... 20

2.3 SCOPE OF PERSONAL INSPECTION ............................................................................ 21

2.4 LIST OF ABBREVIATIONS .......................................................................................... 21

3.0 RELIANCE ON OTHER EXPERTS ...................................................................... 25

4.0 PROPERTY DESCRIPTION AND LOCATION ................................................. 26 4.1 LOCATION AND GENERAL DESCRIPTION .................................................................. 26 4.2 LAND TENURE ......................................................................................................... 27 4.3 TENURE RIGHTS AND RISK FACTORS ....................................................................... 30

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY .......................................................................................... 31 5.1 ACCESSIBILITY ........................................................................................................ 31 5.2 CLIMATE .................................................................................................................. 32 5.3 LOCAL RESOURCES AND INFRASTRUCTURE ............................................................. 32 5.4 PHYSIOGRAPHY ........................................................................................................ 32

6.0 HISTORY .................................................................................................................. 34 6.1 DISCOVERY HISTORY ............................................................................................... 34 6.2 HISTORICAL STUDY AND EVALUATION WORK ........................................................ 34 6.3 HISTORICAL MINERAL RESOURCE ESTIMATES ........................................................ 37

6.3.1 1981 and 1997 ICP Mineral Resource Estimates ....................................................... 38 6.3.2 1998 ICP Mineral Resource Estimate ........................................................................ 38 6.3.3 MDA 2001 Resource Estimate ................................................................................... 38 6.3.4 MDA 2005 Resource Estimate ................................................................................... 39 6.3.5 MDA 2006 Resource Estimate ................................................................................... 39

6.4 PRODUCTION HISTORY ............................................................................................ 40

7.0 GEOLOGICAL SETTING AND MINERALIZATION ....................................... 41 7.1 OVERVIEW ............................................................................................................... 41

7.2 REGIONAL GEOLOGY ............................................................................................... 41 7.3 LOCAL GEOLOGY ..................................................................................................... 43

7.3.1 Lithology and Stratigraphy ......................................................................................... 44 7.3.2 Structural Geology of the Deposits ............................................................................ 44 7.3.3 Ram Deposit Stratigraphy .......................................................................................... 45 7.3.4 Sunshine Deposit Stratigraphy ................................................................................... 47 7.3.5 Alteration .................................................................................................................... 48

7.4 MINERALIZATION .................................................................................................... 49 7.4.1 Global Overview ........................................................................................................ 49 7.4.2 ICP Mineralization ..................................................................................................... 49

8.0 DEPOSIT TYPES ..................................................................................................... 51 8.1 PRE-2005 CONCEPTIONS .......................................................................................... 51 8.2 POST-2005 CONCEPTIONS ........................................................................................ 51

9.0 EXPLORATION ....................................................................................................... 52 9.1 PROGRAMS ............................................................................................................... 52

9.1.1 1995-1996 Campaign ................................................................................................. 52 9.1.2 1997 Campaign........................................................................................................... 52 9.1.3 1998-2001 Campaign ................................................................................................. 52 9.1.4 2002-2006 Campaign ................................................................................................. 53 9.1.5 2007-2016 Campaign ................................................................................................. 53

9.2 EXPLORATION RESULTS ........................................................................................... 53

10.0 DRILLING ................................................................................................................ 54 10.1 DRILLING CAMPAIGNS ............................................................................................. 54 10.2 FCC DRILLING PROCEDURES ................................................................................... 56

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10.3 MICON OBSERVATIONS DURING SITE VISIT/COMMENTS ......................................... 56

10.4 DRILLING RESULTS .................................................................................................. 57 10.5 MICON COMMENTS .................................................................................................. 58

11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ............................... 60 11.1 SAMPLE PREPARATION ............................................................................................ 60

11.1.1 Sample Preparation at Site ......................................................................................... 60 11.1.2 Laboratory Sample Preparation .................................................................................. 60

11.2 ANALYSES ............................................................................................................... 60

11.3 SECURITY................................................................................................................. 61 11.4 QUALITY CONTROL/ASSURANCE (QA/QC) ............................................................. 62

11.4.1 MDA Verification ...................................................................................................... 62 11.4.2 Micon Verification ..................................................................................................... 63

11.5 SUMMARY STATEMENT/COMMENTS ........................................................................ 67

12.0 DATA VERIFICATION .......................................................................................... 68 12.1 SITE VISITS .............................................................................................................. 68

12.1.1 Discussions on Geological Attributes ........................................................................ 68 12.1.2 Discussions on Mine Planning Parameters ................................................................. 68 12.1.3 Field Examination of Out Crops................................................................................. 69 12.1.4 Examination of Drill Cores ........................................................................................ 69 12.1.5 Data Collection Techniques/Sampling ....................................................................... 69 12.1.6 Down-hole Surveys .................................................................................................... 69 12.1.7 Analysis of QA/QC Monitoring Charts ...................................................................... 70 12.1.8 Specific Gravity .......................................................................................................... 70

12.2 REVIEW OF MDA DATA VERIFICATION ................................................................... 71 12.2.1 Database Audit for the 2006 Resource ....................................................................... 71 12.2.2 Database Audit for the 2015 ....................................................................................... 71 12.2.3 QA/QC for the 2006/2015 Resources......................................................................... 71

12.3 DATABASE VALIDATION .......................................................................................... 71

12.4 DATA VERIFICATION CONCLUSIONS ........................................................................ 72

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................... 73 13.1 METALLURGICAL TESTWORK PROGRAMS ................................................................ 73 13.2 METALLURGICAL SAMPLES ..................................................................................... 74 13.3 MINERALOGY .......................................................................................................... 75

13.4 COMMINUTION ......................................................................................................... 76 13.5 FLOTATION .............................................................................................................. 76

13.5.1 Bulk Concentrate Flotation ........................................................................................ 77 13.5.2 Copper Scalping Flotation .......................................................................................... 79 13.5.3 Concentrate Characteristics ........................................................................................ 81 13.5.4 Concentrator Flotation Recoveries ............................................................................. 82

13.6 SOLID-LIQUID SEPARATION ..................................................................................... 84 13.6.1 Tailings ....................................................................................................................... 84 13.6.2 Concentrate ................................................................................................................. 85

13.7 HYDROMETALLURGICAL PROCESS ........................................................................... 85 13.7.1 Leaching Circuit ......................................................................................................... 86 13.7.2 Leach Residue Thickening and Filtration................................................................... 90

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13.7.3 Iron Removal (Fe Precipitation and Thickening) ....................................................... 91 13.7.4 Cobalt Precipitation and Re-dissolution ..................................................................... 92 13.7.5 Cobalt Solvent Extraction .......................................................................................... 94 13.7.6 Cobalt Sulphate Crystallization .................................................................................. 94 13.7.7 Copper Solvent Extraction ......................................................................................... 96 13.7.8 Copper Sulphate Crystallization ................................................................................. 97 13.7.10 Gold Recovery Circuit ................................................................................................ 97

13.8 RECOMMENDATIONS FOR FUTURE TESTWORK ......................................................... 98 13.8.1 Copper Flotation ......................................................................................................... 98 13.8.2 Cobalt Solvent Extraction .......................................................................................... 98 13.8.3 Copper Solvent Extraction ......................................................................................... 98 13.8.4 Crystallization ............................................................................................................ 98 13.8.5 Gold Recovery Circuit ................................................................................................ 99 13.8.6 CPF Pilot Plant ........................................................................................................... 99 13.8.7 Process Modelling and Simulation ............................................................................. 99 13.8.8 HAZOP Studies .......................................................................................................... 99

14.0 MINERAL RESOURCE ESTIMATES ................................................................ 100 14.1 DATABASE DESCRIPTION ....................................................................................... 100 14.2 OVERVIEW OF MDA’S ESTIMATION METHODOLOGY ............................................ 100 14.3 GLOBAL/GENERAL STATISTICS .............................................................................. 102 14.4 GEOLOGIC AND DOMAIN MODEL ........................................................................... 104

14.4.1 MDA Modelling ....................................................................................................... 104 14.4.2 Micon Review and Wireframing .............................................................................. 106

14.5 GRADE CAPPING, COMPOSITING AND DOMAIN STATISTICS ................................... 107

14.6 GEOSTATISTICS ...................................................................................................... 111 14.6.1 Density ..................................................................................................................... 114

14.7 ESTIMATION ........................................................................................................... 114 14.7.1 Block Model Definition ............................................................................................ 114 14.7.2 Estimation/Search Parameters .................................................................................. 115 14.7.3 Grade Interpolation ................................................................................................... 115 14.7.4 Block Grades Validation .......................................................................................... 117 14.7.5 Mineral Resource Parameters and Categorization.................................................... 120 14.7.6 Mineral Resource Statement..................................................................................... 121 14.7.7 Risks/Uncertainties ................................................................................................... 123

15.0 MINERAL RESERVE ESTIMATES ................................................................... 124 15.1 RESOURCE MODEL ................................................................................................. 124

15.2 CUT-OFF GRADE (COG) CRITERIA AND ESTIMATE ................................................ 124 15.3 STOPE OUTLINE ..................................................................................................... 125

15.4 DILUTION AND LOSSES .......................................................................................... 125 15.5 MINING RECOVERY................................................................................................ 126 15.6 MINERAL RESERVE ESTIMATE ............................................................................... 127

16.0 MINING METHODS ............................................................................................. 128 16.1 MINING METHODS ................................................................................................. 128 16.2 MINE DESIGN PARAMETERS .................................................................................. 129 16.3 GEOTECHNICAL CONSIDERATIONS ......................................................................... 129

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16.3.1 Principal Rock Type ................................................................................................. 129 16.3.2 Rock Quality Designation (RQD) ............................................................................ 131 16.3.3 Joint Data .................................................................................................................. 131 16.3.4 Rock Mass Rating .................................................................................................... 132

16.4 GROUND SUPPORT RECOMMENDATIONS ................................................................ 133 16.4.1 Background .............................................................................................................. 133 16.4.2 Underground Geotechnical Design Parameters Ram/Sunshine Deposits (Minefill,

2006) ........................................................................................................................ 133 16.4.3 Updated ‘Preliminary’ Ground Support Recommendations .................................... 134 16.4.4 Conclusion – Geotechnical Consideration ............................................................... 136

16.5 MINING CUT-OFF GRADE AND SPECIFICATIONS .................................................... 138 16.6 SELECTIVITY, DILUTION AND RECOVERY .............................................................. 139

16.6.1 Mining Selectivity .................................................................................................... 139 16.6.2 Dilution ..................................................................................................................... 139 16.6.3 Mining Recovery ...................................................................................................... 139

16.7 MINE DESIGN ......................................................................................................... 140 16.7.1 Underground Excavation Dimensions ...................................................................... 140 16.7.2 Mine Access ............................................................................................................. 140 16.7.3 Underground Mine Layout ....................................................................................... 141

16.8 MINE DEVELOPMENT AND PRODUCTION SCHEDULE .............................................. 142 16.8.1 Mine Development ................................................................................................... 143 16.8.2 Production Schedule ................................................................................................. 146

16.9 MANPOWER REQUIREMENTS ................................................................................. 149 16.10 EQUIPMENT SELECTION ......................................................................................... 152

16.11 UTILITIES, SERVICES FOR UNDERGROUND ............................................................. 152 16.11.1 Temporary Mine Area Building ............................................................................... 152 16.11.2 Explosive Storage ..................................................................................................... 153 16.11.3 Underground Communication System ..................................................................... 153

16.12 VENTILATION ......................................................................................................... 153 16.13 BACKFILL SYSTEM ................................................................................................. 154

16.13.1 Backfill Reticulation and Pumping System .............................................................. 155 16.13.2 Backfill Material Testing .......................................................................................... 156 16.13.3 Design Criteria ......................................................................................................... 158

16.14 MINE DEWATERING ............................................................................................... 159 16.15 COMPRESSED AIR .................................................................................................. 160 16.16 POWER REQUIREMENTS AND DISTRIBUTION .......................................................... 160

17.0 RECOVERY METHODS ...................................................................................... 161 17.1 MINE SITE PROCESS PLANT DESIGN ...................................................................... 161

17.1.1 Process Description .................................................................................................. 161 17.2 COBALT PROCESSING FACILITY (CPF) .................................................................. 163

17.2.1 Process Description .................................................................................................. 163 17.2.2 Acidulation ............................................................................................................... 166 17.2.3 Pressure Oxidation Acid Leaching ........................................................................... 166 17.2.4 Cu-SX and Crystallization ........................................................................................ 168 17.2.5 Leach Filtration ........................................................................................................ 169 17.2.6 Sulphur Flotation ...................................................................................................... 169 17.2.7 Gold Recovery .......................................................................................................... 169

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17.2.8 Secondary Belt Filter/Cyanide Destruction .............................................................. 170 17.2.9 Cu-Fe Removal......................................................................................................... 170 17.2.10 Cobalt Precipitation .................................................................................................. 171 17.2.11 Cobalt SX ................................................................................................................. 171 17.2.12 Crud Treatment......................................................................................................... 172 17.2.13 Cobalt Sulphate Crystallization ................................................................................ 172 17.2.14 Trace Metal Precipitation ......................................................................................... 173 17.2.15 Magnesium Sulphate Crystallization ........................................................................ 173

18.0 PROJECT INFRASTRUCTURE .......................................................................... 175 18.1 MINE AND MILL SITE ............................................................................................. 175

18.1.1 Site Layout ............................................................................................................... 175 18.1.2 Work Completed to Date .......................................................................................... 176 18.1.3 Mine Site Access Roads ........................................................................................... 177 18.1.4 Buildings .................................................................................................................. 178 18.1.5 Electrical Power Supply and Distribution ................................................................ 178 18.1.6 Surface Facilities Fire Protection ............................................................................. 179 18.1.7 Mine and Concentrator Communications ................................................................. 179 18.1.8 Water Supply, Treatment and Discharge .................................................................. 179 18.1.9 Tailings and Waste Rock Storage (TWSF) .............................................................. 181 18.1.10 Explosives Storage and Transport ............................................................................ 183

18.2 CPF INFRASTRUCTURE .......................................................................................... 183 18.2.1 CPF Site Access ....................................................................................................... 183 18.2.2 Process Plant Layout ................................................................................................ 184 18.2.3 Buildings .................................................................................................................. 184 18.2.4 Crystallizer Pads ....................................................................................................... 185 18.2.5 Administration Complex .......................................................................................... 186 18.2.6 Rail Spur Line and Loading Area ............................................................................. 186 18.2.7 Hydrometallurgical Facility Fire Protection ............................................................. 186 18.2.8 Power Supply and Distribution ................................................................................ 186 18.2.9 Process Control System ............................................................................................ 187 18.2.10 Communications ....................................................................................................... 187 18.2.11 Water Supply ............................................................................................................ 187 18.2.12 Waste Disposal ......................................................................................................... 188

19.0 MARKET STUDIES AND CONTRACTS ........................................................... 189 19.1 INTRODUCTION ...................................................................................................... 189 19.2 COBALT ................................................................................................................. 189

19.2.1 Cobalt Sulphate ........................................................................................................ 190 19.3 COPPER SULPHATE................................................................................................. 192 19.4 MAGNESIUM SULPHATE ......................................................................................... 192 19.5 GOLD ..................................................................................................................... 192

19.6 COPPER CONCENTRATE ......................................................................................... 193 19.7 PROJECTED REVENUE AND MARKET POSITION ...................................................... 193 19.8 CONTRACTS ........................................................................................................... 194

20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT ....................................................................................... 195

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20.1 ENVIRONMENTAL BASELINE STUDIES AND IMPACT ASSESSMENTS........................ 195 20.1.1 Mine and Mill ........................................................................................................... 195 20.1.2 CPF ........................................................................................................................... 201

20.2 SOCIAL COMMUNITY RELATIONS .......................................................................... 201 20.2.1 Mine and Mill ........................................................................................................... 201 20.2.2 CPF ........................................................................................................................... 201

20.3 PLAN OF OPERATIONS ............................................................................................ 201 20.3.1 Tailings and Waste Rock Storage Facility ............................................................... 202 20.3.2 Water Management .................................................................................................. 203 20.3.3 Reclamation – Closure ............................................................................................. 206 20.3.4 Closure Considerations ............................................................................................. 207

20.4 CPF OPERATIONS .................................................................................................. 207

20.5 PERMITS ................................................................................................................. 208 20.5.1 Mine/Mill ................................................................................................................. 208 20.5.2 CPF ........................................................................................................................... 209

21.0 CAPITAL AND OPERATING COSTS................................................................ 210 21.1 CAPITAL COST ESTIMATE ...................................................................................... 210

21.1.1 Mining Capital Cost ................................................................................................. 210 21.1.2 Mill/Concentrator and Infrastructure - Direct Capital Cost ...................................... 211 21.1.3 Indirect Capital Costs ............................................................................................... 212 21.1.4 Contingency – Mine and Mill................................................................................... 212 21.1.5 Cobalt Production Facility – Direct Capital Cost ..................................................... 212 21.1.6 Cobalt Production Facility – Indirect Capital Cost .................................................. 213 21.1.7 Contingency – CPF .................................................................................................. 213 21.1.8 Closure Costs ............................................................................................................ 214

21.2 OPERATING COST ESTIMATE.................................................................................. 214 21.2.1 Mining Operating Cost ............................................................................................. 215 21.2.2 Mill/Concentrator Operating Cost ............................................................................ 215 21.2.3 Cobalt Production Facility Operating Cost .............................................................. 216 21.2.4 Residue Disposal ...................................................................................................... 216 21.2.5 General and Administrative Operating Costs ........................................................... 216 21.2.6 Selling Costs and Royalty ........................................................................................ 217

22.0 ECONOMIC ANALYSIS ...................................................................................... 218 22.1 BASIS OF EVALUATION .......................................................................................... 218 22.2 MACRO-ECONOMIC ASSUMPTIONS ........................................................................ 218

22.2.1 Exchange Rate and Inflation .................................................................................... 218 22.2.2 Weighted Average Cost of Capital ........................................................................... 218 22.2.3 Expected Metal Prices .............................................................................................. 219 22.2.4 Taxation Regime ...................................................................................................... 219 22.2.5 Royalty ..................................................................................................................... 219 22.2.6 Selling Expenses....................................................................................................... 220

22.3 TECHNICAL ASSUMPTIONS ..................................................................................... 220 22.3.1 Mine Production Schedule ....................................................................................... 220 22.3.2 Operating Costs ........................................................................................................ 222 22.3.3 Capital Costs............................................................................................................. 222

22.4 BASE CASE CASH FLOW ........................................................................................ 223

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22.5 DISCOUNTED CASH FLOW EVALUATION ................................................................ 224

22.6 SENSITIVITY STUDY ............................................................................................... 226 22.7 CONCLUSION.......................................................................................................... 226

23.0 ADJACENT PROPERTIES .................................................................................. 227

24.0 OTHER RELEVANT DATA AND INFORMATION ........................................ 228

25.0 INTERPRETATION AND CONCLUSIONS ...................................................... 229 25.1 GEOLOGY AND MINERAL RESOURCES ................................................................... 229 25.2 MINING AND MINERAL RESERVES ......................................................................... 229 25.3 ECONOMIC EVALUATION ....................................................................................... 230

26.0 RECOMMENDATIONS ........................................................................................ 232 26.1 GEOLOGY/MINERAL RESOURCES ........................................................................... 232 26.2 MINING .................................................................................................................. 232

26.3 PROCESSING – FUTURE TESTWORK ........................................................................ 233

27.0 DATE AND SIGNATURE PAGE ......................................................................... 235

28.0 REFERENCES ........................................................................................................ 236

29.0 CERTIFICATES ..................................................................................................... 242

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List of Tables

Table 1.1 ICP RAM Deposit at 0.20% Co Cut-off Grade as at 10 March 2015 ............... 3

Table 1.2 Mineral Reserve for ICP at 0.25% Co Cut-off Grade ....................................... 4

Table 1.3 LOM Capital Estimate ....................................................................................... 9

Table 1.4 Summary of LOM Operating Costs ................................................................. 10

Table 1.5 Life-of-Mine Cash Flow Summary ................................................................. 14

Table 2.1 List of Abbreviations ....................................................................................... 22

Table 4.1 ICP Mining Claims .......................................................................................... 28

Table 6.1 FCC’s 1998 ICP Mineral Resources at 0.20% Co Cut-off .............................. 38

Table 6.2 MDA 2001 Ram Deposit Mineral Resource Estimate @ 0.30% Co Cut-off .. 39

Table 6.3 MDA 2005 Resource Estimate (Ram & Sunshine Deposits) at 0.20% Co &

0.30% Co Cut-off ............................................................................................. 39

Table 6.4 MDA 2006 Resource Estimate at 0.30% Co Cut-off ...................................... 40

Table 7.1 Summary of the Stratigraphy of the Ram Deposit .......................................... 46

Table 10.1 ICP Drilling Campaigns .................................................................................. 54

Table 11.1 Summary of Certified Values for Standards used at the ICP .......................... 64

Table 13.1 Summary of Metallurgical Samples ................................................................ 74

Table 13.2 Comparison of Metallurgical Sample Head Grades ........................................ 75

Table 13.3 Comminution Test Results .............................................................................. 76

Table 13.4 Summary of Bulk Concentrate Flotation LCT Results ................................... 78

Table 13.5 Summary of Bulk Concentrate Flotation Variability Results .......................... 79

Table 13.6 Summary of the Differential Flotation LCT Results ....................................... 81

Table 13.7 Multi-Element Bulk Concentrate Analyses ..................................................... 81

Table 13.8 Mintek TCLP Test Results on Leach Residue................................................. 88

Table 13.9 SGS 2017 Leach Test Results ......................................................................... 89

Table 13.10 Mintek 2007 Pilot Plant Average Cobalt Precipitate Analysis ........................ 93

Table 13.11 2005 Mintek Mini Pilot Plant Average Copper Solvent Extraction Results ... 96

Table 14.1 Descriptive Statistics of the Assay Database ................................................. 102

Table 14.2 List of Mineralized and Dilutionary Domain Codes ..................................... 105

Table 14.3 Statistics of Uncapped Composites ............................................................... 110

Table 14.4 Statistics of Capped Composites ................................................................... 110

Table 14.5 Comparison of MDA and Micon Average Values of Composites ................ 110

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Table 14.6 Details of Assay Capping Values by Horizon ............................................... 111

Table 14.7 Summary of Variography Results for Horizon 3023..................................... 113

Table 14.8 Summary Statistics on Specific Gravity Samples ......................................... 114

Table 14.9 Ram Deposit Block Model Attributes ........................................................... 115

Table 14.10 Estimation Parameters for Co, Cu and Au .................................................... 115

Table 14.11 Ram Deposit Mineral Resources at 0.2% Co Cut-off ................................... 121

Table 15.1 Cut-off Grade Criteria ................................................................................... 125

Table 15.2 Mineral Reserve for ICP at 0.25% Co Cut-off Grade ................................... 127

Table 16.1 Principal Structural Trends from 6930 Level Adit ........................................ 130

Table 16.2 Rock Strength for the Ram Deposit ............................................................... 130

Table 16.3 Average RQD Data from 2000 and 2004 for Ram Deposit .......................... 131

Table 16.4 2004 Drilling Longest Stick Measurement .................................................... 132

Table 16.5 Rock Mass Rating Estimate for Ram Deposit from year 2000 drilling ......... 132

Table 16.6 Recommended Ground Support Requirements for Ram Deposit (Minefill,

2006) ............................................................................................................... 134

Table 16.7 Summary of Rock Mass Classification for Mineralized Zones and MFQ at

ICP .................................................................................................................. 135

Table 16.8 Updated ‘Preliminary’ Ground Support Recommendations for Permanent

Openings for ICP ............................................................................................ 136

Table 16.9 Estimated Mine Development Distance ........................................................ 140

Table 16.10 Estimated Mine Development Summary (Footage) ...................................... 144

Table 16.11 Estimated Mine Development Summary (Tonnage) ..................................... 145

Table 16.12 Mining Production Schedule ......................................................................... 147

Table 16.13 Mine Staff ...................................................................................................... 150

Table 16.14 Underground Mine Labour ............................................................................ 151

Table 16.15 Mining Equipment List .................................................................................. 152

Table 16.16 2008 UCS Testing Results ............................................................................. 157

Table 16.17 2017 UCS Testing Results ............................................................................. 157

Table 16.18 Summary of Estimate Binder Addition ......................................................... 159

Table 18.1 Total CPF Make up Water ............................................................................. 187

Table 20.1 Water Treatment Concentrations and Limits................................................. 204

Table 20.2 Water Treatment Systems Comparison ......................................................... 205

Table 20.3 ICP Permits .................................................................................................... 208

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Table 20.4 CPF Permits ................................................................................................... 209

Table 21.1 LOM Capital Estimate ................................................................................... 210

Table 21.2 LOM Mining Capital Estimate ...................................................................... 210

Table 21.3 Mill/Concentrator Capital Estimate ............................................................... 211

Table 21.4 Mine + Mill Indirect Capital Estimate ........................................................... 212

Table 21.5 CPF- Direct Capital Estimate ........................................................................ 212

Table 21.6 CPF Indirect Capital Estimate ....................................................................... 213

Table 21.7 Summary of LOM Operating Costs ............................................................... 214

Table 21.8 LOM Mine Operating Cost Estimate............................................................. 215

Table 21.9 LOM Mill/Concentrator Operating Cost Estimate ........................................ 215

Table 21.10 LOM CPF Operating Cost Estimate .............................................................. 216

Table 21.11 LOM G&A Operating Cost Estimate ............................................................ 216

Table 22.1 Operating Cost Estimate ................................................................................ 222

Table 22.2 Life-of-Mine Cash Flow Summary ............................................................... 223

Table 22.3 Base Case Life of Mine Annual Cash Flow .................................................. 225

Table 25.1 Summary of the Ram Deposit Mineral Resources at 0.2% Co Cut-off ......... 229

Table 25.2 Mineral Reserve for the ICP at 0.25% Co Cut-off ........................................ 229

Table 25.3 Life-of-Mine Cash Flow Summary ............................................................... 230

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List of Figures

Figure 1.1 Cobalt Price and Cobalt Sulphate Premium (Recent and Forecast) ................ 11

Figure 1.2 Annual Mining Schedule ................................................................................. 12

Figure 1.3 Annual Processing Schedule ........................................................................... 12

Figure 1.4 Annual Sales Revenues by Product ................................................................. 13

Figure 1.5 LOM Cash Operating Costs ............................................................................ 13

Figure 1.6 Life-of-Mine Cash Flows ................................................................................ 15

Figure 1.7 NPV Sensitivity Diagram ................................................................................ 16

Figure 4.1 Location Map of the Idaho Cobalt Project ...................................................... 26

Figure 4.2 Plan Showing Layout of the ICP Claims ......................................................... 27

Figure 5.1 Idaho Project Site Access Roads ..................................................................... 31

Figure 6.1 Image of ICP Showing Mill Site and Completed Earthworks after Completion

of Stages I and II Construction ......................................................................... 36

Figure 7.1 Regional Geology of the ICP........................................................................... 42

Figure 7.2 Local Geology of the ICP ................................................................................ 43

Figure 10.1 Ram Deposit Drill Hole Locations .................................................................. 55

Figure 10.2 FCC’s Resident Geologist Displaying 1996 Drill Cores/Sampling Records

during Micon Visit............................................................................................ 57

Figure 10.3 Typical Cross Section through the Ram Deposit ............................................. 58

Figure 11.1 FCC Core Storage Facility in Salmon ............................................................. 62

Figure 11.2 Summary of Blank Samples Results: 1997 to 2006 Drilling........................... 63

Figure 11.3 Control Chart for Co: Standard 1 .................................................................... 64

Figure 11.4 Control Chart for Co: Standard 2 .................................................................... 65

Figure 11.5 Control Chart for Cu: Standard 1 .................................................................... 65

Figure 11.6 Control Chart for Cu: Standard 2 .................................................................... 66

Figure 11.7 Control Chart for Au: Standard 1 .................................................................... 66

Figure 11.8 Control Chart for Au: Standard 2 .................................................................... 67

Figure 12.1 Example of the Ram Deposit Drill Core Photograph ...................................... 69

Figure 12.2 Section Showing the 2016 Metallurgical Drill Hole ....................................... 70

Figure 13.1 SGS-L 2005 LCT Flowsheet ........................................................................... 78

Figure 13.2 SGS-L 2016/17 Differential Flotation LCT Flowsheet ................................... 80

Figure 13.3 Bulk Concentrate Cobalt Recovery Test Results ............................................ 83

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Figure 13.4 Bulk Concentrate Copper Recovery Test Results ........................................... 83

Figure 13.5 Flotation Test Results – Cobalt Recovery vs Cu:Co Ratio ............................. 84

Figure 13.6 Cobalt Sulphate Crystallization Test Process Flowsheet ................................ 95

Figure 14.1 Quantile-Quantile Plot of Co, Cu and Au ...................................................... 101

Figure 14.2 Ram Deposit Global Log Probability Plot for Co ......................................... 102

Figure 14.3 Ram Deposit Global Log Probability Plot for Cu ......................................... 103

Figure 14.4 Ram Deposit Global Log Probability Plot for Au ......................................... 103

Figure 14.5 Section Showing Interpreted Mineral Domains ............................................ 106

Figure 14.6 Ram Deposit Isometric Projection Showing Wireframes Honouring the 5 ft.

Level Plans ..................................................................................................... 107

Figure 14.7 Cobalt Log-probability Plot for Horizon 3021 .............................................. 108



Figure 14.10 Co Variogram along the Major Axis (Strike Direction) for Horizon 3023 ... 112

Figure 14.11 Cu Variogram along the Major Axis (Strike Direction) for Horizon 3023 ... 112

Figure 14.12 Au Variogram along the Major Axis (Strike Direction) For Horizon 3023 .. 113

Figure 14.13 Long Section Distribution of Co Grades in Horizon 3021 ............................ 116



Figure 14.16 Section Through the Ram Block Model for Horizon 3023 ........................... 118

Figure 14.17 Cobalt Swath Plot for the Main Domain Horizon 3021 of the Ram Deposit 119

Figure 14.18 Cobalt Swath Plot for Domain 3022 of the Ram Deposit ............................. 119

Figure 14.19 Cobalt Swath Plot for Domain 3023 of the Ram Deposit ............................. 120

Figure 14.20 Long Section of Horizon 3023 Showing Resource Categories ..................... 122



Figure 16.1 Estimated Ground Support for ICP ............................................................... 136

Figure 16.2 Weak Rock Mass Design Span Curve for Man Entry (Ouchi and Brady,

2004) ............................................................................................................... 137

Figure 16.3 Stability Graph (Unsupported Stopes) ........................................................... 138

Figure 16.4 ICP Mine Development Layout ..................................................................... 141

Figure 16.5 ICP Stope Layout........................................................................................... 142

Page

xiv

Figure 16.6 Schematic of the Mine and Stope Layout ...................................................... 148

Figure 16.7 Schematic of ICP Ventilation Layout ............................................................ 154

Figure 16.8 Backfill Schedule and Material Source. ........................................................ 155

Figure 17.1 Mine Site Block Flow Diagram ..................................................................... 161

Figure 17.2 CPF Refinery Block Flow Diagram – Concentrate Feed Circuit .................. 164

Figure 17.3 CPF Refinery Block Flow Diagram – Cobalt Simplified Circuit .................. 165

Figure 18.1 Site Facility Map ........................................................................................... 176

Figure 18.2 General Mine Site Area Map ......................................................................... 177

Figure 18.3 Water Treatment Block Flow Diagram ......................................................... 181

Figure 18.4 TWSF Plan View ........................................................................................... 182

Figure 18.5 CPF Site Location .......................................................................................... 184

Figure 19.1 ICP Breakdown of Projected Revenue .......................................................... 194

Figure 20.1 Groundwater Sampling Locations ................................................................. 198

Figure 20.2 Mine and Mill - Surface Water and Spring Monitoring Network ................. 199

Figure 20.3 Mine and Mill - Tailings and Waste Storage Facility ................................... 200

Figure 20.4 Mine/Mill Water Balance .............................................................................. 205

Figure 20.5 Cobalt Processing Facility, Blackfoot, Idaho ................................................ 208

Figure 21.1 LOM Annual Mining Capital Expenditure .................................................... 211

Figure 22.1 Cobalt Price and Cobalt Sulphate Premium (Recent and Forecast) .............. 219

Figure 22.2 Annual Mining Schedule ............................................................................... 220

Figure 22.3 Annual Processing Schedule ......................................................................... 221

Figure 22.4 Annual Sales Revenues by Product ............................................................... 221

Figure 22.5 LOM Cash Operating Costs .......................................................................... 222

Figure 22.6 Life-of-Mine Cash Flows .............................................................................. 224

Figure 22.7 NPV Sensitivity Diagram .............................................................................. 226

1

1.0 SUMMARY

1.1 AUTHORIZATION AND PURPOSE

In June, 2016, Formation Capital Corporation, U.S. (FCC), a wholly-owned subsidiary of

eCobalt Solutions Inc., commissioned Micon International Limited (Micon) and its sub-

consultants, SNC Lavalin Inc. (SLI), to prepare a Feasibility Study (FS) for the production of

battery-grade cobalt sulphate along with copper, magnesium sulphate and gold in by-products

from its Idaho Cobalt Project (ICP) in east central Idaho, USA, and to summarise the results

of that study in this Technical Report prepared in accordance with the reporting requirements

of Canadian National Instrument (NI) 43-101. The purpose of this report is to support the

public disclosure of the ICP mineral resources, reserves and the economic results of the FS.

1.2 PROPERTY DESCRIPTION AND OWNERSHIP

The ICP property is 100% owned by FCC and consists of 243 contiguous unpatented lode

mining claims located in east central Idaho, approximately 25.8 miles west of the town of

Salmon. The property covers 4,475 acres centered on 45°07’50” north latitude and 114°21’42”

west longitude.

Presently, the ICP property is not subject to any royalties, other agreements or encumbrances.

1.3 GEOLOGY AND MINERALIZATION

The ICP is hosted in Proterozoic age meta-sediments found on the east side of the central Idaho

Batholith comprising granitic-to-granodioritic rocks. The host sedimentary rocks are believed

to have been part of a large fault-bounded marine sedimentary basin in which dominantly

clastic sediments were deposited. The basin is now part of a supergroup of dominantly quartzite

and argillite metasedimentary rock, the base of which is referred to as the Apple Creek

Formation. All significant copper-cobalt deposits and occurrences are found in the Apple

Creek Formation in a 30 to 35-mile-long linear belt known as the Idaho Cobalt Belt. The

deposits are tabular/stratiform, strike north-northwest, with dips of between 50° and 60° to the

west. Aside from the Ram deposit, which is the focus of this report, there are two other sub-

parallel deposits, namely the Sunshine and East Sunshine which are located about a mile to the

south of the Ram.

Mineralization at the ICP is closely associated with the mafic sequences of the middle unit of

the Apple Creek Formation. Dominant ore minerals include cobaltite (CoAsS) and

chalcopyrite (CuFeS2), with lesser, variable occurrences of gold. Other minerals present in

small quantities are pyrite (FeS2), pyrrhotite (FeS), arsenopyrite (FeAsS), linnaeite ((Co

Ni)3S4), loellingite (FeAs2), safflorite (CoFeAs2), enargite (Cu3AsS4) and marcasite (FeS2).

The nature of the cobalt mineralization suggests that high purity cobalt (>99.9% Co) can be

produced from the deposit. The current study is focussed on producing battery-grade cobalt

2

sulphate heptahydrate. By-products include copper sulphate, magnesium sulphate, copper

concentrate and gold.

1.4 STATUS OF EXPLORATION

There are no current exploration activities. However, the Ram deposit resource remains open

at depth and along strike offering opportunities for expansion. The Sunshine and East Sunshine

deposits are within a mile trucking distance of the Ram and represent additional potential to

the mineral resources of the ICP.

1.5 MINERAL PROCESSING/METALLURGICAL TESTING

A number of metallurgical testwork programs comprising batch and continuous tests have been

completed using representative samples of the RAM deposit mineralization that support the

Feasibility Study process flowsheet. The main testwork programs completed to date include

the following:

Initial milling and flotation testwork on bulk samples and drill composites performed

by Noranda’s nearby Blackbird Mining Company (BMC) in the 1980’s. BMC

reportedly was successful in producing separate copper and cobalt concentrates using

a differential flotation flowsheet.

Early work by The Center for Advanced Mineral and Metallurgical Processing

(CAMP) in 2001 used approximately 1 ton of large diameter drill core from the RAM

deposit. This testwork included a comprehensive milling and flotation test program and

nitrogen species-catalyzed (NSC) leaching of the batch flotation concentrate.

In 2005 SGS Lakefield (SGS-L) conducted a number of flowsheet development

testwork programs including detailed comminution and flotation testing as well as

preliminary leach testing that confirmed CAMP’s NSC test result.

The initial hydrometallurgical tests completed by SGS-L in 2005 provided the design

criteria used for a Mini Pilot Plant testwork campaign undertaken in 2005 by Mintek,

South Africa. This program was directed by Hatch and was successful in developing a

basic hydrometallurgical process.

Pocock Industrial Inc. conducted solids-liquid separation tests in 2005, including

settling/thickening and filtration studies on samples of cleaner concentrate and rougher

flotation tailings.

A pilot plant was operated at Mintek in 2007. This work resulted in improved Fe/Cu

removal, solution purification steps, consistently high grade cobalt product (>99.9%

Co) and introduced of flash cooling technology.

In 2015 Hazen Research completed further flotation and hydrometallurgical testwork

under the direction of Samuel Engineering Inc. (Samuel).

CYTEC Solvay Group (Cytec), conducted bench scale and continuous pilot plant scale

cobalt solvent extraction testwork in 2015 using pregnant leach solution (PLS)

3

generated by Hazen. The objective of this work was to produce a clean cobalt sulphate

solution that could be fed to the crystallizers.

GE Water & Process Technologies (GE) performed crystallizer bench tests in 2015

with the objective of gathering adequate design data in order to confidently size and

estimate the cost of a commercial cobalt sulphate crystallizer. GE also prepared a

capital cost estimates for the magnesium sulphate and copper sulphate crystallizer

packages for the feasibility study.

In 2016 and 2017 SGS-L completed a program of bench scale testwork to confirm the

Feasibility Study design. This work included differential flotation, copper/iron

removal, NSC leaching, leach residue elemental sulphur recovery and gold leaching.

In 2017 SGS-L completed a series of tests to produce copper and cobalt sulphate

crystals.

1.6 MINERAL RESOURCE ESTIMATE

The mineral resource estimate was prepared by Mining Development Associates (MDA) and

was incorporated into the Samuel Engineering Inc. technical report titled “Preliminary

Economic Assessment NI 43-101 Technical Report, Idaho Cobalt Project, Salmon, Idaho,

USA” dated 29 April 2015. Micon has audited and validated the MDA estimation procedures

and mineral resources as detailed in Sections 12 and 14 of this report and summarizes the ICP

mineral resources for the Ram deposit as presented in Table 1.1 below. The mineral resources

are reported at a cut-off grade of 0.20% Co; the copper and gold resources are those resources

carried within the resource blocks which attain the cobalt cut-off grade.

Table 1.1

Ram Deposit Mineral Resources at 0.2% Co Cut-off

Category Zone Co%

Cut-off

Resource

(Tons)

Co

(%)

Co

(000 lbs)

Au

(oz/t)

Au

(oz)

Cu

(%)

Cu

(000 lbs)

Measured All Zones 0.2 1,725,000 0.54 18,590 0.014 24,300 0.76 26,325

Indicated All Zones 0.2 1,711,000 0.64 21,988 0.017 29,900 0.71 24,111

M+I All Zones 0.2 3,436,000 0.59 40,578 0.016 54,200 0.73 50,436

Inferred All Zones 0.2 1,543,000 0.51 15,594 0.012 18,700 0.68 21,032

i. CIM Definition Standards (2014) were followed for mineral resource estimation.

ii. The effective date of this resource estimate is 27 September, 2017.

iii. The mineral resource is estimated at a cut-off grade of 0.20% Co.

iv. The Mineral Resources are estimated using an average long-term cobalt price of USD 14.50 per lb.

v. Totals may not add correctly due to rounding.

1.7 MINERAL RESERVE ESTIMATE

For the ICP, the Measured and Indicated mineral resources from horizons 3021, 3022 and 3023

were considered in the mine plan for conversion into a mineral reserve. The considered

Measured and Indicated mineral resource from these horizons comprise approximately 87% of

the total mineral resource at ICP.

4

Conversion of the mineral resource estimates to mineral reserve was inclusive of the modifying

factors, diluting material and allowances for losses which are to be expected when the material

is mined or extracted.

Stope outlines were generated from 10 ft. vertical level interval shells, honoring the cut-off

grade of 0.25% Co. The resulting shells were transformed into solids and sectioned into

individual stopes of approximately 70 ft. H by 300 ft. L.

Table 1.2 summarizes the mineral reserve estimate for the Idaho Cobalt Project.

Table 1.2

Mineral Reserve for ICP at 0.25% Co Cut-off Grade

Mineral Reserve Class Unit Total or Average

Proven Reserve t’000 1,987

Cobalt Grade % Co 0.43

Copper Grade % Cu 0.69

Gold Grade oz/t 0.013

Cobalt content 000 lb 17,107

Copper content 000 lb 27,384

Gold content oz 25,276

Probable Reserve t’000 1,675







Proven + Probable Reserve t’000 3,662







1.8 MINING METHODS

The mining methods proposed for the Ram deposit are longitudinal longhole stoping and

overhand cut and fill. The selection of these mining methods for the deposit was determined

primarily by the geometry of the mineralized horizons, including factors such as its continuity,

dip and width, and the geotechnical parameters of the rock mass.

The Ram deposit is composed of numerous parallel mineralized horizons, with thickness

ranging from one foot to more than 20 ft., at an average dip of 55° (Samuel, 2015). Three

horizons (3021, 3022 and 3023) contain the majority of the mineralization and only these zones

are considered in the mine design, plan and mineral reserve.

5

Cut and fill mining will be applied to areas dipping less than 50°, or in stopes having widths

ranging from 6 to 10 ft. Conventional cut and fill mining, using hand-held pneumatic drills,

will be carried out in areas having economic mineralized width ranging from 6 to 8 ft. Areas

having widths ranging from 8 to 10 ft. will be mined by mechanized cut and fill. Horizons

wider and steeper than 10 ft. and 50° will be mined with longitudinal longhole stoping. Small

and mid-sized mining equipment was selected to provide a higher selectivity for the proposed

mining methods.

Longhole stoping and cut and fill mining methods will be used to extract 70% and 30% of the

mineral reserve, respectively. In combination, these two mining methods provide a production

capacity in the underground mine that is approximately 10% higher than the nominal mill

capacity (800 t/d). The mine has the capacity to supply approximately 323,000 t/y of ore to the

mill during steady state operation, equivalent to approximately 880 t/d for 365 d/y.

Conservatively, the mine operating cost estimates have been based on achieving this higher

rate of production, whereas in practice it is anticipated that production would keep pace with

milling, and so mining expenditures that are projected over a period of 12 years would in fact

be spread over the whole of the mill operating life of approximately 12.5 years.

Paste prepared from mill tailings will be utilised as backfill material in combination with waste

rock fill arising from mine development

Excavated material will be hauled by 22-t payload low profile trucks to the tram loading area

located at the mine portal, and then loaded into an aerial tramway for final transportation to

the processing plant or waste storage facility, as appropriate.

1.9 PROCESSING

1.9.1 Mill/Concentrator

The recovery of all products is completed in a two-step process in separate locations. Initially,

a concentrate is produced at the mine site near Salmon, Idaho and is then transported for

processing at the proposed Cobalt Processing Facility (CPF) near Blackfoot, Idaho.

The primary facilities at the mine site include the concentrator, paste backfill plant and the

water treatment plant.

Material from the tram is discharged onto separate ore and waste stockpiles located near the

concentrator building. Each stockpile has a live capacity of 800 tons which is sufficient for one

day’s operation in the event of a shutdown for maintenance or repair of the tram. A front end

loader will transfer material from the stockpiles to the hopper of the primary crusher.

Primary and secondary crushing is performed via a jaw crusher, and cone crusher, respectively.

Ore from the crushing circuit is transported via the mill feed conveyor into the ball mill feed

chute and is milled at the rate of 36.2 t/h. Milled slurry is pumped to the flotation circuit,

6

comprising a conditioning tank, rougher flotation and cleaner flotation. The cleaner

concentrate is pumped to the concentrate thickener, where flocculant is added to increase the

density of the slurry in the thickener underflow. The thickened concentrate is pumped to the

concentrate stock tank for storage prior to final dewatering in the concentrate filter press. The

filtered concentrate is trucked to the CPF refinery for final processing.

Tailings are thickened and used in the underground mine as pastefill, or dry-stacked on surface

in the tailings and waste rock storage facility (TWSF).

1.9.2 Cobalt Processing Facility (CPF)

The proposed CPF is a hydrometallurgical plant located near Blackfoot Idaho. It is a

sophisticated processing facility that uses a complex series of processes including pressure

leaching in autoclaves, solvent extraction, crystallization, precipitation, thickening and

filtration to produce a number of products. The products are primarily cobalt sulphate, with

by-product copper sulphate, magnesium sulphate and gold, the latter being produced as loaded

carbon that is sold a third party for elution and gold-winning to conventional doré bullion.

The residual tailings produced in the refinery is shipped via rail using side dump cars to an

offsite waste facility. Waste effluent from the site is disposed of in the municipal sewer line

for treatment in Blackfoot Idaho.

1.10 INFRASTRUCTURE

1.10.1 Mill/Concentrator

Infrastructure at the ICP mine/mill site was partly constructed during an earlier stage of project

development, including:

Completion of the access road from highway 93 to the mine site.

Security/Gate House has been purchased and installed at entrance to the mine site.

Site preparation including stripping and grading.

Earthworks for the first cell of the Tailings Waste Storage Facility (TWSF) was nearly

completed during the 2011 construction phase. Subject to testing and repair, liner

material on site may still be suitable for use.

Some footings have been installed for the crusher building and the mill and

concentrator building.

The administration building has been purchased and installed at site. No utilities have

been installed to the building.

The incoming power supply line was completed during the last phase of construction.

Tie-ins to the supply line and the site distribution system will be made during the next

phase of construction.

7

The road to the portal location and portal bench have also been completed. A Hilkfiker

wall will be constructed during final construction prior to mine development.

A small warehouse and yard south of Salmon Idaho has been purchased. The Salmon

Depot is currently used for storage of the purchased equipment. In future, this site will

be used as a mustering point for construction and operations employees who will be

bussed to site. It will also serve as temporary storage of concentrate prior to shipment

to the CPF and incoming shipments bound for the mine site.

In addition, the following structures will be required:

Crusher Building – has been purchased and is stored at the Salmon Depot.

Concentrator Building – has been purchased and is stored at the Salmon Depot.

o Control Room (enclosure within the Concentrator Building).

o Sample Prep Room (enclosure within the Concentrator Building).

Dry/Change House.

Facilities are also planned for water supply, storage, treatment and discharge.

The tailings and waste rock storage facility (TWSF) is a single surface disposal facility is used

to store both the tailings from the concentrator and the waste rock material. This facility serves

to minimize the area of disturbance by sharing containment and drainage collection facilities

while providing storage for these materials.

1.10.2 CPF Infrastructure

Owing to its location within the Blackfoot municipal area, many of the infrastructural

requirements at the CPF are already in place or relatively simple to complete. These include:

Power supply, from utility connection point adjacent to the Blackfoot site.

Municipal water supply and waste water disposal.

Rail spur and connection to the adjoining rail line.

Communications

1.11 MARKET STUDIES AND CONTRACTS

1.11.1 Market Studies

The feasibility study is based on the recovery of battery grade cobalt sulphate heptahydrate,

together with copper sulphate, magnesium sulphate and gold, and a minor volume of copper

concentrate as saleable by-products.

CRU Consulting (CRU) provided a report, Market Study for the Idaho Cobalt Project (ICP),

dated September, 2017, and which includes the following:

8

An assessment of the battery market and the technologies in use and under development

to support electric vehicles and other rechargeable battery applications.

Analysis of the market for cobalt, with particular emphasis on the use of cobalt sulphate

in the battery market.

Analysis of the current and future supply of cobalt sulphate and accessibility of that

market to the ICP.

An assessment of the market for the associated by-products of the ICP (i.e., copper

sulphate, magnesium sulphate, gold and copper concentrate).

Micon has reviewed the CRU report and supports its rationale for projections of unit revenues

for cobalt, copper and magnesium sulphates, gold and copper in concentrate.

1.11.1.1 Cobalt Sulphate

CRU estimates global cobalt sulphate consumption at 14,544 t contained metal in 2016, a

25.3% y/y increase. It is driven by strong growth in the electric vehicles (EV) sector; in

particular, 23.7% y/y increase in EV, plug-in hybrid electric vehicles (PHEV) and hybrid

electric vehicles (HEV) production.

As a result of its analysis, CRU concludes that the ICP has the opportunity to become a reliable

source of cobalt sulphate to markets within the United States and internationally.

1.11.1.2 By-products

Copper sulphate and magnesium sulphate are used in a wide variety of applications. CRU’s

analysis of trade data indicates that the United States is a net importer of both commodities

and indicates that material from ICP will have the opportunity to displace imported material in

market regions where there is a freight advantage.

1.11.2 Contracts

FCC has not entered into any material contracts relating to development of the ICP.

1.12 ENVIRONMENT STUDIES, PERMITTING AND SOCIAL/COMMUNITY IMPACT

The mine and mill are located on National Forest lands managed by the Salmon-Challis

National Forest. As such it is subject to the National Environmental Policy Act (NEPA). This

requires a thorough series of environmental baseline studies and an Environmental Impact

Statement that provides the Company and state and federal government agencies a complete

property description, identification of all environmental impacts both positive and negative and

the development of mitigation methods to reduce or eliminate negative impacts utilizing best

practices.

9

The Final Environmental Impact Statement (FEIS, June, 2008) discussed the project,

alternatives to the project, environment effects (direct, indirect and cumulative) and

consultation with aboriginal groups, communities and other stakeholders. No issues were

identified that could not be mitigated using best practices.

An extensive environmental monitoring plan has been developed that covers the following:

Water Quality Monitoring

Biological Monitoring

Wetlands Monitoring

Storm Water Monitoring

Weather Monitoring

Air Quality Monitoring

Geochemical Monitoring

A list of permits and authorizations required for the project and their current status is given in

Section 20.0 of this report.

1.13 CAPITAL AND OPERATING COSTS

The capital cost estimate for this project presented herein is considered to be at a feasibility

study level with an accuracy of +15%/-15% and carrying contingencies totaling approximately

12% on initial capital and 9% on LOM capital expenditures.

The LOM capital cost estimate is summarised in Table 1.3. The estimate is given in US dollars

($), with a base date of third quarter, 2017. Owing to rounding of the estimates, some totals

may not agree.

Table 1.3

LOM Capital Estimate

Area Initial Capital

$'000

Sustaining Capital

$'000

LOM Total Capital

$'000

Mining 22,463 70,661 93,124

Processing + Infrastructure 26,355 5,000 31,355

Indirect costs 8,764 0 8,764

Contingency 5,165 3,207 8,373

Sub-total Mine/Mill/Concentrator 62,748 78,869 141,616

Direct - CPF 88,861 5,000 93,861

Indirect - CPF 20,495 0 20,495

Contingency 14,644 0 14,644

Sub-total Cobalt Production Facility 124,000 5,000 129,000

Rehabilitation and Mine Closure 588 16,942 17,530

TOTAL 187,336 100,810 288,146

10

The estimated life-of-mine total project operating costs are summarized in Table 1.4.

Table 1.4

Summary of LOM Operating Costs

Area

LOM total

Operating Costs

($’000)

Unit cost $/tonne

milled

$/lb Contained

Co in sulphate

Mining 196,692 53.71 6.19

Mill/Concentrator 52,494 14.34 1.65

Transport (residue disposal) 5,199 1.42 0.16

Hydromet Plant (CPF) 149,121 40.72 4.69

G&A 37,309 10.19 1.17

Sub-total Direct Operating Costs 440,815 120.38 13.88

Selling Costs 2,117 0.58 0.07

Total Cash Operating Costs before

by-product credits

442,932 120.96 13.94

Less By-product credits (282,510) (77.15) (8.89)

Cash Operating Costs (net) 160,422 43.81 5.05

1.14 ECONOMIC ANALYSIS

1.14.1 Global Assumptions

Micon has prepared its assessment of the project on the basis of a discounted cash flow model,

from which Net Present Value (NPV), Internal Rate of Return (IRR), payback and other

measures of project viability can be determined. Assessments of NPV are generally accepted

within the mining industry as representing the economic value of a project after allowing for

the cost of capital invested. Micon has applied a real discount rate of 7.5% as its base case.

Price assumptions for each product and by-product are given in United States dollar ($) terms

and, unless otherwise stated, all financial results are also expressed in U.S. dollars. All material

capital and operating cost estimates and other inputs to the cash flow model for the project

have been prepared using constant, third quarter 2017 money terms, i.e., without provision for

escalation or inflation. Since these costs are estimated in U.S. dollars, no exchange rate

assumptions are relevant.

The base case cash flow projection assumes a variable price of cobalt metal, with cobalt

sulphate heptahydrate (with a minimum grade of 20.5% Co) trading at a premium of around

2% on a 100% cobalt basis. The basis for these price assumptions are discussed in Section 19

of this report. Figure 1.1 shows the annual prices and premium applied for cobalt in sulphate.

11

Figure 1.1

Cobalt Price and Cobalt Sulphate Premium (Recent and Forecast)

Copper sulphate sales are forecast at a constant price of $2.60/lb Cu, with a premium of 54%

for the sulphate resulting in gross revenue of $4.00/lb Cu. Copper concentrate sales are forecast

with payability of 98%, treatment charges of $185/t including transport, and $0.10/lb Cu

refining. Gold revenue and credits are based on a price of $1,200/oz Au, and magnesium

sulphate sales are forecast on a price averaging $250/t MgSO4.

Idaho state and U.S. federal income taxes payable on the project have been provided for in the

cash flow forecast after deductions for relevant depreciation allowances.

The net taxes payable on the forecast project cash flow has been estimated by an independent

third party with specialist expertise in this area, and Micon has relied on this analysis in its

economic evaluation of the project.

No royalty has been provided for in the cash flow model.

1.14.2 Technical Assumptions

Figure 1.2 shows the annual tonnage of mill-feed material mined from underground, as well as

the mill head grades for cobalt, copper and gold content.

0.0%

1.0%

2.0%

3.0%

4.0%

5.0%

6.0%

-

5.00

10.00

15.00

20.00

25.00

30.00

35.00

Pre

miu

m (

%)

$/l

b

Co (99.3%) Premium for Sulphate (%)

12

Figure 1.2

Annual Mining Schedule

As shown in the figure, the grade of the mill feed demonstrates the focus on higher cobalt

grades in the early part of the production period. Material with a relatively high copper/cobalt

ratio of 2.0 or more is extracted later in the mine life. Treatment of this material necessitates

the commissioning of a copper scalping circuit, the construction of which is provided for in

the sustaining capital estimate.

Annual production of cobalt and by-products over the LOM period is shown in Figure 1.3.

Figure 1.3

Annual Processing Schedule

Annual sales value of cobalt and by-products over the LOM period is shown in Figure 1.4.

0.000

0.200

0.400

0.600

0.800

1.000

1.200

0.0

50.0

100.0

150.0

200.0

250.0

300.0

350.0

1 2 3 4 5 6 7 8 9 10 11 12 13

Gra

de

(p

pm

Au

, % C

o, C

u)

Mill

ed

(0

00

t)

Mill feed (ROM) Cobalt Grade Copper Grade Gold Grade

0

5,000

10,000

15,000

20,000

25,000

30,000

0

2

4

6

8

10

12

Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12 Yr13

MgS

O4

.7H

2O

(t)

Co

SO4

.7H

2O

, Cu

SO4

.5H

2O

, C

u-C

on

c (t

),G

old

(o

z)

Cobalt Sulphate Copper Sulphate Gold in doré Copper Conc. Magnesium Sulphate

13

Figure 1.4

Annual Sales Revenues by Product

Over the LOM period, cobalt sulphate sales account for 75% of total revenue. Copper sulphate

contributes a further 15%, magnesium sulphate 5%, gold 4%, and copper concentrates 1%.

Figure 1.5 shows cash operating expenditures over the LOM period.

Figure 1.5

LOM Cash Operating Costs

0

20,000

40,000

60,000

80,000

100,000

120,000

140,000

160,000


Rev

enu

e ($

'00

0)

Cobalt Sulphate Copper Sulphate Magnesium Sulphate Gold in doré Copper Conc.

0

5,000

10,000

15,000

20,000

25,000

30,000

35,000

40,000

45,000

50,000


Rev

enu

e ($

'00

0)

Selling Costs (Cu-conc, gold) Mining Mill/Concentrator Transport CPF G&A

14

Pre-production capital expenditures are estimated to total $186.75 million. This sum includes

$22.46 million for mining, $26.36 million in the milling/concentrator plant, $88.86 million in

the hydrometallurgical plant (CPF), $29.26 million indirect costs and owner’s costs, and

contingencies totalling $19.81 million.

Sustaining capital is estimated at $83.87 million over the LOM period, mainly for underground

development but including $10 million for retro-fitting a copper sulphide scalping circuit. A

further $17.53 million is required to cover mine closure and associated bonding costs.

Working capital has been estimated to include 30 days allowance for product inventory on site,

in transit, and accounts receivable on concentrates delivered. Stores provision is for 30 days of

consumables and spares inventory, less 60 days accounts payable. On this basis, an average of

$4.92 million of working capital is required during the mine/mill operating period.

The LOM base case project cash flow is presented in Table 1.5. Annual cash flows are

summarized in Figure 1.6 (following page).

Table 1.5

Life-of-Mine Cash Flow Summary

Item LOM total

($ 000) $/t milled $/lb Cobalt

Cobalt Sales 846,837 231.26 26.66


Mining 196,692 53.71 6.19


Transport 5,199 1.42 0.16

CPF 149,121 40.72 4.69

G&A 37,309 10.19 1.17

Total Operating Costs 442,932 120.96 13.94

By-product credits (282,510) (77.15) (8.89)

Net Operating Costs 160,422 43.81 5.05

EBITDA 686,415 187.45 21.61

Capital Costs 288,146 78.69 9.07

Net cash flow before tax 398,269 108.76 12.54

Tax 66,814 18.25 2.10

Net cash flow after tax 331,454 90.51 10.43

15

Figure 1.6

Life-of-Mine Cash Flows

The project demonstrates an undiscounted pay back of 3.3 years, or approximately 4.0 years

when discounted at 7.5%, leaving a tail of over 8 years of production.

1.14.3 Discounted Cash Flow Evaluation

The base case evaluates to an IRR of 25.1% before taxes and 21.3% after tax. At a discount

rate of 7.5%, the net present value (NPV7.5) of the cash flow is $177 million before tax and

$136 million after tax.

1.14.4 Sensitivity

The sensitivity of project returns to changes in all revenue factors (including grades, recoveries,

prices and exchange rate assumptions) and also to capital and operating costs was tested over

a range of 30% above and below base case values. See Figure 1.7, showing net present values

on an after-tax basis.

The chart suggests that the project is most sensitive to revenue drivers, moderately sensitive to

operating costs and least sensitive to changes in capital cost. Within a range of 30% above and

below base case values, operating and capital costs both maintain a positive NPV outcome.

16

Figure 1.7

NPV Sensitivity Diagram

1.14.5 Conclusion

Micon concludes that this study demonstrates the potential viability of the project within the

range of accuracy of the estimated capital and operating costs, production forecast, and price

assumptions.

Micon and SLI have concluded that the study contains adequate detail and information to

support this positive outcome. Standard industry practices, equipment and design methods

were used in the study. Micon and SLI further conclude that the ICP contains a viable cobalt

and base metal resource that can be mined by underground methods and recovered with a

combination of both conventional and state of the art processing technologies. Using the

assumptions described herein, the project is economic and further development is warranted.

1.15 CONCLUSION AND RECOMMENDATIONS

1.15.1 Geology and Resources

The Ram deposit consists of a Hanging-wall Zone with 3 primary and 4 minor horizons, a

Main Zone comprising 3 horizons, and a Footwall Zone with 3 horizons. These sub-parallel

horizons generally strike N15oW and dip 50o-60o to the northeast.

The Main Zone (horizons 3021, 3022 and 3023) of the Ram deposit contribute about 87% of

the Measured and Indicated resource. The exact extents and significance of the Hanging-wall

and Foot-wall Zones horizons of the deposit remain to be fully investigated and could have

material effect in terms of increasing the resource and life of mine.

(50,000)

0

50,000

100,000

150,000

200,000

250,000

300,000

70 75 80 85 90 95 100 105 110 115 120 125 130

$'0

00

Percentage of Base Case

Prices Operating Costs Capital Expenditure

17

The mineralization of the Ram deposit remains open at depth (down-dip) and along strike. The

geological corridor/structure controlling the mineralization is persistent for the entire strike

length of FCC’s ICP area and beyond. The already known Sunshine deposit is within easy

reach (i.e. only a mile south) from the infrastructure at the Ram. Hence, the outlook in terms

of increasing the resource is favourable.

It should also be noted that previous drill-testing by earlier operators in the greater region

identified additional areas of mineralization near the ICP deposits. These mineralized zones

represent promising targets for future drilling.

In Micon’s view, the critical issues pertaining to the successful development of the ICP are

precision in predicting the grade and geometry of the various components of the deposit and

availability of additional resources to sustain the operations. To address these issues, Micon

makes the following recommendations:

While the block size of 6 ft. by 2 ft. by 5 ft. is an appropriate size for the narrow deposit

widths encountered at the RAM deposit and the envisaged SMU, the ability to estimate

grades and geometry with precision to this resolution requires a much closer drill

spacing. Accordingly, infill development drilling and/or development drifting is

recommended prior to commercial underground mining production and before final

stope design. The suggested infill drill hole spacing is 30 to 35 ft.

Concurrently with infill development drilling, a drilling program to upgrade the

Inferred mineral resources should be initiated to increase the life of mine.

Additional exploration in the form of systematic step-out drilling should be conducted

following the main trend of mineralization in the north-westerly and south easterly

direction along strike and down dip.

A review and mineral resource update of the Sunshine and East Sunshine deposits is

recommended together with economic studies on trucking ore from these deposits to

the facilities at the Ram.

1.15.2 Mining

The following summarizes the recommendations observed during the preparation of the current

feasibility, even though “all drill-hole information, geotechnical data, and hydrological data

have been developed to a feasibility level” (PEA, 2015) in previous studies:

Backfill testing – Results from the 2017 pastefill material testing indicates that the

strength of the pastefill is not dependent of the type of cement or binder types but rather

on the water:cement ratio (P&C, 2017). It is recommended that additional material

testing to be carried out with increased binder addition to the current testing matrix

(50% cement: 50% slag) to potentially reduce the cement costs.

Backfill plant – Currently the backfill plant has only one silo for the cement storage. A

trade-off study to identify the technical and cost benefit for an additional binder storage

18

system will be advantageous to the project. Micon agrees with P&C that a re-evaluation

“of the paste delivery pipeline system design completed in 2008 to ensure that the

expected pumping pressures are appropriate given the change in tailings properties

from 2008 to 2017” (P&C, 2017) and to ensure these are compatible to the purchased

backfill equipment.

Geotechnical – Minefill indicated that the available geotechnical data is limited, and

additional data collection is warranted, including laboratory uniaxial compressive

strength testing. To date, there have been no geotechnical drillholes completed at the

Ram and no oriented core measurements collected from this deposit. Adit mapping

from the neighbouring Black Bird mine was carried out at adjacent adits near the Ram

portal, however the mapping of those adits encountered none of the principal units

expected in the Ram deposit (Minefill, 2006). Additional classification of the rock mass

in relation to it spatial location will assist in stope dimension and overall mine design.

The current mine design is very similar to the mine design in which the mine ventilation

study was performed. An updated mine ventilation study is recommended in the next

phase of engineering study before construction when a finalized mine design is

available.

Optimization – Additional optimization of the mine design, plan and especially the

production schedule can potentially improve the economics of the project.

1.15.3 Processing – Future Testwork

Copper Flotation – Additional tests are recommended to verify the copper scalping and

cleaning flotation performance using fresh samples that represent the relatively high

Cu:Co ratio mineralization planned to be mined and processed in the later years of the

mine life.

Cobalt Solvent Extraction – Pilot plant cobalt solvent extraction testwork needs to be

completed in order to provide design details for the process. The objective of this

additional testwork will be to confirm extraction kinetics, determine optimum percent

solids MgO vs. cobalt recovery, confirm Co/Mg selectivity, determine strip liquor

impurities and confirm the overall circuit mass balance. The cobalt and zinc stripping

conditions also need to be confirmed.

Copper Solvent Extraction – The design of the copper solvent extraction circuit is based

on the 2005 mini pilot plant test program, the object of which was to produce cathode

copper not copper sulphate crystals. There may be a benefit of reviewing this circuit as

the differences in the optimal PLS specifications for these two applications

(electrowinning vs crystallization) could result in a simpler system and lower capital

costs.

Crystallization – Although adequate bench scale testwork has been completed to

provide a design for the cobalt crystallizer circuit, additional detailed work needs to be

completed to establish the actual maximum recovery rate per pass and the critical

impurity concentration prior to the finalized design and procurement of the system. It

is recommended that extended continuous operations be performed using a high purity

19

feed electrolyte to produce additional cobalt sulphate crystals and investigate the

impact of impurity build-up of the product over a more prolonged period of operation.

A process to treat the bleed stream and recycle cobalt will also need to be developed.

Successful production of cobalt crystals from project representative concentrate based

solutions rather than synthetically prepared solutions should also be demonstrated.

Testwork needs to be completed using representative solution samples to provide

detailed design details of the magnesium sulphate crystallizer circuit.

Based on the recent copper crystallization testwork at SGS-L, it is recommended to

perform additional neutralization tests on both the feed solution and the copper raffinate

with the objective to (i) minimize cobalt and copper losses in the primary precipitate

stage and (ii) reduce the copper concentration in the feed to cobalt recovery, without

losing cobalt to the copper precipitate. This work should also include an evaluation of

a two stage precipitation process at two target pH levels for both processes.

Gold Recovery Circuit – Additional testwork is required to optimize the elemental

sulphur flotation and the cyanide leaching circuit circuits. Testwork also needs to be

completed in order to model the CIL circuit and gold/silver carbon loading as well as

the cyanide destruction circuit.

CPF Pilot Plant – Much of the CPF processing circuits have been designed using batch

tests or continuous pilot tests using synthetic solutions. It is therefore recommended

that the complete CPF process be tested using a continuous pilot plant using composite

samples of flotation concentrate.

During the pilot plant testwork program it is suggested that solid/liquid separation and

washing of precipitates should be evaluated using pressure filtration and/or

centrifuging to develop an industrially robust methodology for removing the

precipitates produced within the process flowsheet.

Process Modelling and Simulation – As part of the feasibility study process engineering

completed by SLI, a MetSim model was developed for the CPF. This model needs to

be developed to a higher level of detail using the results from the additional testwork

recommended above. The more robust model will be available to stress test the final

detailed design of the CPF.

HAZOP Studies – During the detailed design phase it is important to complete a hazard

and operability study (HAZOP) in order to identify and evaluate potential risks to

personnel or equipment so that the design can mitigate these risks.

20

2.0 INTRODUCTION

2.1 AUTHORIZATION AND PURPOSE

In June, 2016, Formation Capital Corporation, U.S. (FCC), a wholly-owned subsidiary of

eCobalt Solutions Inc. (eCobalt), commissioned Micon International Limited (Micon) and its

sub-consultants, SNC Lavalin Inc. (SLI), to prepare a Feasibility Study (FS) for the production

of battery-grade cobalt sulphate along with copper and gold in by-products from its Idaho

Cobalt Project (ICP) in east central Idaho, USA, and to summarise the results of that study in

this Technical Report prepared in accordance with the reporting requirements of Canadian

National Instrument (NI) 43-101. The purpose of this report is to support the public disclosure

of the ICP mineral resources, reserves and the economic results of the FS.

This report is intended to be used by FCC subject to the terms and conditions of its agreement

with Micon. That agreement permits eCobalt to file this report as an NI 43-101 Technical

Report with the Canadian Securities Administrators (CSA) pursuant to provincial securities

legislation. Except for the purposes legislated under provincial securities laws, any other use

of this report, by any third party, is at that party’s sole risk.

The requirements of electronic document filing on SEDAR (System for Electronic Document

Analysis and Retrieval, www.sedar.com) necessitate the submission of this report as an

unlocked, editable pdf (portable document format) file. Micon accepts no responsibility for

any changes made to the file after it leaves its control.

The conclusions and recommendations in this report reflect the authors’ best judgment in light

of the information available to them at the time of writing. The authors and Micon reserve the

right, but will not be obliged, to revise this report and conclusions if additional information

becomes known to them subsequent to the date of this report. Use of this report acknowledges

acceptance of the foregoing conditions.

Micon does not have nor has it previously had any material interest in FCC or related entities.

The relationship with FCC is solely a professional association between the client and the

independent consultant. This report is prepared in return for fees based upon agreed

commercial rates and the payment of these fees is in no way contingent on the results of this

report.

This report includes technical information, which requires subsequent calculations or estimates

to derive sub-totals, totals and weighted averages. Such calculations or estimations inherently

involve a degree of rounding and consequently introduce a margin of error. Where these occur,

Micon does not consider them material.

2.2 SOURCES OF INFORMATION

The principal sources of information for this report are:

21

Previous NI 43-101 Technical Reports on the ICP filed on SEDAR, as noted and

referenced in Sections 6 and 27, respectively.

Drill hole databases supplied by FCC.

ICP Ram deposit block model supplied by Mining Development Associates (MDA).

Observations made during the site visits by Micon.

Discussions/meetings with FCC management/staff/consultants familiar with the

property.

Data/reports supplied by FCC and its consultants.

Metallurgical Reports by SGS (2015, 2016 and 2017).

Experience gained while working on similar deposits.

Micon is pleased to acknowledge the helpful cooperation of FCC’s

management/staff/consultants who made all data requested available and responded openly

and helpfully to all questions, queries and requests for material.

All units of length in this report are reported in the imperial system. For rock and base metals,

mass units are in pounds avoirdupois (lb) and short tons (T, each of 2,000 lb). Precious metal

values are typically given in troy ounces (oz) and grades as oz. per short ton. Currency values

are expressed in United States dollars ($ or USD), unless otherwise indicated.

2.3 SCOPE OF PERSONAL INSPECTION

Micon conducted a site visit to the ICP from 13 to 14 July 2016. During the visit, Micon

discussed the geologic model, examined drill cores, reviewed drill hole logs, reviewed

mineralization types and discussed the quality assurance/quality control (QA/QC) protocols

used by FCC. Micon also inspected the existing infrastructure/facilities and materials already

purchased in anticipation of future mining activities.

Previously, Micon visited the ICP on 9 December 2010 whilst undertaking a due diligence

study of the property on behalf of a financial institution.

2.4 LIST OF ABBREVIATIONS

All currency amounts in this report are stated in US dollars ($), unless otherwise stated.

Quantities are generally stated in imperial units, following US conventional practice, including

pounds avoirdupois (lb), short tons (T) of 2,000 lbs each; feet, yards and miles for distance;

acres for area; weight percent (%) for cobalt (Co) and copper (Cu) grades and troy ounces per

short ton (oz/t Au) for gold grades. Precious metal grades may be expressed in parts per billion

(ppb) or parts per million (ppm) and their quantities may also be reported in troy ounces (oz).

Abbreviations used in this report are listed in Table 2.1.

22

Table 2.1

List of Abbreviations

Abbreviation Term o Degree(s) oC Degree(s) Centigrade o C-days Degree Centigrade days oF Degree(s) Fahrenheit

< Less than

> Greater than

μg/L Micrograms per litre

μm Micrometre(s) (micron = 0.001 mm)

% Percent, percentage

’ Minutes of latitude and longitude

3D Three dimensional

A Ampere(s)

AAS Atomic absorption spectroscopy

acfm Actual cubic feet per minute

Ag Silver

Al Aluminum

amsl Above mean see level

ANFO Ammonium nitrate-fuel oil

ARD/ML Acid Rock Drainage /Metal Leaching

As Arsenic

Au Gold

B Billion

C Carbon

Ca Calcium

cfm Cubic feet per minute

cm Centimetre(s)

CPF Cobalt Processing Facility

Co Cobalt

Cu Copper

CV Coefficient of variation

d Day(s)

dB(A) Decibel(s) (adjusted)

EPCM Engineering, procurement and construction management

F Fluorine

Fe Iron

FOB Free on board

ft Foot, feet

FW Footwall

g Gram(s)

g Acceleration due to gravity

g/L Grams per litre

g/t Grams per tonne

Ga Billion years (old, ago)

GA General arrangement

gal Gallon(s) (US)

GHG Green House Gas (emissions)

gpm Gallons per minute

GPS Global positioning system

23

Abbreviation Term

GWh Gigawatt-hour

H Hydrogen

h Hour(s)

h/d Hours per day

h/w Hours per week

ha Hectare(s)

HAZOP Hazard and operability study

HDPE High density polyethylene

HP horsepower

HQ Diamond drill core size 63.5 mm (inside diameter of core tube)

Hz Hertz

ICP Idaho Cobalt Project

in Inch(es)

IRR Internal Rate of Return

J Joule(s)

K Potassium

k Kilo (thousand)

kcfm Thousand cubic feet per minute

kg Kilogram(s)

kg/h Kilograms per hour

kg/m3 Kilograms per cubic metre

km Kilometre(s)

km/h Kilometres per hour

kPa Kilopascal(s)

kV Kilovolt(s)

kVA Kilovolt-ampere(s)

kW Kilowatt(s)

kWh Kilowatt hour

kWh/t Kilowatt hours per tonne

L Litre(s)

lb Pound(s)

LCT Locked cycle test

LHD Load-haul-dump

LME London Metal Exchange

LOM Life of mine

M mega (million)

m Metre(s)

m3/h Cubic metres per hour

m/min Metres per minute

Ma Million years (old, ago)

masl Metres above sea level

MCC Motor control centre

min Minute(s)

ML Million litres

mL Millilitres

mm Millimetre(s)

mg/L Milligrams per litre

Mg Magnesium

MPa Megapascal(s)

MW Megawatt(s)

Na Sodium

24

Abbreviation Term

NAG Net acid generating

NI 43-101 Canadian National Instrument 43-101

NO2 Nitrous oxide

NPV Net present value

NSR Net smelter return

NWT Northwest Territories

oz Ounce(s), troy ounces

oz/ton Ounces per ton (short ton, 2,000 pounds)

P&ID Process and instrumentation diagram

Pa Pascal(s)

Pa.s Pascal-second

Pb Lead

ppb Parts per billion

ppm Parts per million

P3 Public private partnerships

QA Quality assurance

QA/QC Quality assurance/quality control

QC Quality control

RBC Rotating biological contactor

RMB Chinese Renminbi

ROM Run-of-mine

rpm Revolutions per minute

RQD Rock quality designation

s Second(s)

S Sulphur

SAG Semi-autogenous grinding

Sb Antimony

SEM Scanning electron microscope

SG Specific gravity

SI International system of units

Si Silicon

SO2 Sulphur dioxide

T Ton(s) – short (2,000 lb)

t/h Short Tons per hour

t/y Short Ton per year

t/d Short Tons per day

TWSF Tailings and Waste Rock Storage Facility

UCS Unconfined compressive strength

US$ United States dollar(s)

V Volt(s)

XRF Energy-dispersive x-ray fluorescence

y Year(s)

yd3 Cubic yard(s)

Zn Zinc

25

3.0 RELIANCE ON OTHER EXPERTS

Micon has reviewed and analyzed data provided by FCC, and has drawn its own conclusions

therefrom, augmented by its direct field examination. Micon has not carried out any

independent exploration work, drilled any holes or carried out any sampling or assaying on the

property, other than examining/verifying mineralization in drill cores and reviewing analytical

and QA/QC procedures/results. While exercising all reasonable diligence in checking,

confirming and testing it, the authors of this report have relied upon FCC’s presentation of data

for the ICP and the findings of its consultants in formulating their opinion.

The various agreements under which FCC holds title to the mineral lands for this project have

not been thoroughly investigated or confirmed by the authors and no opinion is offered as to

the validity of the mineral title claimed. The descriptions were provided by FCC.

The description of the property is presented here for general information purposes only, as

required by NI 43-101. The authors are not qualified to provide professional opinion on issues

related to mining and exploration lands title or tenure, royalties, permitting and legal and

environmental matters. Accordingly, the authors have relied upon the representations of the

issuer, FCC, for Section 4.0 of this report, and have not verified the information presented

therein.

Those portions of the report that relate to the location, property description, infrastructure,

history, deposit types, exploration, drilling, sampling and assaying (Sections 4.0 to 11.0) are

taken, at least in part, from current and previous texts prepared by Formation Capital staff, the

2015 Preliminary Economic Assessment Technical Report by Samuel Engineering Inc. and

other updated information provided by FCC. Micon has relied on these data, supplemented by

its own observations at site.

Some of the figures and tables for this report were reproduced or derived from reports written

for FCC, but the majority of the photographs were taken by Micon during the site visit. Where

the figures and tables are derived from sources other than Micon, the source is acknowledged

below the figure or table.

26

4.0 PROPERTY DESCRIPTION AND LOCATION

The following description is largely excerpted from the 2015 PEA Technical Report by Samuel

Engineering Inc., with minor edits and additions.

4.1 LOCATION AND GENERAL DESCRIPTION

The ICP Property consists of 243 contiguous unpatented lode mining claims located in east

central Idaho, approximately 25.8 miles (41.5 km) west of the town of Salmon, as shown on

the location map provided in Figure 4.1.

Figure 4.1

Location Map of the Idaho Cobalt Project

Source: Map supplied by FCC, 2016

The property covers approximately 4,475 acres centered on 45°07’50” north latitude and

114°21’42” west longitude. It is within the Gant Mountain 7.5-minute quadrangle of the USGS

Topographic Map Series. More specifically, the ICP unpatented mining claims are located in

27

Sections 8, 9, 15, 16, 17, 18, 20, 21, 22, 23, 26, 27, 28, 29, 33, 34 and 35, Township 21 North,

Range 18 East (Figure 4.2). The claim block is within the Salmon-Cobalt Ranger District of

the Salmon-Challis National Forest (Prenn, 2005), lands under surface use administration by

the United States Forest Service (USFS). The mine portal is located at an elevation of

approximately 7,060 ft above sea level, and the processing plant and most of the site

infrastructure is located on Big Flat, which is approximately 930 ft above the mine.

4.2 LAND TENURE

The layout of the claims that comprise the ICP is shown in Figure 4.2 and the claims are listed

in Table 4.1. All claims are owned by Formation Capital Corporation, U.S.

Figure 4.2

Plan Showing Layout of the ICP Claims

Source: Map supplied by FCC, 2017

28

Table 4.1

ICP Mining Claims

Claim Name County # IMC # Claim Name County # IMC # Claim Name County # IMC #

Chelan No. 1 (A) 248345 175861 HZ 22 224194 174660 NFX 49 307262 218717

Chip 1 248956 184883 HZ 23 224195 174661 NFX 50 307263 218718

Chip 2 248957 184884 HZ 24 224196 174662 NFX 56 307269 218724

Chip 3 (A) 277465 196402 HZ 25 224197 174663 NFX 57 307270 218725

Chip 4 (A) 277466 196403 HZ 26 224198 174664 NFX 58 307271 218726

Chip 5 (A) 277467 196404 HZ 27 224199 174665 NFX 59 307272 218727

Chip 6 (A) 277468 196405 HZ 28 224200 174666 NFX 60 307273 218728

Chip 7 (A) 277469 196406 HZ 29 224201 174667 NFX 61 307274 218729

Chip 8 (A) 277470 196407 HZ 30 224202 174668 NFX 62 307275 218730

Chip 9 (A) 277471 196408 HZ 31 224203 174669 NFX 63 307276 218731

Chip 10 (A) 277472 196409 HZ 32 224204 174670 NFX 64 307277 218732

Chip 11 (A) 277473 196410 HZ Frac. 228967 177254 Powder 1 269506 190491

Chip 12 (A) 277474 196411 JC 1 224165 174631 Powder 2 269505 190492

Chip 13 (A) 277475 196412 JC 2 224166 174632 Ram 1 228501 176757

Chip 14 (A) 277476 196413 JC 3 224167 174633 Ram 2 228502 176758

Chip 15 (A) 277477 196414 JC 4 224168 174634 Ram 3 228503 176759

Chip 16 (A) 277478 196415 JC 5 (A) 245689 174635 Ram 4 228504 176760

Chip 17 (A) 277479 196416 JC 6 224170 174636 Ram 5 228505 176761

Chip 18 (A) 277480 196417 JC 7 Frac. 224171 174637 Ram 6 228506 176762

Chip 21 Frac. 306059 218113 JC 8 Frac. 224172 174638 Ram 7 228507 176763

Chip 22 Frac. 306060 218114 JC 9 228054 176750 Ram 8 228508 176764

Chip 23 306025 218115 JC 10 228055 176751 Ram 9 228509 176765

Chip 24 306026 218116 JC 11 228056 176752 Ram 10 228510 176766

Chip 25 306027 218117 JC-12 228057 176753 Ram 11 228511 176767

Chip 26 306028 218118 JC-13 228058 176754 Ram 12 228512 176768

Chip 27 306029 218119 JC 14 228971 177250 Ram 13 (A) 245700 181276

Chip 28 306030 218120 JC 15 228970 177251 Ram 14 (A) 245699 181277

Chip 29 306031 218121 JC 16 228969 177252 Ram 15 (A) 245698 181278

Chip 30 306032 218122 JC 17 259006 187091 Ram 16 (A) 245697 181279

Chip 31 306033 218123 JC 18 259007 187092 Ram Frac.1 (A) 245696 178081

Chip 32 306034 218124 JC 19 259008 187093 Ram Frac.2 (A) 245695 178082

Chip 33 306035 218125 JC 20 259009 187094 Ram Frac.3 (A) 245694 178083

Chip 34 306036 218126 JC 21 259010 187095 Ram Frac.4 (A) 245693 178084

Chip 35 306037 218127 JC 22 259011 187096 South ID 1 (A) 248725 175874

Chip 36 306038 218128 LDC Frac.1 (A) 248720 175880 South ID 2 (A) 248726 175875




DEWEY Frac.(A) 248739 177253 LDC Frac.5 (A) 248724 175884 South ID 6 (A) 248716 176744

Goose 2 (A) 259554 175863 LDC-1 224140 174579 South ID 7 306433 218216

Goose 3 227285 175864 LDC-2 224141 174580 South ID 8 306434 218217


Goose 6 227282 175867 LDC-5 224144 174583 South ID 10 306436 218219

29

Claim Name County # IMC # Claim Name County # IMC # Claim Name County # IMC #





Goose 12 (A) 259548 175873 LDC-10 224149 174588 Sun 1 222991 174156

Goose 13 228028 176729 LDC-11 224150 174589 Sun 2 222992 174157

Goose 14 (A) 259547 176730 LDC-12 224151 174590 Sun 3 (A) 245690 174158

Goose 15 228030 176731 LDC-13 (A) 248718 174591 Sun 4 222994 174159

Goose 16 228031 176732 LDC-14 (A) 248719 174592 Sun 5 222995 174160

Goose 17 228032 176733 LDC-16 224155 174594 Sun 6 222996 174161

Goose 18 (A) 259546 176734 LDC-18 224157 174596 Sun 7 224162 174628

Goose 19 (A) 259545 176735 LDC-20 224159 174598 Sun 8 224163 174629

Goose 20 228035 176736 LDC-22 224161 174600 Sun 9 224164 174630

Goose 21 228036 176737 NFX 17 307230 218685 Sun 16 (A) 245691 177247

Goose 22 228037 176738 NFX 18 307231 218686 Sun 18 (A) 245692 177249

Goose 23 228038 176739 NFX 19 307232 218687 Sun 19 277457 196394

Goose 24 228039 176740 NFX 20 307233 218688 Sun 20 306042 218133

Goose 25 228040 176741 NFX 21 307234 218689 Sun 21 306043 218134

HZ 1 224173 174639 NFX 22 307235 218690 Sun 22 306044 218135

HZ 2 224174 174640 NFX 23 307236 218691 Sun 23 306045 218136

HZ 3 224175 174641 NFX 24 307237 218692 Sun 24 306046 218137

HZ 4 224176 174642 NFX 25 307238 218693 Sun 25 306047 218138

HZ 5 224413 174643 NFX 30 307243 218698 Sun 26 306048 218139

HZ 6 224414 174644 NFX 31 307244 218699 Sun 27 306049 218140

HZ 7 224415 174645 NFX 32 307245 218700 Sun 28 306050 218141

HZ 8 224416 174646 NFX 33 307246 218701 Sun 29 306051 218142

HZ 9 224417 174647 NFX 34 307247 218702 Sun 30 306052 218143

HZ 10 224418 174648 NFX 35 307248 218703 Sun 31 306053 218144

HZ 11 224419 174649 NFX 36 307249 218704 Sun 32 306054 218145

HZ 12 224420 174650 NFX 37 307250 218705 Sun 33 306055 218146

HZ 13 224421 174651 NFX 38 307251 218706 Sun 34 306056 218147

HZ 14 224422 174652 NFX 42 307255 218710 Sun 35 306057 218148

HZ 15 231338 178085 NFX 43 307256 218711 Sun 36 306058 218149

HZ 16 231339 178086 NFX 44 307257 218712 Sun Frac.1 228059 176755

HZ 18 231340 178087 NFX 45 307258 218713 Sun Frac.2 228060 176756

HZ 19 224427 174657 NFX 46 307259 218714 Togo 1 228049 176769

HZ 20 224428 174658 NFX 47 307260 218715 Togo 2 228050 176770

HZ 21 224193 174659 NFX 48 307261 218716 Togo 3 228051 176771

Note: ‘(A)’ = Amended; ‘Frac.’=Fractional

Ownership of unpatented mining claims in the U.S. is in the name of the holder (locator), with

ownership of the minerals belonging to the United States of America, under the administration

of the U.S. Bureau of Land Management (BLM). Under the Mining Law of 1872, which

governs the location of unpatented mining claims on federal lands, the locator has the right to

explore, develop and mine minerals on unpatented mining claims without payments of

30

production royalties to the federal government. It should also be noted that in recent years there

have been U.S. Congressional efforts to change the 1872 mining law to include the provision

of federal production royalties. Currently, however, annual claim maintenance and filing fees

are the only federal encumbrances to unpatented mining claims.

The mining claims covering the northwest end of the property which includes the Ram deposit,

mill site and the tailings and waste rock storage facility were surveyed by Taylor Mountain

Survey; fractional claims were located to cover all fractions.

4.3 TENURE RIGHTS AND RISK FACTORS

To maintain the claims in good standing, FCC must pay annual claim maintenance and filing

fees to the BLM before September 1 of each calendar year. Other than maintenance and filing

fees, Micon is not aware of any other significant factors and risks that may affect access, title,

or the right or ability to perform work on the ICP property.

Presently, the ICP property is not subject to any royalties, other agreements and encumbrances.

Information relating to mineral claims was supplied by FCC. Micon has not carried out an

independent verification of land title and ownership for any of the above-mentioned claims.

31

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY

The following description is excerpted from the March 2015 PEA Technical Report by Samuel

Engineering Inc. with minor edits and additions.

5.1 ACCESSIBILITY

Vehicle access to the ICP is via a series of well-maintained, public-access gravel roads that

lead west from a point on paved Highway 93, approximately 6 miles south of Salmon, Idaho,

as shown in Figure 5.1. This gravel road leads to the Blackbird Mine, which is currently not

operating; however, the road is kept open year-round and a potential mining operation can

operate year-round. The total driving distance from Salmon to the ICP proposed mill site is

approximately 48 miles.

Figure 5.1

Idaho Project Site Access Roads

Source: Map supplied by FCC, 2016.

32

5.2 CLIMATE

The Natural Resources Conservation Service (NRCS) Morgan Creek SnoTel station is located

approximately 20 air miles south-southeast of the ICP at an elevation of 7,600-feet (NRCS,

2004). Based on 12 years of data (1991-2003), the average annual temperature at the station is

34.8 degrees Fahrenheit (ºF), with a low of –34.6ºF and a high of 89.4ºF. Based on 23 years of

data (1981-2004), annual precipitation is 24.4inches. About 60 percent of the precipitation

occurs as snow during the winter months (14.7 inches).

5.3 LOCAL RESOURCES AND INFRASTRUCTURE

Salmon, Idaho, is the nearest town and is located about 26 miles east of the property. The 2000

Census reported a population of about 3,120 people (www.city-data.com, 2005). Salmon is a

local supply and transportation center, with an airport paved with a 5,510 x 75-ft. airstrip at an

elevation of 4,044 ft. The nearest railroad is at Dubois, a smaller town 100 miles southeast of

Salmon. A 4 MW power line extends from Salmon to BMC’s Blackbird Mine site.

Although Salmon currently does not provide services for mining activities, it has functioned

in this manner for past mining activities at Noranda’s former Blackbird mine, and at Meridian

Gold’s former Beartrack gold mine. Salmon has, and can again, serve as a location for

personnel housing and a staging point for mine support services.

The area covered by the Idaho claims is sufficiently large to accommodate open pit and

underground operations, including ancillary installations.

5.4 PHYSIOGRAPHY

The ICP is located in the Salmon River Mountains of central Idaho, within the Northern Rocky

Mountain physiographic province. Major waterways in the area include the Salmon River and

Panther Creek. These waterways are located in the upper reaches of the Snake River Basin,

which drains to the Columbia River. The ICP is within the Panther Creek sub-basin of the

Salmon River. The project area contains flat-topped mountains and moderate to steep V-shaped

canyons, and covers an area ranging in elevation from 6,100 ft. to 8,100 ft. The area that may

potentially be affected by mining and mill operations is bounded by the divides of the streams

that generally drain the project area which are Bucktail Creek and Big Flat Creek. Bucktail

Creek drains into the South Fork of Big Deer Creek, which drains to Big Deer Creek, which

then drains to Panther Creek. Big Flat Creek drains directly into Panther Creek, which reports

to the Salmon River.

The terrain in the mine area is made up of slopes approaching 35% and cut by narrow valleys.

The mineralized material outcrops between elevations of 7,400 ft. and 7,800 ft., with most

facilities located at 6850 feet. Soils in the area are generally comprised of sandy loam averaging

5 ft. in depth, with frequent rock outcroppings. Bedrock exposure amounts to only about 1%

to 3% of the property area. Large boulder fields are found in many areas along the higher

mountain ridges.

33

During the summer of 2000 the Clear Creek Fire burned over 200,000 acres, including the area

of the ICP. The severity of the fire was high over most of the area, with all of the canopy cover

and most of the litter and duff burned off. A preliminary assessment indicates that the degree

of change that occurred was influenced by the various fuel loads, species, ladder fuels, canopy

closures, slope and aspect components interacting with fire weather conditions at the site. As

a consequence, typical mosaic patterns now prevail that are consistent with large fire behavior

in this type of ecosystem. Post-fire vegetation establishment in the project area in 2004 was

variable, with vegetation cover ranging from 30% to 80% depending on slope, aspect, fire

intensity and severity, soil type and post- fire seeding.

34

6.0 HISTORY

Sections 6.1 and 6.2 of this chapter have been excerpted from the March 2015 PEA Technical

Report by report by Samuel Engineering Inc. with some minor edits and additions.

6.1 DISCOVERY HISTORY

Copper mineralization in the Blackbird Creek area was discovered in 1892, and the area was

soon explored as both a copper and gold prospect. The area was first mined by Union Carbide

at the Haynes-Stellite Mine located south of the present FCC claim block, during World War

I. Union Carbide mined approximately 4,000 tons of cobalt-bearing ore before ceasing

operations, reportedly due to excessive mining costs. From 1938 to 1941, the Uncle Sam

Mining and Milling Company operated a mine at the south end of the present Blackbird mine

and reportedly mined about 3,600 tons of ore.

Calera Mining Company, a division of Howe Sound Company, developed and mined the

Blackbird deposit between 1943 and 1959 under a contract to supply cobalt to the U.S.

government. Calera mined approximately 1.74 M tons of ore grading 0.63% Co, 1.65% Cu,

and 0.03 oz. Au/ton during this period, accounting for the majority of production from the

district. Calera stopped mining when the government contract was terminated in 1960.

Reportedly, poor payment for cobalt from smelters hindered continued development of the

district, with minor exceptions.

Machinery Center Inc. mined 343,000 tons grading 0.36% Co and 0.64% Cu from the district

between 1963 and 1966, when Idaho Mining Company (owned by Hanna Mining Company)

purchased the property. Noranda optioned the property from Hanna in 1977 and carried out

extensive exploration, mine rehabilitation and metallurgical testing. In 1979 Noranda and

Hanna formed the Blackbird Mining Company (BMC) to develop the property. BMC

completed an internal feasibility study of their property at the time, including material from

the Sunshine deposit in 1982. BMC allowed perimeter claims to lapse in 1994, and FCC

restaked much of that ground. From 1995 to the present, FCC has completed surface

geochemical sampling and drilled 158 diamond drill holes on their ground.

6.2 HISTORICAL STUDY AND EVALUATION WORK

A prefeasibility-level Technical Report on the ICP property was prepared by MDA and filed

with SEDAR on October 31, 2006. Following this report, FCC decided to push forward with

further development work, drilling, a new resource model and metallurgical testwork.

In September 2007, a technical report on the ICP (the 2007 Technical Report), describing a

feasibility study, was filed on SEDAR (www.sedar.com). The 2007 Technical Report was

subsequently amended and refiled on SEDAR filed in May 2008. In August 2014, a Technical

Disclosure Review of Formation Metals Inc. by the British Columbia Securities Commission

determined that certain information in the 2007 Technical Report was deemed to be out of date

with respect to, among other things, commodity prices, capital cost estimates and operating

35

cost estimates and as such, was not to be relied upon. Subsequently, a new technical report

describing a PEA on the ICP was filed in March, 2015.

The United States Department of Agriculture Salmon Challis National Forest (the Forest

Service) issued a revised Record of Decision (the ROD) for the ICP in January 2009. The ROD

described the decision to approve a Mine Plan of Operations (MPO) for mining, milling and

concentrating mineralized material from the ICP. The ROD was subsequently affirmed by the

Forest Service in April 2009. As there are no significant changes to the mining methods,

milling and concentrating procedures from the previously filed Technical Report in

comparison to this 2015 PEA, this Plan of Operations at the ICP mine and mill remained

unchanged and the ROD remains in place. In December 2009, the Forest Service approved the

FCC’s MPO allowing for the commencement of ICP construction.

Construction on the ICP was planned in three stages; the first two have been completed. Stage

I construction commenced in January 2010 and concluded in April 2010. Stage I consisted of

timber clearing operations for the tailings waste storage facility (TWSF), topsoil stock pile

area, roads around the mill site and concentrator pads. Stage II construction comprised

primarily of earthworks preparation of all surface structures including mill and concentrator

pads, access and haul roads, TWSF and portal bench preparation, was dependent on securing

additional financing discussed below.

In October 2010, the FCC concluded a 5,727.5-ft. diamond drill program drilled in six holes

in a previously untested area on the project along the southern extension of the Ram deposit.

Data from this drill program was used for subsequent mine plan optimization studies. This

drilling extended the previously defined strike length of the Ram deposit an additional 14%

from 2,800 to 3,200 ft. The results of this drill program were incorporated into an updated

resource estimate for the ICP and form a part of the 2015 PEA report.

In March 2011, FCC announced that it had concluded an equity financing for gross proceeds

of CDN$80M. Proceeds of the financing were used to fund the continuation of engineering,

procurement and construction at the ICP (Stage II), for reclamation bonding requirements and

for general corporate purposes. Stage II construction commenced in July 2011 and concluded

in late 2012. Stage II construction also included mine site portal bench development,

geotechnical core drilling comprised of three HQ sized oriented core holes totaling 575 feet.

Drilling was completed in December 2011.

A Google Earth Image from September 2013 outlines the earthworks completed to date in

Figure 6.1 below.

36

Figure 6.1

Image of ICP Showing Mill Site and Completed Earthworks after Completion of Stages I and II

Construction

Source: Google Earth September 2013

The decision was made to defer Stage III construction of the ICP in May 2013. This final stage

of construction was to include the commencement of the underground development of mine

workings and the construction of all surface buildings including the mill, concentrator and

water treatment plant. The decision to defer construction was made in response to weakened

commodity prices and the enhanced adversity to risk by potential financiers in the prevailing

turbulent financial and commodity markets.

Falling commodity prices also affected Formation Metal Inc.’s ability to operate its Sunshine

Precious Metals refinery at a profit and in October 2013, the refinery was sold. The sale of the

refinery included land adjacent to the refinery building that was originally intended to house

the Cobalt Production Facility (CPF). As a consequence, a tradeoff study was undertaken to

determine the optimal location of the new CPF which is to be located along a railhead in

southern Idaho. Blackfoot, Idaho Falls, and Pocatello were all considered to be potential future

locations of the CPF.

Positive developments in the cobalt sector were realized in early 2014, fueled largely by

expansion projects for the development of electric vehicles, grid storage and the associated

projected explosive growth in the demand for rechargeable batteries requiring cobalt. In

August 2014, the price of cobalt metal attained a twenty-nine-month high of $16.00 per lb. In

response to these developments, in early 2014, FCC undertook a review of the cobalt chemicals

37

utilized in this sector and determined that pursuing the viability of producing cobalt chemicals

for the rechargeable battery sector was warranted. This developed into an in-house economic

analysis returning positive results and by August 2014, Requests for Proposals by independent

engineering firms to review the in-house engineering work was initiated and awarded to

Samuel Engineering, Inc. The results of these efforts culminated in the completion of the 2015

PEA and the initiation of feasibility level metallurgical testwork by Hazen Research Inc. of

Golden, CO, on ICP core and rejects. This metallurgical testwork was completed in Q2 2015.

A number of changes from the Technical Report were proposed, primarily at the CPF, with the

goal of maximizing the economic viability of producing cobalt chemicals for the rechargeable

battery sector. These changes included:

Increase in resources by including data from the 2010 drilling.

Inclusion of inferred material in resources in the 2015 PEA.

Re-estimated the resource and created a block model.

Redesign of a block mine model and mine schedule using the block model.

Reduction of development workings on the north end of the Ram deposit where

narrower, lower grade and isolated mineralization occurs.

Attention to dilution factors by utilizing slusher mining as opposed to LHD’s in smaller

width stopes.

Relocation of the CPF on a railhead.

Scalping of copper at the CPF.

The use of Cyanex 272 solvent extraction reagent at the CPF.

The production of battery grade cobalt sulphate heptahydrate chemicals at the CPF

resulting in the removal of numerous circuits required for high purity cobalt metal

production.

The production of copper sulphate as opposed to copper metal at the CPF.

The use of MgO instead of lime for neutralization resulting in less residue disposal at

the CPF.

The production of saleable MgSO4.

Inclusion of gold revenues (at 85% recovery).

6.3 HISTORICAL MINERAL RESOURCE ESTIMATES

Several mineral resource estimates have been prepared for the ICP prior to the 2016 estimate

of mineral resources presented herein. The reader is cautioned that these mineral resource

estimates are being treated as historical in nature and therefore are not confirmed to be NI 43-

101 compliant. They were prepared prior to the involvement of Micon and a Qualified Person

(QP) from Micon has not verified them as current. The relevance and reliability of the estimates

38

are not known. The 2005 and 2006 ICP estimates are classified using the categories set out in

the then current versions of the Canadian Institute of Mining, Metallurgy and Petroleum's CIM

Standards on Mineral Resources and Reserves, Definitions and Guidelines as required by NI

43-101. It is not known what reporting codes were used for the earlier estimates. It should also

be noted that all existing mineral resource estimates prepared prior to this report have since

been superseded by the 2016 Micon validated mineral resource estimate for the ICP, as

described in Sections 12 and 14.0 of this report.

6.3.1 1981 and 1997 ICP Mineral Resource Estimates

The resource estimates conducted in 1981 and 1997 pertain to the Sunshine and Sunshine East

deposits and not the Ram deposit which is the subject of this Technical Report.

6.3.2 1998 ICP Mineral Resource Estimate

The 1998 mineral resource estimate was conducted by FCC for the Ram, Sunshine and East

Sunshine deposits utilizing data from 92 drill holes, including historic and FCC’s drill

campaigns in 1995, 1996, and 1997. FCC performed the estimation by means of long-sectional

polygonal methods for the various stratiform mineralized horizons in each target area. The

resources are summarized in Table 6.1 at a cut-off grade of 0.20% Co.

Table 6.1

FCC’s 1998 ICP Mineral Resources at 0.20% Co Cut-off

Deposit Tons %Co %Cu Oz Au/t

Measured & Indicated (M & I)

Ram (I) 770,921 0.496 0.68 0.015

Sunshine (M & I) 245,554 0.965 0.47 0.022

East Sunshine (I) 100,466 0.422 0.94 0.014

Project Total (M & I) 1,116,941 0.592 0.657 0.016

Inferred

Ram 1,722,822 0.463 0.47 0.012

Sunshine 96,830 0.624 1.29 0.027

East Sunshine 430,748 0.404 1.06 0.017

Project Total (Inferred) 2,250,400 0.459 0.618 0.014

Caution: This resource is historical. It does not conform to NI 43-101 and the

CIM Definition Standards (2014); it is superseded by Micon’s resource estimate

presented in Section 14 of this report.

FCC’s resource estimates were independently audited by MDA in 1998, 1999, and again as

part of the 2001 MDA pre-feasibility study.

6.3.3 MDA 2001 Resource Estimate

FCC conducted additional drilling on the Ram deposit in 1999 and 2000 and this drilling was

included in the 2001 MDA resource estimate. Resource estimates for the Sunshine and East

Sunshine deposits remained unchanged from the 1998 estimate (see Table 6.1). The updated

39

Ram deposit resources were reported at a cut-off of 0.30% Co and are summarized in Table

6.2.

Table 6.2

MDA 2001 Ram Deposit Mineral Resource Estimate @ 0.30% Co Cut-off

Deposit Tons %Co %Cu Oz Au/t Lbs Co Lbs Cu Oz Au

Measured & Indicated (M & I) (000’s) (000’s)

Ram (M & I) 945,00 0.690 0.57 0.018 13,043 10,824 16,700

Inferred

Ram 1,807,000 0.644 0.47 0.021 23,298 17,128 38,560

Caution: This resource does not follow the CIM Definition Standards (2014) and is superseded by Micon’s resource

estimate presented in Section 14 of this report.


MDA updated the ICP mineral resources during 2005. For the Ram deposit, correlation of

horizons between drill holes and between cross sections were made based on a combination of

lithology, structure, style of mineralization, and grade.

The 2005 resource estimates for the Ram and Sunshine deposit were based on a long-section

polygonal method.

The 2005 combined Measured and Indicated Resources for the Sunshine and Ram deposits are

summarized in Table 6.3.

Table 6.3

MDA 2005 Resource Estimate (Ram & Sunshine Deposits) at 0.20% Co & 0.30% Co Cut-off

Ram & Sunshine Deposits Cut-off % Co Tons %Co %Cu Oz Au/t Avg TH Ft

Measured & Indicated (M & I) 0.30 1,895,400 0.667 0.598 0.016 7.7

Measured & Indicated (M & I) 0.20 2,282,300 0.596 0.561 0.014 6.1

Caution: This resource does not follow the CIM Definition Standards (2014) and is superseded by Micon’s

resource estimate presented in Section 14 of this report.


The MDA 2006 resource estimate was disclosed in the May 2008 Technical Report and

Feasibility Study prepared by Samuel Engineering Inc. (Kunter and Prenn, 2008). The

estimated resources are summarized in Table 6.4.

40

Table 6.4

MDA 2006 Resource Estimate at 0.30% Co Cut-off

Deposit Category Tons %Co %Cu Oz Au/t Avg TH Ft

Ram Measured & Indicated 2,393,700 0.631 0.651 0.016 8.2

Sunshine Measured & Indicated 260,700 0.604 0.327 0.013 3.8

Total Measured & Indicated 2,654,400 0.628 0.619 0.016 7.8

Caution: This resource does not follow the CIM Definition Standards (2014) and is superseded by

Micon’s resource estimate presented in Section 14 of this report.

6.4 PRODUCTION HISTORY

There has been no prior production from the ICP project and there are no historical mineral

reserve estimates.

41

7.0 GEOLOGICAL SETTING AND MINERALIZATION

Section 7.0 of this report relies heavily upon material contained in the March 2015 PEA

Technical Report by Samuel Engineering Inc. and MDA with minor edits/additions.

7.1 OVERVIEW

The ICP is located on the east side of the central Idaho Batholith Cretaceous-age granitic to

granodioritic rocks, hosted in Proterozoic-age sedimentary rock. The host sedimentary rocks

are on the southern flank of, and perhaps were part of, a large Proterozoic-age marine

sedimentary basin in which dominantly clastic sediments were deposited; now these

metamorphosed rocks are known as the Belt Supergroup and consist of dominantly quartzite

and argillite.

Unique to the Proterozoic rocks in this region, are cobalt-copper (Co-Cu) occurrences in the

Proterozoic age Apple Creek Formation of east-central Idaho. The Co-Cu mineralization at the

Blackbird Mine has been described as a type locality for this occurrence of stratiform Co-Cu

mineralization. The ICP is located adjacent to the former Co-Cu producing Blackbird Mine.

Work by the USGS published in Tysdal (2000), correlating the rocks of the Lemhi Range with

the rocks of the Salmon River Range, led to the recommended nomenclature of Apple Creek

Formation for the middle Yellow Jacket Formation, which includes the cobalt-bearing strata.

7.2 REGIONAL GEOLOGY

The regional geology is summarized in Figure 7.1. The ICP is situated in the Idaho Cobalt Belt,

a 30- to 35-mile long metallogenic district characterized by stratiform/tabular copper-cobalt

deposits. The deposits are hosted by a thick, dominantly clastic sequence of Middle Proterozoic

age sandwiched between late Proterozoic quartz monzonitic intrusions. The clastic sediments

were deposited in a large fault-bounded basin, probably as large submarine fan complexes

and/or deltaic aprons that were frequently “drowned” by continuing subsidence within the

basin. All significant copper-cobalt deposits and occurrences are found in the Proterozoic

Apple Creek Formation, which constitutes the base of this sequence. This formation was

originally correlated with Pritchard Formation metasediments of the Belt supergroup to the

north, its age being constrained by dates of 1.37 Ga for adamellites intruding the sequence and

1.7 Ga from mafic dykes and sills emplaced along the basin margin faults (Hughes, 1983).

The structure of the Apple Creek Formation is dominated by the regional rift structure. Cobalt-

copper-gold mineralization occurs along a northwest-southeast trending structure parallel to

and west of the central axis of the rift.

There is a series of northerly trending faults that are considered to represent initial growth

faults, reactivated by Laramide and younger events. The district has also been affected by

north-easterly structures of the Trans-Challis Fault Zone (Gow, 1995).

42

Figure 7.1

Regional Geology of the ICP


43

7.3 LOCAL GEOLOGY

The ICP is hosted in Proterozoic age meta-sediments found on the east side of the central Idaho

Batholith comprising granitic-to-granodioritic rocks. The local geology is summarized in

Figure 7.2. Figure 7.2

Local Geology of the ICP


Most of the following geologic discussion (except where otherwise indicated) is summarized

from an internal report dated April 1998 and entitled “Report on the Reserve/Resource

Estimates for Sunshine Lode, East Sunshine, and Ram Prospects, Sunshine Property, Idaho,

USA” by FCC field staff.

44

7.3.1 Lithology and Stratigraphy

The Idaho Cobalt Belt represents a distinct district dominated by stratabound cobalt + copper

± gold mineralization, with a remobilized constituent. The district is underlain by strata of the

middle Proterozoic-age Apple Creek Formation, which is an upward-thickening, upward-

coarsening clastic sequence at least 49,000 feet thick (Nash, 1989) that represents a major

basin-filling episode (Connor, 1990) and was formerly considered part of the Yellow Jacket

Formation.

Detailed work by Noranda geologists and the USGS showed that the Apple Creek can be

divided into three units. The lower unit of the Apple Creek Formation is over 15,000 feet thick

and consists mainly of argillite and siltite, with lesser occurrences of fine-grained quartzite and

carbonates. Graded bedding and planar to wavy laminae are common in the lower unit, which

is locally metamorphosed to phyllite. The middle unit of the Apple Creek Formation is up to

3,600 feet thick and comprises several upward-coarsening sequences of argillite, siltite, and

quartzite, with distinctive biotite-rich interbeds (Nash, 1989) that generally have a direct

correlation to mineralization. The middle unit hosts the majority of the known cobalt, copper

and gold occurrences in the Idaho Cobalt Belt. The upper unit exceeds 9,800 feet in thickness

and is predominantly composed of thin- to thick bedded, very fine- to fine-grained quartzite

(Connor, 1990).

Mafic tuffs within the Apple Creek Formation are the oldest igneous rocks exposed in the

Sunshine-Blackpine district. They are accompanied by felsic tuffs and carbonatitic tuffs. Some

mafic dikes and sills intrude the Apple Creek Formation and may be comagmatic with the

mafic tuff beds. Several small lamproitic diatremes may also be coeval with mafic volcanism

(Gow, 1995).

The Apple Creek Formation has undergone varying degrees of regional metamorphism,

ranging from greenschist facies in the southern part of the district to amphibolite grade facies

in the northern part of the district. Several types of mafic dikes and sills, ranging from 3 ft. to

100 ft. thick, intrude the Apple Creek Formation and are interpreted as feeders to the exhalative

mafic tuffs, which are most abundant in areas of intrusive activity.

7.3.2 Structural Geology of the Deposits

The dominant structures in the area are steep, north- to northwest-trending normal faults and

shear zones. The prominent White Ledge Shear, which displays substantial apparent strike-slip

movement, marks the western extent of the mafic strata and associated stratiform

mineralization in the project area (Nash, 1989).

Noranda Exploration Inc. interpreted the Sunshine stratigraphy as having been folded into a

tight syncline about a northerly-plunging axis (Daggett and Baer, 1981). Small-scale fold

hinges and transposed bedding visible in the Sunshine Trench indicate parasitic folding and

locally severe deformation. Large-scale transposition faults roughly parallel the axial plane of

the Sunshine syncline.

45

7.3.3 Ram Deposit Stratigraphy

Stratigraphy in the Ram deposit area is predominantly medium- to fine-grained quartzite

metamorphosed to upper greenschist to amphibolite facies. Stratigraphically, the Ram deposit

is subdivided into three zones: Hanging-wall, Main and Footwall zones, with each zone

containing distinct mineralized horizons. Typical cross sections are provided in Section 10 that

deals with drilling results.

7.3.3.1 Hanging-wall Zone

FCC subdivided the Ram Hanging-wall zone into three lithologic packages. The upper hanging

wall contains medium- to coarse-grained, locally poorly bedded to well-bedded quartzite.

Occurrences of biotitic tuffaceous exhalite (BTE) are generally restricted to discontinuous,

irregular coarse-grained garnetiferous pods (interpreted locally as diapirs). The middle hanging

wall is dominated by medium grained, generally well-bedded quartzite that is locally

conformably interbedded with chloritic/biotitic cobaltiferous tuffaceous exhalites. The lower

hanging wall includes medium- to coarse-grained quartzite with poorly defined, chaotic

bedding. BTE material is restricted to sporadic, irregular diapirs.

The current resource model described in Section 14.0 contains four additional hanging wall

horizons that occur above or between and often coalesce with the primary three hanging wall

horizons. Each of the four is limited in spatial extent.

7.3.3.2 Main Zone

The Main zone is dominated by fine- to medium-grained, thin- to medium-bedded quartzites

that are interbedded with biotitic and chloritic tuffaceous exhalites and local siliceous

tuffaceous exhalites (STE). Mineralization in the Ram Main zone is generally found within a

confined stratigraphic package containing three, closely spaced, stratiform horizons, of

variable thickness and continuity, which strike between 340° and 355° and dip between 50°

and 55° to the northeast. The three mineralized horizons have been coded from upper to lower

in the geologic model as 3021, 3022 and 3023 horizons. The 3023 horizon is the lowest

member of the main zone and is the thickest and most continuous horizon. The main zone is

up to 21 feet in true thickness.

The Main zone horizons contain fine- to coarse-grained disseminations, bands, blebs, and

stringers of cobaltite, chalcopyrite, and minor pyrite. This mineralization is dominantly

concordant with bedding, but locally has been remobilized into thin quartz veins (i.e., ‘sweat

veins’) and/or crosscutting structures. The main zone represents the bulk of the potentially

economic mineralization identified in the Ram deposit to date.

7.3.3.3 Footwall Zone

The Footwall zone was subdivided into two rock packages. The upper footwall is characterized

by poorly to well-bedded silty quartzite, often intercalated with chloritic and biotitic tuffaceous

46

exhalite. Frequently distorted bedding (soft sediment deformation) and a lack of tuffaceous

exhalite differentiate the lower footwall from the upper.

Table 7.1 summarizes the stratigraphy of the Ram Deposit.

Table 7.1

Summary of the Stratigraphy of the Ram Deposit

Component Description

Upper Hanging

Wall Medium to coarse grained quartzites, locally poorly bedded to well bedded

BTE is restricted to irregular, coarsely garnetiferous diapirs

Middle Hanging

Wall Medium grained quartzites interbedded with locally conformable cobaltiferous

chloritic/biotitic exhalites, generally well bedded

Lower Hanging

Wall Medium to coarse grained quartzite, poorly defined locally chaotic bedding

Local clastics

Sporadic, irregular diapirs of biotitic exhalite, often cobaltiferous

Main Zone

Mineralization Fine to medium grained, poorly bedded but with locally well-developed thin to

medium bedding • generally three conformable cobaltiferous horizons, best developed

down dip

3021 horizon Not always well defined • often comprised of clastic horizon with biotitic matrix and

considerable chalcopyrite and minor pyrite, on section 0+00 comprised of fine grained

cobaltite fracture fill associated with minor chloritic to biotitic BTE in STE matrix. •

biotite becomes dominant down dip • not well developed up dip on northern sections

3022 horizon Not always well defined

Often comprised of disseminated cobaltite in biotitic gangue

Best developed down dip

3023 horizon Variably comprised of disseminated and/or banded cobaltite in chloritic BTE, or

fracture filling cobaltite in STE with attendant chloritic or biotitic BTE

In general, chloritic component increases up dip and biotite increases down dip

STE increases down dip

Chalcopyrite and pyrite appear to increase down dip • foot wall contact is best

developed down dip

A distinct, black, biotitic MDS is encountered near the lower horizon in all

intersections

Upper Footwall Silty quartzite, intercalated with BTE, thin to medium bedded, poorly to well-defined •

Chloritic BTE component absent or restricted to horizons down dip

Footwall

horizons Commonly characterized by biotitic matrix with chloritic overprint

Locally contains STE

Commonly contains chalcopyrite and locally pyrite and minor pyrrhotite

In general, chloritic component increase up dip

In general, biotitic component and sulphides increase down dip and to the north

Lower Footwall Silty quartzite, thin to medium bedded, poorly to well defined, often distorted

Frequently soft sediment textures

Locally abundant MDS, frequently calcareous • occasional biotitic garnetiferous bands

Note: MDS = Mafic dyke/sill

7.3.3.4 Faulting

A north-trending vertical to steeply east dipping normal fault is evident in drill core. The fault

cuts the main mineralized zone in the south on section 2+00 north and diverges to the north.

47

The mineralized horizons were found in drill holes R97-02 and R97-03 west of the fault. The

down dropped side of the fault appears to be around 40 feet lower. Shearing is locally common

in the footwall as well as in the upper hanging wall. Small-scale folds are evident in drill core

(FCC 1998 resource report).

7.3.4 Sunshine Deposit Stratigraphy

Stratigraphy, including the BTE horizons, strikes north northwest and dips moderately to

steeply to the east-northeast. Individual sulphide-bearing beds may not be continuous over a

distance of a few hundred feet, but generally, the overall mineralized zones within the BTE

horizons can be traced along strike for over 1,500 feet.

The description that follows is copied from the 1998 resource report that was completed by

FCC:

The Sunshine Lode’s Main Zone is comprised of fine- to medium-grained metaquartzite

interbedded with siltite and mafic sequences. The mafic sequences, comprised of green biotite

and lesser chlorite, have been interpreted to be metamorphosed tuffs or exhalites (BTE) (Clark,

L.A., 1995). Portions of the mafic sequence contain significant amounts of chert of exhalative

origin (STE) (Clark, L.A., 1995).

The hanging-wall stratigraphy is dominated by upward-coarsening and thickening quartzite.

In the lower hanging wall, quartzite is intercalated with local siltite and minor mafic sequences

(BTE), while in the upper hanging wall quartzite contains little siltite and no mafic sequences.

The footwall stratigraphy is dominated by a thick sequence of monotonous siltite or pelite with

minor interbedded sandy units. Mafic sequences are rare and cannot be correlated except

locally. Shearing is prevalent within this package.

The boundary between the footwall and the main zone is defined by a sedimentary interface

based on grain size, indicating a change between shallow and deeper water.

Concordant to sub-concordant discontinuous quartz veins are found throughout the Sunshine

Lode’s stratigraphy. These are diagenetic in origin, and while they occasionally carry grade,

they are not traceable for any appreciable distance along strike or down dip.

Folding, at least locally and on the bedding scale, has been noted within the (drill) core from

changes in bedding attitudes and fold noses. This may lend support to the idea that the horizons

are folded repetitions. However, no definitive evidence of overturning could be documented

in the core.

The Sunshine Lode mineralization appears to be cut by a number of discontinuous, shallow to

moderate, west-dipping, dip-slip faults/shears. In addition, drilling has revealed a number of

discontinuous, crosscutting tectonic breccias, which may affect the continuity of the Sunshine

Lode, at least locally.

48

The north-trending, steeply west-dipping, Green Dyke fault, which parallels the Sunshine Lode

for much of its strike length, may truncate the mineralization down dip and to the south.

Drilling below the fault has been limited, and some of the holes may not have reached the

mineralized horizons. Two Noranda drill holes, 80-03A and 80-13A, which do penetrate below

the fault, intersected a core length of 2.30 ft. of 0.320% cobalt, 0.08% copper and 0.003 oz

gold/ton and 4.00 ft. of 0.217% cobalt, 0.21% copper 0.003 oz gold/ton respectively. These

holes suggest that higher-grade pods of mineralization may remain undiscovered below the

fault. Neither a sense of movement nor a displacement has been determined for this fault.

The Sunshine Lode’s mineralized zone is found within a confined stratigraphic section that

contains a main mineralized horizon (1003), a lower footwall horizon (1001) and an upper

hanging-wall horizon (1007). Although the mineralized zone is continuous along strike, the

individual horizons do not always display good continuity along strike or down dip. The

footwall and hanging-wall horizons attenuate rapidly both along strike and down dip.

However, within the main horizon and hanging-wall horizon, tabular deposits of mineralization

with sufficient grade and size exist, which should be mineable. These deposits appear to have

their long axis down plunge towards north.

The stratabound mineralization revealed by the drilling consists of fine- to medium-grained

disseminations, blebs and stringers of cobaltite and minor chalcopyrite and pyrite. Two types

of (mineralization) occur within the Sunshine Lode, fine- to coarse-grained cobaltite within

siliceous gangue and fine-grained cobaltite within micaceous gangue. The micas are black

biotite, green biotite and chlorite. The horizons are typically composed of both mineralized

material types and are hosted by medium grained biotite rich quartzites.

7.3.5 Alteration

The Apple Creek Formation has been subjected to varying degrees of regional metamorphism

resulting in the southern part of the district displaying greenschist facies while the northern

part is dominated by amphibolite grade facies. On a broad scale, alteration related to

mineralizing events is manifested by the presence of tourmaline and ankerite in the middle unit

that hosts the mineralized zones. However, these alteration minerals can be found up to several

thousands of feet from the nearest known sulphide occurrences, and thus, do not provide any

reliable indications of proximity to ore targets. In places, ankerite changes to disseminated

siderite. Silicification and chloritization have been noted within the mineralized zones but

chlorite-rich rocks may be found as much as several hundreds of feet from known

mineralization. Alteration at the Idaho Cobalt Project has been likened by FCC personnel to

that found at the nearby Blackbird deposits which has been described as being strata-bound

and coincident with biotite and intercalated rocks with the alteration zoning consisting of

pyrite-siderite-quartz-muscovite in the core zone and grading outward into quartz-muscovite-

lesser pyrite. Potassic alteration has enhanced biotite crystallization across the entire ore zone.

49

7.4 MINERALIZATION

7.4.1 Global Overview

A number of significant stratiform/tabular cobalt-copper-gold deposits and prospects define

the Idaho Cobalt belt. As far as can be determined at this point, they are associated with two

or more distinctive, regional stratigraphic horizons within the Apple Creek Formation that are

distinguished by diagnostic Fe minerals. In the Blackbird area, the mineralized sequence is

characterized by the presence of biotite-rich beds often referred to as “biotitic” within a

sequence of up to 3,000 feet of interbedded quartzite, siltite and argillite. Approximately 10

miles to the southeast, probably within the same stratigraphic sequence, FCC has been

exploring stratiform copper-cobalt mineralization at their Blackpine project.

Three types of cobalt-copper-gold occurrences have been reported in the Idaho Cobalt Belt

(Nash, 1989, reported in Pegg, 1997):

Type 1: Cobalt-copper-arsenic rich deposits of the Blackbird Mine type. Generally, these

contain approximately equal amounts of cobalt and copper, with variable amounts of gold and

pyrite. The dominant minerals include cobaltite (CoAsS) and chalcopyrite (CuFeS2). The

cobaltite accounts for nearly all of the arsenic content in these occurrences. This syngenetic

and stratabound mineralization is closely associated with mafic sequences of the Apple Creek

Formation. The deposits are found in tabular form. Examples of these types of deposits include

the Blackbird Mine and the mineralized zones found within FCC’s Sunshine and Ram deposits.

Type 2: Cobaltiferous-pyrite-magnetite deposits with a variable chalcopyrite and low arsenic

content. These occurrences are hosted by fine-grained metasediments from the lower unit of

the Apple Creek Formation. Mineralization is stratabound, locally stratiform and is found

within syn-sedimentary soft sediment structures. The deposits are found in the area of Iron

Creek, approximately 17 miles southeast of the Blackbird Mine.

Type 3: Cobaltiferous, tourmaline-cemented breccias. These are relatively common in the

lower unit of the Apple Creek Formation, especially south and east of the Blackbird Mine.

Only a few of these, apparently, contain in excess of 0.1% cobalt.

7.4.2 ICP Mineralization

Mineralization at the ICP is Type 1 characterized as syngenetic, stratiform/tabular exhalative

deposits within, or closely associated with, the mafic sequences of the Apple Creek Formation.

This mineralization is dominantly bedding concordant and the deposits range from nearly

massive to disseminated. Some crosscutting mineralization is present that may be in feeder

zones to the stratiform mineralization or may be due to remobilization locally into fracture

quartz veins and/or crosscutting structures.

Dominant minerals include cobaltite (CoAsS) and chalcopyrite (CuFeS2), with lesser, variable

occurrences of gold. Other minerals present in small quantities are pyrite (FeS2), pyrrhotite

50

(FeS), arsenopyrite (FeAsS), linnaeite ((Co Ni)3S4), loellingite (FeAs2), safflorite (CoFeAs2),

enargite (Cu3AsS4) and marcasite (FeS2).

Recently, rare-earth minerals have been identified in samples from the deposit as monazite,

xenotime and allanite. At this time, these minerals have not been considered for potential

recovery as by-products of the Co-(Cu-Au).

The Ram is the largest and best-known deposit in the ICP area. It consists of a Hanging-wall

Zone with 3 primary and 4 minor horizons, a Main Zone comprising 3 horizons, and a Footwall

Zone with 3 horizons (Figure 10.3). These sub-parallel horizons generally strike N15oW and

dip 50o – 60o to the northeast. Most of the significant Co mineralization is associated with

exhalative lithologies i.e. biotitic tuffaceous exhalate (BTE), siliceous tuffaceous exhalate

(STE), and quartzite with impregnations of biotitic tuffaceous exhalate (QTZ/BTE) or siliceous

tuffaceous exhalate (QTZ/STE).

The Sunshine/East Sunshine deposit is FCC’s second best known deposit within the ICP area

and is located about a mile south of the Ram deposit. Mineralized zones are typically multiple,

stacked sulphide-bearing beds. Individual mineralized beds or horizons are intimately

associated with biotite-rich tuffaceous exhalative (BTE) horizons. An increase in silica content

generally indicates an increase in cobalt, copper and gold grades.

51

8.0 DEPOSIT TYPES

8.1 PRE-2005 CONCEPTIONS

Geoscientific work/observations prior to 2005 suggested a sedimentary exhalative deposit

class for the ICP deposits as reflected in the following descriptions excerpted from the March

2015 PEA Technical Report by report by Samuel Engineering Inc.:

The deposits comprising the ICP belong to a class of deposits variably described as “Blackbird

Co-Cu” (Evans et. al., 1986) or “Blackbird Sediment-hosted Cu-Co” (Hõy, 1995). The ICP

lies in the type locality for these deposits and includes some of the type deposits.

According to Evans et. al. (1986), “These deposits are stratabound iron-, cobalt-, copper-, and

arsenic-rich sulphide mineral accumulations in nearly carbonate-free argillite/siltite couplets

and quartzites”.

Hoy (1995) suggested the following “associated deposit types: Possibly Besshi volcanogenic

massive sulphide deposits, Fe formations, base metal veins, tourmaline breccias.”

8.2 POST-2005 CONCEPTIONS

Recent geoscientific work and observations suggests an iron oxide-copper-gold (IOCG)

deposit class with a magmatic-hydrothermal origin for the ICP deposits. The following is an

excerpt from the abstract of a paper by Slack J. F. (2006) – see references under Section 28.

“Analysis of 11 samples of strata-bound Co-Cu-Au ore from the Blackbird district in Idaho

shows previously unknown high concentrations of rare earth elements (REE) and Y, averaging

0.53 wt percent ∑REE + Y oxides. Scanning electron microscopy indicates REE and Y

residence in monazite, xenotime, and allanite that form complex intergrowths with cobaltite,

suggesting coeval Co and REE + Y mineralization during the Mesoproterozoic. Occurrence of

high REE and Y concentrations in the Blackbird ores, together with previously documented

saline-rich fluid inclusions and Cl-rich biotite, suggest that these are not volcanogenic massive

sulphide or sedimentary exhalative deposits but instead are iron oxide-copper-gold (IOCG)

deposits”.

52

9.0 EXPLORATION

Most of the exploration work on the ICP was conducted between 1995 and 1998. This work is

described in detail in the FCC field staff report of 1998 and is summarized below.

9.1 PROGRAMS

9.1.1 1995-1996 Campaign

In 1995, soil sampling of selected areas was conducted on lines spaced 200 ft. and 400 ft. apart,

with samples collected at intervals of 100 ft. along the lines. This program discovered the

southern end of the previously unknown Ram target.

In 1996, the soil grid was extended north and soil samples were collected on lines spaced 200

ft. apart with samples collected at 25 ft. intervals along the lines. Some infill samples were

collected from the 1995 soil grid. Other parts of the grid were also extended and sampled on

25 ft. intervals where it was deemed warranted.

A total of 8,427 soil samples were collected during the 1995/1996 campaign.

Other exploration activities conducted during 1995/1996 include surface geological mapping

at a scale of 1 in. to 100 ft., mapping of old trenches and prospect pits, and collection of 979

surface rock samples including those from trenches.

9.1.2 1997 Campaign

The Ram soil grid was extended northward, with the collection of an additional 95 soil samples;

concurrently, the north and south extensions of the Ram prospect were geologically mapped.

In the same year, FCC built 3,100 ft. of benched drill road into the Ram zone; the road was

laid out to cross the Ram soil geochemical anomaly, in order to facilitate trenching. Three

trenches, 623 ft. long in aggregate, were excavated within the “prism” of the road; the trenches

were mapped and 83 rock samples were collected. The newly opened 6930 drift was mapped,

and 163 rock samples were collected.

For a topographic base, FCC had a five-foot contour map of the project area, produced

photogrammetrically, using aerial photography.

9.1.3 1998-2001 Campaign

Permitting baseline studies were initiated.

53

9.1.4 2002-2006 Campaign

Various baseline studies were completed in support of project activities. The Plan of

Operations (POO) and the United States (USFS) Environmental Impact Statement (EIS) were

also completed. An updated POO was submitted in April 2006;

9.1.5 2007-2016 Campaign

No exploration work other than drilling was carried out.

9.2 EXPLORATION RESULTS

The surface geological and geochemical work were important contributors to the discovery

and expansion of the Ram deposit both in the northerly and southerly directions. Whilst both

soil and rock chip samples are not representative, they serve primarily to detect mineralization

for further investigation by trenching and ultimately drilling.

54

10.0 DRILLING

10.1 DRILLING CAMPAIGNS

The ICP drilling campaigns are summarized in Table 10.1. Total drilling by FCC is 158 holes

for 103,185.5 ft. completed between 1995 and 2010.

Table 10.1

ICP Drilling Campaigns

Year Drilled Operator Deposit Number Feet

1959 Calera Mininig Company Sunshine 3 982

1979 – 1981 Blackbird Mining Company Sunshine 29 17,826.0

Noranda

1995 – 1996 Formation Capital Sunshine 48 29,144.0

1995 – 1996 Formation Capital East Sunshine 24 14,723.5

1997 Formation Capital Ram 20 12,045.0


2000* Formation Capital Ram 8 2,613.0





Totals 104 62,675.5

Totals 86 60,300.0

Grand Total Ram+Sunshine 190 121,993.5

*Metallurgical Test holes – Not used in Grade Model

The Ram deposit has been tested by 86 diamond drill holes totalling 60,300 ft. drilled in 1997

through 2010 by FCC. Although drilling has been intermittent over the years, there has been

continuity over the campaigns. The Ram deposit comprises several sub-parallel horizons which

generally strike N15°W and dip 50°-60° to the northeast and were drill tested to depths of

1,200 ft. vertically. The Main Zone horizons, which are the most extensive, were drill tested

over 3,000 ft. in strike extent, 500 ft.-1,200 ft. in vertical extent, and have true thicknesses that

average about 8 ft. (true thicknesses range from less than 3 ft. to greater than 20 ft. for horizon

3023). Figure 10.1 shows locations of the drill collars, the surface projection of drill-hole traces

(azimuths), and the current resource outline for the Ram deposit.

The Sunshine deposit is located about a mile (1.6 km) due south of the Ram deposit (Figure

7.2). It consists of multiple, stacked sulphide-bearing beds of limited strike length. Individual

mineralized beds or horizons range in thickness from inches to several feet and are associated

with biotite-rich tuffaceous exhalative (BTE) horizons. The deposit horizons strike north-

northwest and dip moderately to steeply to the east-northeast.

The resources considered in the current Technical Report are those of the Ram deposit only.

The Sunshine and other deposits within the property represent additional potential for the ICP

resources. All holes drilled on the Ram deposit are diamond core holes.

55

Figure 10.1

Ram Deposit Drill Hole Locations

Source: Map supplied by FCC, 2016; modified by Micon, 2017.

56

10.2 FCC DRILLING PROCEDURES

The following description has been excerpted from the March 2015 PEA Technical Report by

Samuel Engineering Inc. and is based on MDA’s observations from 1998 to 2010.

All drill data was obtained by core drilling, with the exception of reverse circulation pre-collars

for the holes completed by FCC in 2000 to obtain metallurgical samples. Exploration holes

were drilled with either NQ- or HQ-size core; the metallurgical holes were drilled with PQ-

size core. NQ, HQ and PQ core have diameters of 1.875 inches (47.6mm), 2.500 inches

(63.5mm) and 3.345 inches (85.0mm), respectively.

FCC routinely logged the drill core in considerable detail, with particular emphasis placed on

mineralized intervals.

The collars of all drill holes were located using tight chain and compass from the nearest known

point. Most of the pre-1998 drill-hole collar locations were resurveyed by Harper-Leavitt

Engineering Inc., using a transit (1998 report by FCC Staff). Collar locations for the 2010 drill

holes were professionally surveyed by Taylor Mountain Surveying, of Salmon, Idaho, using a

combination of Global Positioning Systems and conventional survey methods.

A single-shot, Sperry Sun instrument was used for down-hole surveys to check the drill-hole

orientations. Down-hole surveys were done every 150 feet in the hole.

Drilling was conducted as angle holes oriented approximately normal to the strike of the

mineralized horizons, and crosscutting mineralized horizons at appropriate angles that allowed

true thicknesses of mineralization to be determined.

It is MDA’s opinion that FCC’s drilling methods used at the Ram Deposit follow industry

standard procedures, and are appropriate methods to adequately interpret the geology and

mineralized zones used in the resource model.

10.3 MICON OBSERVATIONS DURING SITE VISIT/COMMENTS

Micon’s site visit was made long after FCC’s drilling campaigns. However, detailed

inspection of stored drill cores, drill core logs, reports and transcripts prepared by FCC and

MDA confirm that all the drill holes were either NQ- or HQ-size core with diameters of 1.875

inches (47.6 mm) and 2,5 inches (63.5 mm), respectively. All drill cores and records of

drilling/sampling (Figure 10.2) are kept in a neat condition.

57

Figure 10.2

FCC’s Resident Geologist Displaying 1996 Drill Cores/Sampling Records during Micon Visit

Source: Photo taken by Micon, July 2016

10.4 DRILLING RESULTS

Drill hole logging, sampling and assay results have confirmed the following:

The Ram deposit consists of a Hanging-wall Zone with 3 primary and 4 minor horizons,

a Main Zone comprising 3 horizons, and a Footwall Zone with 3 horizons (Figure 10.3).

These sub-parallel horizons generally strike N15oW and dip 50o – 60o to the northeast.

The mineralized zones are tabular/stratiform as shown in Figure 10.3.

Most of the significant Co mineralization is associated with exhalative lithologies i.e.

biotitic tuffaceous exhalate (BTE), siliceous tuffaceous exhalate (STE), and quartzite

with impregnations of biotitic tuffaceous exhalate (QTZ/BTE) or siliceous tuffaceous

exhalate (QTZ/STE).

True thickness of the mineralized horizons averages is about 8 ft.; the range is from

less than 2 ft. to over 20 ft.

There is a weak to fair correlation between Co and Au but none between Co and Cu.

58

Figure 10.3

Typical Cross Section through the Ram Deposit

Source: Micon 2017 – Generated from resource database

10.5 MICON COMMENTS

Micon notes that all the drilling on the Ram deposit has been conducted by FCC and thus

protocols pertaining to the exploration history of the deposit and database build-up have

progressive continuity. The core sizes used yield representative samples and also minimize

core loss in bad ground. However, in a few mineralized areas affected by faults and shears, up

to 10% of core was lost. The frequency of such losses is insignificant to have a material impact

on the assay database.

Drill hole logs produced by FCC are very detailed and include all essential information, i.e.,

drill hole survey information, core losses, rock quality designation, lithology, structure,

alteration, and sampling.

59

FCC’s drilling protocols appear to have been in line with the best practice guidelines of the

prevailing CIM Standards over the drill campaign periods.

There are no drilling, sampling or recovery factors that could materially impact the accuracy

and reliability of the assay results.

60

11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 SAMPLE PREPARATION

11.1.1 Sample Preparation at Site

Sample lengths/intervals were delineated based on lithological, alteration and mineralogical

changes. FCC’s sample lengths ranged from 0.50 ft. to 5 ft. throughout all its drilling

campaigns. Typically, several sample intervals define the mineralization for any single

intercept of the various mineralized horizons. Prospective/anomalous zones were bracketed by

taking two or more samples on the margins.

Once the logging was complete, the drill core selected for sampling was sawn lengthwise into

symmetrical halves resulting in two equally representative samples. One-half of the drill core

was placed in a plastic sample bag with a sample identification tag before being sealed. The

other half of the drill core was returned to its original position in the core box and the

corresponding tag for each sample interval was placed at the end of the sample position in the

core box.

Quality control is achieved by inserting one barren control sample (blank) and certified

reference materials (CRMs) at regular intervals into the sample stream for each batch of core

samples.

Other than the insertion of control samples, there is no other action taken at site.

Sample reference sheets summarizing all the samples taken from each hole are provided during

the core cutting process. The sheets are used to follow along for quality control samples and

for preparing for the assay requisition and shipment forms. When a standard is encountered,

the sticker is removed from the standard packet and placed on the sample sheet with its

associated sample.

11.1.2 Laboratory Sample Preparation

Once at the laboratory, the samples are entered into the internal system. Samples are prepared

by drying, if necessary, then the entire sample is crushed in its entirety to ≥70% at <2 mm,

riffle split to obtain a 250 g sub-sample, which was pulverized to ≥ 85% at < 75 microns.

11.2 ANALYSES

FCC included cobalt, copper and gold assaying as part of their routine analytical procedure. In

addition, multi-element geochemical analyses were completed on nearly all of FCC’s drill

holes at the ICP.

FCC’s sample analyses through 2006 were performed by Chemex Labs, Inc., of Sparks,

Nevada, and Vancouver, British Columbia, and by Bondar Clegg Laboratories, Inc. (USA), of

61

Reno, Nevada, and Bondar Clegg Laboratories, Inc. (Canada), of Vancouver, British

Columbia. EcoTech Laboratories Ltd. of Kamloops, British Columbia, completed additional

check-sample analyses in 1996.

Cobalt and copper analyses for drill samples up through the 2000 drilling were done by 4-acid

(HNO3- HClO4-HF-HCl) digestion and an atomic absorption (AA) finish; gold was analyzed

by 30-gram fire assay followed by an AA finish. Cobalt and copper analyses for the 2004

through 2006 drill samples were done by aqua regia digestion and an atomic absorption (AA)

finish; gold was again analyzed by 30-gram fire assay followed by an AA finish. Multi-element

geochemical analyses for all drill campaigns were performed using aqua regia digestion

followed by induction-coupled plasma atomic-emission spectrometry (ICP-AES). These are

all industry standard analytical techniques appropriate for the types of rocks and mineralization

at the ICP.

Chemex Labs, Inc., which became ALS Chemex and subsequently ALS Global, holds ISO

9002:1994 certification at its North American and Peruvian laboratories and ISO 9001:2000

certification in North America. ALS Global is the successor to Chemex and Bondar Clegg, the

laboratories that did most of FCC’s analyses. Neither Micon nor MDA has determined the date

that ALS Global or its predecessors first obtained ISO 9002 certification, but it is probable that

much of the work for FCC was done before that date.

FCC’s sample analyses in 2010 were performed by ALS Minerals, a division of ALS Global.

Samples were crushed in their entirety to ≥70% at <2mm, riffle split to obtain a 250g sub-

sample, which was pulverized to ≥ 85% at < 75 microns. Analytical techniques similar to those

used prior to 2010 were employed, including aqua regia digestion and AA or ICP-AES finish

for cobalt and copper, and 30-gram fire assay with AA finish for gold. Multi-element

geochemical analyses were performed using lithium metaborate fusion, acid digest and ICP-

AES-Mass Spectrometry. Duplicate samples for verification purposes were analyzed at ACT

Labs of Ontario, Canada and were analyzed for cobalt and copper by sodium peroxide fusion

and ICP-AES finish, and for gold by 30-gram fire assay with AA finish.

All the laboratories involved in the analyses of FCC’s samples are independent of the issuer.

11.3 SECURITY

All activities pertaining to data collection, i.e. sampling, insertion of control samples,

packaging and transportation, were/are conducted under the direct supervision of the project

manager.

FCC’s core and sample security measures were typical for exploration projects in North

America at the time the work was done. The core was received at the drill by an employee of

FCC, and taken to the company’s facility in Salmon for processing.

That facility is a warehouse-like building (Figure 11.1) with lockable doors. Sawed core was

placed in labeled sample bags that were closed with wire ties. An employee of the analytical

62

laboratory picked up the samples at FCC’s facility. Thus, the core has been under FCC’s

control from receipt at the drill, and the parts of core not used for the analytical samples

remained under FCC’s control. The samples were under FCC’s control from the drill to the

core sawing facility, and under the laboratory’s control after leaving FCC’s facility.

Figure 11.1

FCC Core Storage Facility in Salmon

Source: Photo taken by Micon, 2016.

11.4 QUALITY CONTROL/ASSURANCE (QA/QC)

11.4.1 MDA Verification

MDA examined FCC’s data related to QA/QC in 1998 and established that the assays of the

check samples, blanks and standards were in good agreement with the expected values. MDA

also examined the 1999 Ram drilling QA/QC and a further check on assay QA/QC data was

completed in 2004. MDA’s conclusion was “Overall, FCC has demonstrated diligence in

monitoring check assays and standards and blanks results, which is critical to the maintenance

of an accurate database”. In addition to these checks, MDA independently selected 10 samples

from the 2005-2006 drilling program and sent them to ACME laboratories for check assaying

from which they obtained a good agreement between the original assays and the check assays.

63

11.4.2 Micon Verification

Micon has noted that FCC used both blanks and standards in its QA/QC protocols but did not

compile control charts. A blank sample was inserted in the sample batch sequence immediately

after a highly-mineralized sample expected to return high values of cobalt and/or copper. A

standard or certified reference material (CRM) was inserted at the rate of 1 in every 20 samples.

Warning limits were set at +/-2 standard deviations, and control limits were set at +/-3 standard

deviations. When a quality control sample fell outside the control limits, the cause was

thoroughly investigated, and if need be, the entire sample batch was automatically re-assayed

and all the initial test results are rejected.

11.4.2.1 Blanks

FCC used a barren Apple Creek meta-siltite as a blank to monitor and control contamination

between samples. The assay was considered a failure if the value was higher than three times

the detection limit (DL). Micon was provided with the results of the blank samples and

compiled a control chart (Figure 11.2) incorporating cobalt, copper and gold. Except for only

two samples, the control chart demonstrates that there was no contamination between samples;

if any, then it was insignificant. It has been suggested that the two failures indicated in Figure

11.2 are most likely due to typographic errors.

Figure 11.2

Summary of Blank Samples Results: 1997 to 2006 Drilling

64

11.4.2.2 Standards/CRMs

FCC used three varieties of CRMs i.e. low grade, medium grade and high grade. All CRMS

were prepared at Chemex Laboratories Inc. from mineralized material obtained from the ICP

area. The certified values summarized in Table 11.1 below are based on the averages of assays

obtained from several different reputable laboratories.

Table 11.1

Summary of Certified Values for Standards used at the ICP

Item Cobalt % Copper % Gold oz/t

Standard 1 0.814 0.08 0.060

Standard 2 0.251 0.03 0.030

Standard 3 2.732 0.19 0.139

Any results falling outside the failure limit of +/-3 SD (standard deviation) were rejected

pending investigation into the source of error. FCC’s general practice was to use standards 1

and 2. Micon has summarized the QA/QC results by compiling control charts for the drilling

periods from 1997 to 2006 for standards 1 and 2 (see Figure 11.3 to Figure 11.8).

Figure 11.3

Control Chart for Co: Standard 1

65

Figure 11.4

Control Chart for Co: Standard 2

Figure 11.5

Control Chart for Cu: Standard 1

66

Figure 11.6

Control Chart for Cu: Standard 2

Figure 11.7

Control Chart for Au: Standard 1

67

Figure 11.8

Control Chart for Au: Standard 2

For Standard 1, Figure 11.3 shows only one significant failure plus three borderline failures

for Cobalt; Figure 11.5 shows three borderline failures for copper and Figure 11.7 shows two

borderline failures for gold. For Standard 2, there is only one failure (Figure 11.8) for gold.

There were no blanks/standards failures for the 2010 drill campaign. Thus, overall, the number

of failures/borderline failures is very insignificant to have material impact on the assay

database.

11.5 SUMMARY STATEMENT/COMMENTS

Micon considers the sample preparation, security and analytical procedures to have been

adequate to ensure the integrity and credibility of the analytical results used in the mineral

resource estimation.

Micon believes that the QA/QC aspects of the project have been adequately addressed.

68

12.0 DATA VERIFICATION

The steps undertaken by Micon to verify the data in this Technical Report include a site visit

to the project area, analyzing monitoring reports on the performance of control samples,

reviewing previous verification work conducted by MDA (i.e. co-authors of the March 2015

PEA Technical Report) and conducting a resource database validation.

12.1 SITE VISITS

Micon representatives Chris Jacobs, C. Eng., MIMMM., Barnard Foo, P.Eng., David

Makepeace, P.Eng. and Charley Murahwi, P.Geo., FAusIMM accompanied by SLI staff Scott

Baliski, P. Eng, and Luc Coussement, P. Eng. conducted a site visit to the ICP from 13 to 14

July, 2016. The FCC staff in attendance were George King, P.Geo., William Scales, Preston

Rufe’ and Mike Irish, P.Eng.

On 13 July, 2017, Barnard Foo visited the 6930L adit of the nearby Blackbird mine with FCC

personnel Floyd Varley, Rick Honsinger, George King, Mike Lee, Matt Bender from Samuel

Engineering, Keith Jones and Jimmy Green of Small Mine Development, L.L.C., and a

representative from Blackbird mine.

The data verification activities and results achieved are summarized as follows.

12.1.1 Discussions on Geological Attributes

Discussions held with FCC’s resident geologist centred on the geological model/attributes of

the Ram-Sunshine Deposits including/encompassing the genetic model, mineralization trends,

and the role of structures and lithology. Micon concurs with FCC’s current interpretation as

far as the following deposit attributes are concerned:

Continuity of the mineralization in distinct stratiform/tabular zones.

Consanguineous mineralization events despite the separation into individual zones.

The strong association of BTE/mafic sequences of the Apple Creek Formation with the

mineralization.

Micon has utilized these attributes in verifying the modelling of the deposit.

12.1.2 Discussions on Mine Planning Parameters

During the July 2017 visit, some accessible mine openings at the nearby Blackbird mine were

inspected in order to validate planning assumptions regarding the expected requirements for

geotechnical support and in-fill drilling during underground mine development and production

at the Ram deposit.

69

12.1.3 Field Examination of Out Crops

The surface expression of the deposit is not discernible. The project area is hilly/rugged and

this necessitates the need for a detailed digital terrain model (DTM) for modelling of the

deposit.

12.1.4 Examination of Drill Cores

Examination of drill cores from several drill holes confirms the strong association of

BTE/mafic sequences with the mineralization. All the drilling on the Ram deposit has been

conducted by FCC and thus, protocols pertaining to the exploration history of the deposit and

database build-up, have progressive continuity. The core sizes used generally yield good core

recovery and minimize core loss in bad ground.

12.1.5 Data Collection Techniques/Sampling

Drill cores were photographed prior to logging and sampling. An example is shown in Figure

12.1. Drill hole logs produced by FCC are very detailed and include all essential information,

i.e., drill hole survey information, core losses, rock quality designation, lithology, structure,

alteration, mineralization and sampling.

Sample intervals which varied from <1 ft. to 6.0 ft. were determined based on geologic,

mineralogic and alteration features. This is in line with standard industry practice.

Figure 12.1

Example of the Ram Deposit Drill Core Photograph

Photo supplied by FCC geologist George King, 2016

12.1.6 Down-hole Surveys

Down-hole surveys were done at 150 ft. intervals. The reliability of down-hole surveys is

difficult to confirm, particularly for the deep drill holes. However, the most recent

metallurgical drill hole (RMH-16-06) drilled in 2016, intersected mineralization above the

modelled/estimated position of the mineralized envelope resulting in about a third of the

mineralized intercept being left outside the resource block model (see Figure 12.2). It is partly

70

for this reason that Micon has assigned most of the resource areas covered by deep holes into

the Indicated and Inferred categories.

12.1.7 Analysis of QA/QC Monitoring Charts

Monitoring charts on quality control samples have already been discussed in Section 11 of this

report. The use of quality control samples appears to have been in line with prevailing industry

standards over the drill campaign periods.

Overall, Micon considers the sample preparation, security and analytical procedures to have

been adequate over the different drill campaigns to ensure the integrity and credibility of the

analytical results used for mineral resource estimation. Independent QPs from MDA who have

been associated with the ICP since 1995 to date arrived at the same conclusion.

Figure 12.2

Section Showing the 2016 Metallurgical Drill Hole

12.1.8 Specific Gravity

Specific gravity was measured and averaged from 813 samples as detailed in Section 14.5.3.

71

12.2 REVIEW OF MDA DATA VERIFICATION

MDA has been an Independent consultant on the Idaho project for several years dating back to

1995 and have been involved in all FCC’s previous independent mineral resource estimates

and Independent Technical Reports. The two major database audits conducted by MDA are

summarized as follow.

12.2.1 Database Audit for the 2006 Resource

In 2005 MDA made numerous site visits to the Idaho project area during which time they

reviewed and checked original assays, check assays and QA/QC procedures and results;

reviewed and audited the digital database; examined geologic data and interpretations; and

reviewed and re-sampled representative core intervals. Spot re-sampling produced comparable

results to the original assays.

For drill data prior to the 1999 Ram drilling program, MDA checked about five percent of the

sample intervals in the project database for data entry errors. No errors were found for entries

of cobalt, copper, or gold values; however, the footage for one interval was entered incorrectly.

Approximately 10 percent of the 1999 Ram drill data was audited, and no errors were found.

12.2.2 Database Audit for the 2015

Edwin Peralta of MDA visited the ICP site on December 10 to 12, 2014. Data verification of

the 2010 drill data was completed to bring the 2012 resource estimate and block model to status

as current and compliant with NI 43-101. Collar and downhole surveys were checked against

original data supplied by a third-party surveyor while the assay data was digitally checked

against the original assay lab data. No errors or missing data were encountered and no changes

were made to the database.

12.2.3 QA/QC for the 2006/2015 Resources

MDA reports that “FCC’s QA/QC analytical procedures including assays of check samples,

standard reference material samples, and blanks, all show that the ICP assay data is reliable

and verifiable, and is adequate for estimating the ICP mineral resource”.

Micon Comments: Overall, MDA appear to have been diligent in their data verification.

12.3 DATABASE VALIDATION

Micon verified the database using GEMS mining software to ascertain that it contained all the

essential elements required in the estimation of mineral resources. Checks were also made to

ensure that down-the-hole surveys were making sense and that all drill hole collars conformed

to the DTM.

72

No further data entry from assay certificates checks were deemed necessary having noted the

thoroughness of the data entry checks conducted by MDA who have been associated with the

project over an extended period of over 20 years.

12.4 DATA VERIFICATION CONCLUSIONS

Based on the verification procedures described above, Micon considers the database of the

Ram deposit to have been generated in a credible manner and therefore suitable for use in

mineral resource estimation.

Lack of sampling beyond mineralized zones is a notable weakness in the database. It does not

allow for the proper determination dilution grades. Notwithstanding this shortfall, FCC’s

exploration databases were professionally constructed and are sufficiently error‐free to support

mineral resource estimates.

73

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 METALLURGICAL TESTWORK PROGRAMS

A number of metallurgical testwork programs comprising batch and continuous tests have been

completed using representative samples of the RAM deposit mineralization that support the

Feasibility Study process flowsheet. The Feasibility Study process includes grinding and

flotation at the mine site Concentrator with subsequent leaching of the flotation concentrate at

the Cobalt Processing Facility (CPF) and ultimately production of cobalt sulphate, copper

sulphate, and magnesium sulphate crystals. A gold recovery circuit is also included at the CPF

to produce a gold doré.

The main testwork programs completed to date include the following:

Initial milling and flotation testwork on bulk samples and drill composites performed

by Noranda’s nearby Blackbird Mining Company (BMC) in the 1980’s. BMC

reportedly was successful in producing separate copper and cobalt concentrates using

a differential flotation flowsheet.

Early work by The Center for Advanced Mineral and Metallurgical Processing

(CAMP) in 2001 used approximately 1 ton of large diameter drill core from the RAM

deposit. This testwork included a comprehensive milling and flotation test program and

nitrogen species-catalyzed (NSC) leaching of the batch flotation concentrate.

In 2005 SGS Lakefield (SGS-L) conducted a number of flowsheet development

testwork programs including detailed comminution and flotation testing as well as

preliminary leach testing that confirmed CAMP’s NSC test result.

The initial hydrometallurgical tests completed by SGS-L in 2005 provided the design

criteria used for a Mini Pilot Plant testwork campaign undertaken in 2005 by Mintek,

South Africa. This program was directed by Hatch and was successful in developing a

basic hydrometallurgical process.

Pocock Industrial Inc. conducted solids-liquid separation tests in 2005, including

settling/thickening and filtration studies on samples of cleaner concentrate and rougher

flotation tailings.

Following batch desktop scale tests a full scale pilot plant was operated at Mintek in

2007. This work was directed by Grenvil Dunn of Hydromet (Pty) Ltd. (Hydromet) and

resulted in improved Fe/Cu removal, solution purification steps, consistently high

grade cobalt product (>99.9% Co) and introduced of flash cooling technology. The data

derived from this test program was used to finalize the process design criteria for the

2007 Feasibility Study that was completed by Samuel Engineering (Samuel). The 2007

study produced high-purity cobalt metal, copper cathodes and by-product streams

nickel hydroxide, and magnesium sulphate.

In 2015 Hazen Research completed further flotation and hydrometallurgical testwork

under the direction of Samuel. The objective of this work was to investigate differential

flotation to produce separate copper and cobalt concentrates and review the iron

74

removal, acidulation steps, NSC leach conditions, copper solvent extraction and gold

leach processes. It was noted that the sample used for this program of work was

relatively old and tarnished and although the work was useful, the results were tainted

and could not be used to support the Feasibility Study.

CYTEC Solvay Group (Cytec), conducted bench scale and continuous pilot plant scale

cobalt solvent extraction testwork in 2015 using pregnant leach solution (PLS)

generated by Hazen. The objective of this work was to produce a clean cobalt sulphate

solution that could be fed to the crystallizers.

GE Water & Process Technologies (GE) performed crystallizer bench tests in 2015

with the objective of gathering adequate design data in order to confidently size and

estimate the cost of a commercial cobalt sulphate crystallizer. GE also prepared a

capital cost estimates for the magnesium sulphate and copper sulphate crystallizer

packages for the feasibility study.

In 2016 and 2017 SGS-L completed a program of bench scale testwork to confirm the

Feasibility Study design. This work included differential flotation, copper/iron

removal, NSC leaching, leach residue elemental sulphur recovery and gold leaching.

In 2017 SGS-L completed a series of tests to produce copper and cobalt sulphate

crystals.

13.2 METALLURGICAL SAMPLES

The samples used for the major metallurgical development programs are listed in Table 13.1.

Table 13.1

Summary of Metallurgical Samples

Test Laboratory Quantity Sample Description

CAMP (2001) About 1,000 kg Bulk composite

combining 13 samples

PQ Core from 1999/2000

drilling, main zones 3021, 3022

and 3023

SGS-L (April 2005) Head Sample 778 kg Not specified Not specified

SGS-L (April 2005) Core Composite 275 kg Not specified Not specified SGS-L (April 2005) Reject Composite 155 kg Not specified Not specified SGS-L (April 2005) Reject Composite 2 90 kg Not specified Not specified SGS-L (May 2005) – Composite 1 54 kg Assay rejects Not specified SGS-L (May 2005) – Composite 2 65 kg Assay rejects Not specified SGS-L (May 2005) – Composite 3 47 kg Assay rejects Not specified SGS-L (May 2005) – Sample S 54 kg ¼ core samples Siliceous material from a trench

SGS-L (May 2005) – Sample M 65 kg ¼ core samples Micaceous material

SGS-L (May 2005) – Sample Q 47 kg ¼ core samples Quartzite material

Hazen (2015) 228 kg ¼ core samples Four composites with varying

Cu:Co ratios from 1:1 to 4:1.

SGS-L (2016/17) – Composite 1 and 2 277 kg Met hole RMH-16-06,

½ core samples

Drill targeting high Cu:Co ratio

material

75

Table 13.2 provides a summary comparison between the total proven and probable reserves

stated in the Feasibility Study, the process design criteria and the various metallurgical

composite samples used to develop the process.

Table 13.2

Comparison of Metallurgical Sample Head Grades

Estimate/Sample Co (%) Cu (%) Cu/Co

Ratio Au (g/t) As (%) S (%)

M&I Resources (FS 2017) 0.59 0.73 1.24 0.55

P&P Reserves (FS 2017) 0.47 0.68 1.45 0.55 - -

Design Criteria - Concentrator 0.56 0.60 1.07 0.48 - -

CAMP (2001) - Composite 0.57 0.29 0.51 0.68 - -

SGS-L (April 2005) – Head sample 0.36 0.35 0.97 - 0.59 0.44

SGS-L (April 2005) – Core Comp. 0.49 0.45 0.92 - 0.63 0.94

SGS-L (April 2005) – Reject Comp. 1.69 0.68 0.40 - 0.91 1.22

SGS-L (April 2005) – Reject Comp 2 0.42 0.43 1.02 - 0.61 0.93

SGS-L (May 2005) – Composite 1 0.59 0.44 0.75 0.35 0.75 0.99



SGS-L (May 2005) – Sample S 1.14 1.20 1.05 1.57 2.46 2.09

SGS-L (May 2005) – Sample M 0.92 0.33 0.36 0.64 1.25 0.73

SGS-L (May 2005) – Sample Q 0.41 0.70 1.71 0.60 0.57 1.40

Hazen (2015) – Composite C 0.51 1.89 3.71 0.8 0.68 2.91

Hazen (2015) – Composite A 0.44 0.46 1.05 0.6 0.74 0.90

Hazen (2015) – Composite B 0.51 0.9 1.76 0.8 0.68 1.66

Hazen (2015) – Composite D 0.58 1.2 2.07 1.0 0.87 2.03

SGS-L (2016/17) – Composite 1 1.00 1.61 1.61 0.74 1.30 3.29

SGS-L (2016/17) – Composite 2 0.49 2.05 4.18 1.03 0.60 3.17

The samples used for the metallurgical testwork programs were representative of the RAM

deposit mineralization.

13.3 MINERALOGY

The main cobalt and copper minerals occurring within the RAM deposit mineral resources are

cobaltite (CoAsS) and chalcopyrite (CuFeS2), respectively. The mineralization is typically low

in sulphides and oxidation is prevalent near the surface of the deposit, especially in the areas

where pyrite is present. Gold and silver are present in most of the mineralization in the area.

Characterization studies were undertaken by SGS-L in 2005 on six variability composite

samples. The samples were crushed to 100% -48 mesh and representative portions were

analysed using X-ray diffraction (XRD) and optical microscopy using polished sections and

polished thin sections. The overall mineralogy of the six samples was similar in nature with

some variation in mineral proportions. The samples mainly consisted of quartz/feldspar, micas

(including phlogopite and trace muscovite and sericite), chlorite, amphiboles (hornblende),

garnet (almandine), cobaltite, chalcopyrite, marcasite/pyrite, and lesser amounts of other

opaque minerals including arsenopyrite, pyrrhotite, magnetite, ilmenite, spinel (hercynite) and

iron oxy-hydroxide.

76

The majority of the cobaltite and chalcopyrite grains were within the size-range of 10 to 300

μm and 20 to 400 μm, respectively. For most samples, a bimodal distribution of cobaltite was

noted where the majority of the finer cobaltite grains (5-50μm) occurring as fine inclusions

disseminated within non-opaque minerals (mainly chlorite and micas). Liberation

characteristics of the samples indicated that most of the coarse cobaltite (>100 μm) and coarse

chalcopyrite (>75 μm) grains were liberated.

13.4 COMMINUTION

The abrasion and standard grinding work index test results from a number of test campaigns

are presented in Table 13.3

Table 13.3

Comminution Test Results

Test Program Sample Abrasion Index

(g)

Bond Rod Mill

Index (kWh/t)1

Bond Ball Mill

Index (kWh/t)1

SGS-L (April 2005) Head sample - 5.0 9.0

SGS-L (May 2005) Composite 1 - - 11.4



SGS-L (May 2005) Sample S 0.1578 - 9.4

SGS-L (May 2005) Sample M 0.0222 - 9.9

SGS-L (May 2005) Sample Q 0.0985 - 10.1

Hazen (2015) Composite B 0.0618 9.1

Average 0.0851 5.0 10.3

Study design criteria Not specified Not specified 12.7 1 Metric

The comminution test results suggest that the mineralization is not abrasive, and the grinding

work indices are relatively low compared with the industry database.

13.5 FLOTATION

A series of flotation tests have been completed in order to develop and optimise the flotation

conditions to produce a bulk concentrate containing both cobalt and copper and a copper

scalping followed by bulk concentrate flotation, which will be used when the copper to cobalt

ratio of the feed material is above the criteria set by the CPF.

Batch rough, cleaner and locked cycle testwork was completed by CAMP, SGS-L and Hazen.

The parameters investigated during the flowsheet development programs were:

Primary grind size and flotation kinetics.

Optimum reagent suite and reagent addition rates.

Effect of copper sulphate.

Effect of regrinding the concentrate.

77

13.5.1 Bulk Concentrate Flotation

The conclusions from the metallurgical flotation testwork programs were as follows:

Flotation kinetics generally increased with finer grind sizes although the effect tended

to be less pronounced as the cobalt and copper grades increased. The metal recoveries

were not significantly different with P80 grind sizes between 67 and 90 µm but there

tended to be a drop in cobalt recovery with a coarser grind of 105 µm. There was

negligible drop in copper recovery with a coarser grind. The Feasibility Study primary

grind design P80 is 75 µm.

The reagent additions used in the early CAMP studies were relatively high and the 2005

SGS-L work substantially reduced the dosage rates without adversely affecting cobalt

and copper recoveries. The addition of copper sulphate was also eliminated. The

reagents and associated dosage rates used in the Feasibility Study are based on the 2005

SGS-L locked cycle tests.

The 2005 SGS-L series of tests indicated that only minor upgrading took place with

regrinding the rougher concentrate prior to cleaning, it also showed lower cobalt

recoveries while the copper recoveries remained the same. Regrinding was therefore

not included in the Feasibility Study design.

Following development phase batch rougher testwork the metallurgical response of the

samples to the newly developed test conditions was evaluated by locked cycle flotation tests.

Three tests were completed using each of the three composites. A schematic showing the 2005

locked cycle test (LCT) flowsheet is provided in Figure 13.1.

78

Figure 13.1

SGS-L 2005 LCT Flowsheet

The results of these as well as the early CAMP locked cycle tests are summarized in Table

13.4. This table also shows the average Feasibility Study estimated results.

Table 13.4

Summary of Bulk Concentrate Flotation LCT Results

LCT Tests Feed Grades (calc.) Conc. Grades Recoveries

Co (%) Cu (%) Au (g/t) Co (%) Cu (%) Au (g/t) Co (%) Cu (%) Au (g/t)

CAMP (2001) 0.57 0.29 0.69 14.4 7.4 13.6 92.7 92.8 72.9

SGS-L (2005) Comp.1 0.60 0.42 0.33 13.3 9.7 7.0 93.0 96.0 90.3

SGS-L (2005) Comp.2 0.73 1.07 0.64 8.3 12.9 6.8 90.7 96.5 84.5

SGS-L (2005) Comp.3 1.10 0.43 0.64 17.9 7.1 10.0 95.1 97.1 92.0

Average 0.75 0.55 0.57 13.5 9.3 9.3 92.9 95.6 84.9

Average SGS-L (2005) 0.81 0.64 0.54 13.2 9.9 7.9 92.9 96.5 88.9

Feasibility Study LOM 0.47 0.68 0.55 10.0 14.9 9.9 93.4 96.5 88.9

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Following the 2005 LCT, SGS-L completed a series of rougher and cleaner flotation tests using

the standard test conditions using the S, M and Q composites. A summary of these variability

test results is provided in Table 13.5.

Table 13.5

Summary of Bulk Concentrate Flotation Variability Results

Variability Tests

(SGS-L 2005)

Feed Grades (calc.) Conc. Grades Recoveries

Co (%) Cu (%) S (%) Co (%) Cu (%) S (%) Co (%) Cu (%) S (%)

Comp.1 (Test F30) 0.57 0.42 0.99 13.8 10.5 25.0 93.7 96.7 95.5

Comp.2 (Test F31) 0.74 1.06 2.16 8.78 13.3 27.6 89.9 95.5 95.5

Comp.3 (Test 32) 1.08 0.41 1.20 16.9 6.56 19.3 93.9 95.5 94.6

Comp. S (Test V4) 1.14 1.20 .2.08 12.5 13.4 24.5 89.4 92.0 93.3

Comp. M (Test V5) 0.92 0.33 0.77 20.9 5.7 21.5 87.9 78.5 91.6

Comp. Q (Test V6) 0.40 0.72 1.39 7.39 14.2 28.7 91.7 96.5 93.8

Apart from Composite M (relatively high micaceous sample), batch bulk concentrate cobalt

recoveries ranged from around 90% to 94% and the copper recoveries were between 92% and

97%.

13.5.2 Copper Scalping Flotation

The Cobalt Processing Facility (CPF) has been designed to be able to handle a concentrate feed

containing a minimum and maximum quantity of copper. This equates to approximately a

minimum Cu to Co ratio of 0.6, below which iron will need to be added to the leach process to

ensure complete arsenic removal as scorodite, (FeAsO4.2H2O), and a maximum Cu:Co ratio

of around 2.0, above which the capacity of the unit processes within the CPF would be limited.

Initial scoping work was undertaken by Noranda in the 1980s and by SGS-L in 2005 that

demonstrated the viability of differential flotation to produce both a copper concentrate and a

cobalt concentrate by taking advantage of the different flotation kinetics of the chalcopyrite

and cobaltite minerals. More recent work by Hazen in 2015 suggested that a copper scalping

flotation circuit using starvation quantities of collector (PAX) prior to a bulk sulphide circuit

should be capable of producing a copper concentrate with low cobalt values and a bulk sulphide

concentrate containing cobalt and the remaining copper and arsenic.

In 2016, SGS-L commenced a bench scale batch and LCT flotation program using freshly

drilled mineralized core with the objective of providing a Feasibility Study design for the

circuit, which would be used during periods within the life-of-mine (LOM) plan that produced

relatively high copper mineralization. This work at SGS-L was undertaken using two

composite samples which were selected from 25 individual samples to provide a relatively

high (Composite 1) and very high (Composite 2) Cu to Co ratio (see Table 13.6 for the head

grade analyses of these composites).

The 2016/17 differential program was based on the copper scalping rougher flotation and

Co/Cu bulk flotation circuits being undertaken at the concentrator site then separate copper

cleaning at the CPF. The tailings from the copper cleaning circuit would be combined with the

80

bulk concentrate and fed to the CPF. The copper scalping rougher concentrate was therefore

filtered and re-pulped using fresh water prior to cleaning.

A schematic of the LCT test flowsheet, which is the differential flotation circuit proposed for

the concentrator site is provided in Figure 13.2.

Figure 13.2

SGS-L 2016/17 Differential Flotation LCT Flowsheet

A summary of the LCT tests and subsequent batch copper cleaning test results is provided in

Table 13.6. The copper cleaning circuit included the re-grinding of the copper rougher

concentrate to a P80 of approximately 20 µm.

The results from the differential flotation and copper cleaning tests show that a very high grade

copper and low grade cobalt concentrate can be produced containing around 33% Co, 0.4% Co

and 0.5% As.

81

Table 13.6

Summary of the Differential Flotation LCT Results

LCT Tests Feed Grades (calc) Conc Grade Recoveries

Co (%) Cu (%) As (%) Co (%) Cu (%) As (%) Co (%) Cu (%) As (%)

SGS-L (2016/17) Composite 1 – Cu:Co Ratio = 1.62

Cu Rougher Con 1.00 1.62 1.29 2.61 28.00 3.53 13.5 89.5 14.1

Bulk Concentrate - - - 12.90 2.60 16.80 80.3 10.0 80.5

Cu Cleaner Con - - - 0.40 33.00 0.56 0.8 36.8 0.8

Bulk + Cu Cl. Tails1 - - - 9.67 10.66 12.66 93.0 62.7 93.8

SGS-L (2016/17) Composite 2 – Cu:Co Ratio 4.40

Cu Rougher Con 0.45 1.98 0.54 1.27 28.30 1.54 16.9 84.4 16.8

Bulk Concentrate - - - 7.92 7.51 10.30 70.3 15.0 74.6

Cu Cleaner Con - - - 0.41 33.50 0.45 4.2 62.0 3.5

Bulk + Cu Cl. Tails1 - - - 6.14 11.49 7.92 83.0 37.4 87.9 1 The combination of the bulk concentrate and copper cleaner tailings is the feed to the CPF

13.5.3 Concentrate Characteristics

During the 2005 testwork undertaken by SGS-L a bulk flotation concentrate sample grading

15.5% Co, 10.8% Cu and 22.3% S was submitted for mineralogical examination. The primary

focus of the study was to determine the modal abundance and mineral textural associations

with emphasis on the nature and mode of occurrence of the gangue mineralogy (diluting the

concentrate). The mineralogical examinations included XRD, optical microscopy using a

polarizing light microscope, and scanning electron microscopy (SEM) equipped with an

energy dispersive spectrometer (EDS).

The results indicated that the major opaque minerals were cobaltite (41 wt.%), chalcopyrite

(28 wt.%), and pyrite (7 wt.%). Non-opaque minerals consisted primarily of quartz (12 wt.%)

and phyllosilicates (4.5 wt.%).

The results from the SGS-L 2005 multi element chemical analyses of LCT concentrates from Composite 1

through 3 and batch concentrates of Comps S, M and Q are presented in Table 13.7.

The final copper concentrates produced from the batch copper cleaning tests during the

2016/17 differential flotation test program at SGS-L produced a copper concentrate containing

33% Cu, 0.4% Co, 0.5% As, 31% Fe and 33% S.

Table 13.7

Multi-Element Bulk Concentrate Analyses

Element Units Composite Sample and Test Reference

1 2 3 S M Q

LCT 1 LCT 2 LCT 3 V-4 V-5 V-6

Ag g/t 13 18 9 22 19 23

Al g/t 12,000 11,000 15,000 2,700 12,000 6,600

As g/t 76,000 38,000 55,000 110,000 52,000 40,000

Ba g/t 27 38 28 9 28 31

Be g/t 0 0 0 0 0 0

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Element Units Composite Sample and Test Reference

1 2 3 S M Q

LCT 1 LCT 2 LCT 3 V-4 V-5 V-6

Bi g/t 350 2,000 720 1,400 890 470

Ca g/t 910 1,400 1,800 370 880 330

Cd g/t <25 <25 <25 <25 <25 <25

Co g/t 140,000 81,000 190,000 140,000 210,000 79,000

Cr g/t 180 150 140 1,000 160 140

Cu g/t 97,000 130,000 71,000 140,000 56,000 150,000

Fe g/t 180,000 220,000 130,000 190,000 110,000 250,000

K g/t 3,100 3,000 3,600 840 2,500 3,000

Li g/t <5 <5 <5 <5 <5 <5

Mg g/t 4,900 3,600 6,200 600 5,200 1,400

Mn g/t 180 190 150 90 140 82

Mo g/t 11 10 25 8 32 17

Na g/t 56 84 70 30 45 55

Ni g/t 520 320 1,100 1,400 1,300 480

P g/t 300 300 400 <200 310 <200

Pb g/t <100 <100 <100 <100 <100 <100

Sb g/t <60 <60 <60 <60 <60 <60

Se g/t 75 83 120 100 170 68

Sn g/t 60 71 64 77 37 84

Sr g/t 5 6 9 4 5 3

Ti g/t 900 850 1,600 110 1,500 410

Tl g/t <60 <60 <60 <60 <60 <60

V g/t <20 <20 <20 <20 <20 <20

Y g/t 230 210 370 120 260 42

Zn g/t 300 400 180 1,100 2,100 640

13.5.4 Concentrator Flotation Recoveries

The Feasibility Study mine plan suggests that the copper scalping scenario may only be

required during the latter production years of the mine when the Cu to Co ratio of the

concentrator feed raises above two. Therefore, the bulk concentrate scenario will be the option

used for the majority of the projected operating life of the mine.

Figure 13.3 and Figure 13.4 present the LCT and variability test results for cobalt and copper

recovery into a bulk concentrate. The figures show the metal recovery verses the upgrading

ratio (concentrate grade / feed grade [c/f]). Also plotted on these figures are the recoveries and

c/f ratios for the process design criteria (PDC) used for the Feasibility Study and the PDC used

for the 2007 study.

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Figure 13.3

Bulk Concentrate Cobalt Recovery Test Results

Figure 13.4

Bulk Concentrate Copper Recovery Test Results

These results show that both cobalt and copper recoveries were typically above 90% for a

variety of feed mineralization. The copper recoveries were consistently between 95% and 97%

while the cobalt recoveries were more variable with no obvious grade or upgrade to recovery

relationship. The variability test results shown in these figure were from batch tests and

therefore the recoveries are lower than would be expected for a full scale re-cycle or LCT

environment.

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Averaging the SGS locked-cycle test results indicate recoveries of 92.9% Co, 96.5% Cu and

88.9% Au and it is these recoveries that are recommended to be used for the Feasibility Study.

If the copper scalping circuit is required to reduce the copper loading in the CPF then a the

project will produce a high grade copper concentrate containing approximately 33% Cu and

less than 0.5% Co and As. In removing a portion of the copper the feed to the CPF

hydrometallurgical circuit will have a Cu to Co ratio of less than two.

As shown with the 2016/17 testwork at SGS-L, the combined copper recovery into the high

grade copper concentrate and CPF feed bulk concentrate will be greater than 98%. However,

the testwork using various mineralized feed samples has shown that the cobalt recovery into

the CPF feed concentrate tends to reduce with a higher Cu to Co ratio of the concentrator feed.

Figure 13.5 presents the Co recoveries verses Cu:Co ratios for the LCT and variability tests

undertaken by SGS-L in 2005 and 2016/17. These results suggest that a Co recovery of 88.7%

would be a reasonable estimate, which is based on a Cu:Co ratio of two. However, as can been

seen in Figure 13.5, there is only one data point for a feed sample with Cu:Co ratio greater than

2 and therefore more work will be required to more accurately predict the Co and Cu recoveries

when higher Cu:Co ore will be mined during the latter years of the mine life.

Figure 13.5

Flotation Test Results – Cobalt Recovery vs Cu:Co Ratio

13.6 SOLID-LIQUID SEPARATION

13.6.1 Tailings

The pertinent testwork used to design the tailings dewatering system was completed by Pocock

Industrial, Inc. (Pocock) and FLSmidth Minerals (FLS).

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Pocock completed both Kynch-type static thickener tests and dynamic (high rate) tests using a

sample of flotation tailings. Pocock also completed viscosity tests and slump tests as well as

vacuum and pressure filtration tests.

Testwork performed by FLS showed that tailings feed from the flotation circuit should be fed

to a paste thickener for consolidation of up to 70% solids. The tailings thickener was sized

based upon the unit area recommendation from FLS. A series of filtration tests undertaken by

FLS on the paste thickener underflow sample revealed that vacuum filtration was the best

method and could achieve a dewatered tails of approximately 80% solids, which is suitable for

stacking and loadout to the TWSF.

13.6.2 Concentrate

The concentrate dewatering design is based upon test work performed by Pocock. Pocock

completed Kynch-type static thickener tests, vacuum filter tests and pressure filter tests, using

a sample of bulk cleaner flotation concentrate.

In 2011, from discussions between Samuel Engineering and filter vendors during the detailed

design of the concentrator plant, it was recognized that value could be added by removing the

concentrate thickener and directly feeding the filter with the 25% solids slurry from the cleaner

flotation circuit.

13.7 HYDROMETALLURGICAL PROCESS

Current plans call for the initial production of one bulk sulphide concentrate that contains

cobalt, copper, and gold. Further processing for recovery of the individual metals is via a

hydrometallurgical treatment plant known as the cobalt processing facility (CPF). During the

latter period of the mine life when the Cu to Co ratio is above two, the process flowsheet will

include copper scalping flotation in order to limit the copper feed to the CPF. This comprises

copper rougher flotation at the mine site concentrator then copper cleaning at the CPF with the

production of a copper flotation concentrate, which will be shipped directly to a copper smelter.

The copper cleaner tailings will be added to the cobalt rich bulk concentrate and fed to the

CPF.

The hydrometallurgical treatment plant consists of the following unit processes:

Concentrate preparation and regrinding.

Acidulation.

Pressure Leaching.

Neutralization.

Copper recovery and sulphate crystallization.

Iron removal.

Cobalt recovery and sulphate crystallization.

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Magnesium recovery and sulphate crystallization.

Gold recovery.

Residue disposal.

13.7.1 Leaching Circuit

Nitrogen species leaching testwork has been undertaken by a number of testwork facilities,

including CAMP in 2001, SGS in 2005 and 2017, Hazen in 2015, Mintek in 2005 and

Hydromet in 2006/2007. This work has demonstrated that nitrogen species-catalyzed (NSC)

leaching is effective and that high cobalt and copper extractions can be achieved.

The leaching circuit has been tested in a number batch tests (CAMP, SGS and Hazen) as well

as a continuous basis at Mintek in 2005 and 2007.

SGS – 2005

A series of 12 batch leach tests were completed by SGS in 2005 on a sample of flotation

concentrate containing 12.3% Co, 9.1% Cu, 15.8% As, 14.7% Fe, 22.0% S, 13.1 g/t Au and

26.9 g/t Ag. The best results gave Co extraction of 99.1%, Cu extraction of 93.4% and a PLS

containing 16 g/L Co, 7.1 g/L Cu, 3.7 g/L As, 8.1 g/L Fe, 46 g/l FAT.

For each test it was noted that sulphur pellets were formed representing approximately 9% of

the initial weight of concentrate and assaying round 9% Cu, 0.2% Co, 0.6% As, 22% Fe, 7 g/t

Au and 50 g/t Ag.

A standard toxicity characteristic leaching procedure (TCLP) test was conducted on the leach

residue and the results indicated that the arsenic in the solid product was stable.

It was noted that there was a direct positive correlation between residual arsenic in solution

and residual free acid. Acid addition therefore has to be minimized but not as much as to

interfere with copper and cobalt extractions.

Mintek Mini Pilot Plant – 2005

In 2005, Mintek in South Africa, under the guidance of Hatch, completed short mini plant

campaign for the recovery of copper and cobalt from 43 kg of flotation concentrate containing

about 10.6% Co, 9.5% Cu, 13.9% As and 17.3% Fe. The unit processes tested during the mini

plant campaign included the following:

Re-grinding to P80 of 10 microns.

Pre-leaching (acidulation).

NSC Leaching.

Leach residue filtration.

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Copper neutralization (and filtration).

Copper solvent extraction.

Iron and arsenic removal (and filtration).

Copper and aluminium removal (and filtration).

Copper and zinc ion exchange.

Nickel ion exchange.

Cobalt precipitation and redissolution (purification).

Cobalt electrowinning.

Leaching conditions were as follows:

Slurry volume 21.3 L in 45 L autoclave (53% freeboard).

NaNO3 concentration of 6 g/L.

Leaching temperature 155°C.

Acid addition 40 g/L H2SO4.

Pre-treatment 1-hour preleach at 90°C to get rid of volatiles (acidulation).

Leaching retention time of 2 hours at 8.5 bar, using oxygen to maintain the pressure.

The average leaching efficiencies for the test program for Co and Cu were 98% and 95%

respectively. Average leaching efficiencies for Fe and As were 20% and 50%, respectively,

and sulphuric acid consumption in the leach was 92 kg/t of dry concentrate. The average

oxygen consumption was 350 kg/t of dry concentrate and a weight loss of around 20%

(including mass of sulphur balls with residue mass) were experienced during the leach.

Hydromet (2006/7)

In 2006/7 Hydromet conducted a series of NSC batch leach tests to examine whether:

The nitrogen species leach can be operated on a continuous basis.

Any sodium that enters the circuit can be immobilised with the leach-end residue.

The aluminium and silicon that leach are rejected ultimately to the leach-end residue.

Arsenic concentrations can be reduced in the continuously fed autoclave discharge

liquors.

The filterability and thickening characteristics of the leach discharge slurry can be

improved over what was achieved in an earlier test work program.

A standard base case open circuit test using similar conditions applied by Mintek in 2005 gave

Co and Cu leach extractions of 95.2% and 92.6%, respectively.

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Tests simulating flash thickener recycle (FTR), which is an autoclave cooling step that permits

a capacity increase by depleting the aqueous fraction from an exothermic autoclave circuit

thereby increasing the retention time of the solids in the autoclave, gave leach extractions of

approximately 97% for Co and 95% for Cu.

Test including simulating FTR and neutralization to reduce the autoclave discharge acid tenor

from around 30 g/L to 7 g/L using lime to the flashed slurry gave leach extractions of

approximately 97% for Co and 93% for Cu and reduced the soluble As in the leach liquor to

87 mg/L. These tests also confirmed that a large component of the added soluble aluminium

and sodium reported to the residue.

Two residues samples from these batch tests were submitted to Mintek for standard pH 4.5

Acetate TCLP tests, the results from these tests are summarized in Table 13.8.

Table 13.8

Mintek TCLP Test Results on Leach Residue

Test

No.

Elemental Analysis (mg/L)

Mn Co Ni Zn As Sr Pb F Cl Ti

5 1.1 0.32 0.34 0.16 0.52 1.75 0.14 1.16 3.6 <0.01

6 0.57 0.21 0.23 0.04 1.76 3.71 0.18 1.16 1.8 <0.01

The conclusions from batch leach tests simulating heat removal and acid neutralisation test

program were as follows:

Good copper and cobalt recoveries can be achieved.

The minimum free acid in the leach for acceptable copper and cobalt recoveries is

above 5 g/L.

The NSC leach could be operated continuously provided that the significant quantity

of the NOX in the vapor and liquid phase can be recovered and returned to the process

as a catalyst.

Partial neutralisation of the acid in the leachate with lime improved the filtration

process.

The proprietary flash cooling step followed by a solids thickener and recycle of the

thickened solids to the autoclave (FTR) provided a means of removing heat from the

autoclave as well as increasing the capacity of the existing circuit for operation in a

cascade of three reactions in series. Normally three vessels in series in a continuous

operation are insufficient to prevent “by passing”. However, the additional capacity

from the FTR should allow for near completion of the reaction within the three vessels

in series.

Reasonable thickening and filtration fluxes can be expected for the separation of the

leach liquors from the leach residue.

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Sodium and aluminium can be precipitated in the leach residue to the point that these

elements should not build up to high concentrations in a continuous mode.

Hazen 2015

Using the flotation concentrates generated during the flotation testwork by Hazen, a series of

batch leach tests were completed. The best test from this non-optimized program of work

resulted in Co and Cu extractions of 98% and 86%, respectively. Arsenic extraction was 19%

and the final acid in solution was 14 g/L H2SO4. Lower Cu extractions than previously achieved

were attributed to the formation of elemental sulphur particles which tended to encapsulate

sulphides and significantly reduce the sulphide oxidation kinetics.

SGS-2017

In 2016, fresh drill core samples were selected, prepared and transported to SGS in Lakefield,

Canada by eCobalt. These samples were prepared into two composites with different Cu to Co

ratios and used for the copper scalping and cleaning tests discussed in Section 13.5.2. The

concentrates from the locked cycle flotation tests were blended to produce samples with 1:1,

1.5:1, and 2:1 approximate Cu:Co mass ratios. Each concentrate was then reground in a

ceramic pebble mill to a particle size P80 of approximately 15 µm and leached in a 2 L capacity

titanium batch autoclave. The leaching protocol used for these batch tests was as follows:

Sulphuric acid addition of 15 g/L.

Pre-acidification for 1 hour at 90°C and atmospheric pressure.

Nitric acid addition to achieve 4 g/L N in solution.

Oxygen added to achieve initial pressure of 621 kPag (90 psig) at a temperature of

155°C.

Two hours leaching time.

The test results from these tests are presented in Table 13.9.

Table 13.9

SGS 2017 Leach Test Results

Units Test P2 Test P3 Test P4 Test P5

Cu:Co ratio - 1.5:1 2:1 1:1 1.5:1

Feed Co assay % 10.6 12.7 7.5 10.6

Feed Cu assay % 6.9 6.3 7.9 6.9

Grind size - P80 µm 15.2 15.7 15.0 15.2

PLS free acid g/L 25 14 26 13

Acid consumption kg/t 66 150 27 130

Final pressure kPag 758 1048 1014 1048

Sulphide oxidation % 97 97 94 99

Weight loss % 54 56 64 71

Co extraction % 99.1 99.6 99.5 99.7

Cu extraction % 94 95 91 99

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Units Test P2 Test P3 Test P4 Test P5

As extraction % 33 49 81 75

Fe extraction % 53 57 77 76

Co in filtrate g/L 30.7 32.6 30.3 32.9

Cu in filtrate g/L 10.2 14.5 7.1 10.8

As in filtrate g/L 4.4 7.0 11.4 10.6

Fe in filtrate g/L 15.6 19.7 21.9 23.4

SO4 in filtrate g/L 250 260 260 300

Co residue grade % 0.45 0.2 0.4 0.22

Cu residue grade % 0.92 1.16 1.74 0.34

As residue grade % 13.6 10.7 5.98 8.50

Fe residue grade % 21.3 22.3 15.0 18.6

STOT residue grade % 6.76 8.34 14.4 5.35

S0 residue grade % 3.02 4.33 9.70 1.89

13.7.1.1 Feasibility study Leach Extractions

The overall cobalt and copper recoveries used in the Feasibility Study economic model are

98.6% and 93.1% respectively.

13.7.2 Leach Residue Thickening and Filtration

Hydromet 2006/07

Thickening and filtration tests were completed by MacOne Agencies using leach samples from

the 2006/07 Hydromet batch leach test program. The thickening tests were made to quantify

the settling flux of the flash thickener residue. This was found to be approximately 12.5 m2/t.h

for a 6 to 8% feed well density. The flocculant dose was 35 to 45 g/tonne and the test results

suggested that a 40% thickener underflow density could be achieved.

The filtration test data suggested filtration and washing design parameters for a vacuum unit

to be 126 kg/m2.h to achieve a washed product containing 50% solids by weight.

Hazen, 2015

Kynch settling tests were performed by Hazen on NSC discharge slurry at 55°C. The best

thickened solids density achieved was 29% solids, unit area calculated at 0.083 to 0.153

m2/(t/d) and an initial settling rate of 15 to 18 m/h.

Standard modified TCLP tests using two final residue samples suggested that the arsenic would

fail the test and therefore the residue would be classified as hazardous wastes. The test leachate

As analyses for the two tests were 13 mg/L and 23 mg/L, which compares to the US

Environmental Protection Agency limits of 5 mg/L.

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13.7.3 Iron Removal (Fe Precipitation and Thickening)

Prior to feeding the cobalt recovery circuit, raffinate from the copper solvent extraction circuit

is fed to the iron removal circuit, which also removes any remaining arsenic and copper from

the circuit. The thickened solids from this step, containing the precipitated iron, arsenic and

copper are recycled to the pressure leaching circuit for stabilization of the arsenic as scorodite

and re-leaching of any co-precipitated Cu and Co.

Mintek Mini Pilot Plant - 2005

As part of the mini pilot plant, Mintek operated a four-stage Fe removal circuit, with an

increasing pH profile by the across the stages with the addition of lime. Both the Fe and As

removal efficiencies was generally greater than 90%.

Pilot Plant Test – Mintek – 2007 and Hydromet (2007)

Following batch laboratory test work performed by Mintek to determine preliminary process

conditions for the pilot plant, a pilot campaign was launched by Mintek during October 2007

using a synthetic solution representative of the expected process solution as specified by

Hydromet.

At the beginning of the pilot campaign, approximately 300 L of the synthetic solution was

treated by Hydromet for iron and arsenic removal. Hydromet reported that significant iron

removal could be achieved from synthetic liquors containing the typical elemental assays of a

Copper SX raffinate. Near complete removal of iron was achieved (residual ± 15 mg/L) and

this occurred at a pH of 5.2. Arsenic co-removal with the iron was not quantitative and the best

achieved result was approximately 56 mg/L at a pH of 5.0.

Hazen 2015

The copper solvent extraction raffinate from the batch testwork program undertaken by Hazen

in 2015 was used to prepare cobalt solvent extraction feed liquor. The combined raffinate

contained 356 mg/L Cu, 9.51 g/L Co, 2.08 g/L Fe, and 10.4 g/L free acid, with a density of

1.134 g/mL. Although this raffinate had a higher copper concentration than the target, it was

deemed acceptable to prepare the cobalt SX liquor.

The best result at pH 5.18 resulted in greater than 98% Fe and 99% As precipitation with only

7% co-precipitation of Co. The resultant filtrates from this series of experiments were

combined and shipped to Solvay for cobalt SX development work.

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SGS-2017

During the confirmation test program undertaken by SGS in 2016/17, neutralization tests were

performed on a synthetic liquor sample representing Cu SX raffinate using specifications

provided by Mike Irish, an independent consultant employed by FCC. The solution contained

0.37 g/L Cu, 22 g/L Co, 1.1 g/L Fe, 28 g/L Mg, 18 g/L H2SO4 and 0.01 g/L As, and had a pH

of 1.1.

This series of neutralization tests were undertaken at 80 °C with the staged addition of

powdered MgO with a reaction time of between 24 and 48 hours. The tests resulted in the

removal of primary impurity elements to below detection limits (< 3 mg/L As, < 0.1 mg/L Cu,

and < 4 mg/L Fe). The co-precipitation of cobalt of only 2.4% was achieved with 91% copper

precipitation.

SGS also completed a series of static settling tests to confirm the iron removal thickener design

criteria. However, this test was unsuccessful, and SGS-L concluded that additional work needs

to be completed on representative non-synthetic solutions in order to provide adequate support

for the detailed design.

13.7.4 Cobalt Precipitation and Re-dissolution


Following concentrate leaching, copper neutralization using lime, copper solvent extraction,

iron and arsenic removal using lime, copper and zinc removal using ion exchange and nickel

removal using ion exchange, the cobalt was precipitated at pH 8.5 using lime then re-dissolved

using sulphuric acid prior to cobalt electrowinning. At pH of 8.5 and a temperature of 50°C,

more than 99% of the cobalt reported to the cobalt hydroxide precipitate.

The precipitation and re-dissolution tests showed the following”:

Approximately 20 to 30% of the Mg but 100% of the Mn co-precipitated with Co and

re-dissolved in the re-dissolution stage.

All remaining Ni, Cu, Zn, Al and most of the As were carried over to the re-dissolution

section and reported to the Co electrolyte. Some of the Fe also reported to the

electrolyte.

Micon notes that the proposed precipitation circuit in the feasibility study is not designed to

produce an electrowinning feed solution and uses MgO rather than lime for pH control.

Pilot Plant Test – Mintek – 2007

There were six main unit operations piloted by Mintek, these included:

Cobalt precipitation.

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Magnesium leach.

Cobalt re-dissolution.

Solvent extraction.

Nickel ion-exchange.

Cobalt electrowinning.

The synthetic solution prepared by Hydromet, following the iron removal step, was used for

the Mintek pilot plant. The first processing step was cobalt precipitation using MgO powder

followed by magnesium leaching from the Co precipitate then re-dissolution of a pure cobalt

sulphate solution.

The pilot campaign was conducted from the 15th of October 2007 to the 8th of November

2007. The cobalt precipitation unit operation was run for a continuous period of approximately

23 days during this time.

From a feed solution containing approximately 69 g/L cobalt, on average 99.6 % of the cobalt

in the feed was precipitated, with a magnesia consumption of 1.45 kg MgO/ kg cobalt in the

feed solution. The final cobalt tenor of the filtrate was approximately 50 mg/L.

Mintek’s estimate of the cobalt precipitate species was approximately 46% basic cobalt

sulphate (CoOH(SO4)½ ) and 54% cobalt hydroxide (Co(OH)2). The average analyses of the

precipitate from the pilot plant operation is presented in Table 13.10.

Table 13.10

Mintek 2007 Pilot Plant Average Cobalt Precipitate Analysis

Element % Element %

Co 28.22 Cr 0.05

S 5.76 Mn 0.21

Mg 6.06 Fe 0.25

Al 0.05 Ni 0.20

Si 0.20 Cu 0.05

Ca 0.48 Zn 0.05

Ti 0.05 Pb 0.05

V 0.05 As 0.05

Sulphuric acid was used to leach un-reacted magnesia from the metal hydroxide precipitate

during the majority of the pilot plant campaign. This process was termed the “Magnesium

Leach” and has been substituted by the introduction of a hydrocyclone which will separate the

coarse unreacted/cores of MgO from the very fine cobalt precipitate.

Cobalt re-dissolution was the third unit operation tested in the pilot plant as part of the cobalt

purification circuit of the cobalt solvent extraction circuit feed solution. The two-stage cobalt

re-dissolution unit operation was successful at removing iron, copper, aluminum and a large

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portion of manganese from the leach liquor. Cobalt re-dissolution was 55% over the first stage

and in excess of 98% at the end of the second stage with a solution pH of 2.

13.7.5 Cobalt Solvent Extraction

The cobalt solvent extraction process design is mainly based on bench scale and continuous

pilot plant scale testwork undertaken by CYTEC Solvay Group (Cytec) in 2015 using a

pregnant leach solution (PLS) generated by Hazen.

The objective of the Cytec work was to prepare a clean cobalt sulphate solution suitable to feed

to the cobalt sulphate crystallizers.

The Cytec work recommended 1 extraction stage, 4 scrubbing stages to remove Mg and 1

stripping stage.

An additional phase of cobalt solvent extraction testwork to support the feasibility study was

planned for 2016/2017. The results from this optimization testwork will be required for the

project detailed design phase.

13.7.6 Cobalt Sulphate Crystallization

GE – 2015

In 2015, GE Water & Process Technologies (GE) was contracted to perform crystallizer bench

tests.

The objective of the initial phase of testing by GE was to find an operating temperature which

could achieve a separation between magnesium and cobalt. Synthetic solutions for this phase

of work were prepared using technical grade cobalt and magnesium sulphate.

The second phase of testwork was designed to investigate the potential impact of trace

impurities expected in the commercial operation on crystal quality, shape and size. These tests

used solvent extraction PLS provided by FCC.

The results from the testwork determined that cobalt sulphate crystals of proper purity can be

obtained provided that feed concentrations of minor constituents are maintained below at the

specific level. It was also determined that there is no significant favorable partitioning of

magnesium during the crystallization process and that feed to the crystallizer must have

concentration of <20 ppm magnesium to meet the product purity requirement for that

constituent.

Crystals produced at 45ºC and 65ºC were adequately large and well formed to utilize standard

dewatering and drying equipment. The Co analyses of the crystal products varied between

19.7% and 21.7%, which suggested that the products were a mixture of 6 and 7 hydrates.

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GE noted that adequate information was obtained during this program of bench testwork to

provide a design for a crystallizer. However more detailed work was needed to be completed

prior to the finalized design of the system.

SGS-L – 2017

In 2017, SGS-L were engaged by FCC to run a cobalt crystallization testwork program. The

objective of this work was to produce samples that could be used for marketing. Initial testing,

both batch and continuous, was conducted on synthetic solutions prepared using reagent grade

chemicals because limited quantities of high purity cobalt sulphate liquor from Co SX testing

(conducted by Cytec) was available.

The overall test circuit design comprising two evaporation reactors and three crystallization

reactors is shown in (see Figure 13.6).

Figure 13.6

Cobalt Sulphate Crystallization Test Process Flowsheet

Two campaigns of continuous crystallization were completed by SGS-L. The initial 25.5 hour

campaign used synthetic solution with a Co tenor of 103 g/L (average impurity content 0.18

g/L) produced over 2 kg of crystals at a calculated Co purity of 99.980% (based on impurity

content detected). The second continuous test campaign used Co SX PLS from Cytec with a

Co tenor of 95 g/L, consisted of 23 hours of continuous operations and produced almost 1.7

kg of crystals at a calculated purity of 99.998% Co.

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Semi-quantitative X-Ray diffraction analysis of acetone washed PP2 crystals suggested that

65% of the cobalt sulphate was present as the heptahydrate state and the remaining 35% was

as the hexahydrate state. The Co assay of the crystals was 21.1%.

13.7.7 Copper Solvent Extraction

The design of the copper solvent (SX) extraction is based on the initial work undertaken by

Mintek in 2005 and copper extraction tests by Hazen in 2015. The Cytec Solvay Group

(Solvay) was also engaged by FCC to provide copper SX recovery modeling and laboratory

support.

In 2017, SGS-L completed a bench scale Cu SX test using concentrate leach liquor as part of

preparing copper sulphate solution for crystallization trials.


Copper solvent extraction from a sulphate solution is a well-established technology for the

recovery of copper. The pilot plant consisted of three extraction units and two stripping units.

The plant operated for 56 hours on synthetic solution then a further 349 hours using pregnant

leach solution (PLS). The average feed and raffinate metal concentrations, when using the

leach solutions, are presented in Table 13.11.

Table 13.11

2005 Mintek Mini Pilot Plant Average Copper Solvent Extraction Results

Description Cu Co Fe As Al Zn Ni

Feed (PLS) 6.0 12.6 0.8 2.6 0.24 0.044 0.12

Raffinate 0.044 13.1 0.9 2.6 0.25 0.048 0.10

It was noted that the copper SX process is extremely selective for Cu, so very few impurities

were transferred to the loaded strip solution.

Hazen 2015

Hazen conducted tests to recover copper from the NSC liquor using solvent extraction.

Standard shake out tests were completed to generate an extraction isotherm and a McCabe-

Thiele diagram in order to determine conditions of the continuous copper SX operation. The

McCabe-Thiele diagram indicated that two stages at an organic/aqueous volumetric ratio of 1

would give a raffinate containing less than 200 mg/L Cu. No organic stripping tests were

conducted by Hazen.

The feasibility study copper SX process comprises three extraction stages and two stripping

stages which corresponds to the 2005 mini pilot plant testwork. Although no stripping was

undertaken by Hazen, the extraction isotherm and testing suggest that less stages may be

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adequate to produce a copper sulphate solution suitable for the production of good quality

copper sulphate crystals.

13.7.8 Copper Sulphate Crystallization

The design of the feasibility study copper sulphate crystallization circuit is based on the models

and experience of GE, who provided a quote of the equipment.

A batch scale crystallization testwork program using was undertaken by SGS in Lakefield,

Ontario in 2017. A composite of liquor samples from nitrogen species catalyst (NSC) leach

tests that were performed by SGS-L in 2017 was used for the test program. The feed solution

was subjected to primary neutralization, locked cycle copper solvent extraction (loading,

stripping, and crystallization) and finally raffinate neutralization testing.

A five stage locked cycle copper solvent extraction test was performed and the aqueous strip

solution was subjected to copper crystallization by chilling and the spent mother liquor from

crystallization was then used to strip the next stage of loaded organic solution. The acidity of

the mother liquor was adjusted as required with concentrated sulphuric acid. SGS-L noted that

there were no signs of decreased extraction or stripping rates throughout the locked cycle test.

The purity of the copper sulphate crystals produced ranged from 99.9% to 99.99%. SGS-L

noted that the concentrations of cobalt and magnesium in the crystals increased through the

five cycles of the test, from 4 g/t Co and 5 g/t Mg in the first cycle to 85 g/t Co and 92 g/t Mg

in the fifth cycle.

Neutralization of the copper raffinate samples with MgO resulted in 8% co-precipitation of

cobalt. This precipitate will be recycled internally in the process to recover the cobalt. The

copper raffinate contained a residual amount of 190 mg/L copper after precipitation, and

additional testing was recommended by SGS-L in order to reduce this.

13.7.9 Magnesium Sulphate Crystallization

No testwork using representative samples has been undertaken on the magnesium sulphate

crystallization circuit.

13.7.10 Gold Recovery Circuit

Gold is recovered from the leach circuit residue. The circuit comprises flotation to

recover/remove any elemental sulphur balls followed by conditioning with MgO (or lime) to

raise the pH to above 9, then carbon-in-leach (CIL) cyanidation to recover gold and silver. The

most recent testwork completed on this circuit was completed by SGS in 2017 using residue

samples from the batch leach tests.

The leach residue sample used for the testwork graded 7.4 % Sº, 10% As, 1.8% Cu, 0.30% Co,

19 g/t Au and 399 g/t Ag, based on the test calculated head.

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The single rougher sulphur flotation test consisted resulted in a mass pull to the sulphur

concentrate of 15% with 27% of the elemental sulphur reporting to the concentrate along with

31% Au and 20% Ag.

The CIL test achieved recoveries of 88% Au and 16% Ag. Cyanide consumptions was 8.9 kg/t

of NaCN and the MgO addition was 87 kg/t. Carbon loading was 191 g/t Au and 986 g/t Ag,

although this is likely to be significantly higher for the continuous process. Due to the

excessively high MgO addition, the project will use lime for pH control as this reagent is less

expensive and more suited to the control of pH around the CIL target of 10.

13.8 RECOMMENDATIONS FOR FUTURE TESTWORK

13.8.1 Copper Flotation

Additional tests are recommended to verify the copper scalping and cleaning flotation

performance using fresh samples that represent the relatively high Cu:Co ratio mineralization

planned to be mined and processed in the later years of the mine life.

13.8.2 Cobalt Solvent Extraction

Pilot plant cobalt solvent extraction testwork needs to be completed in order to provide design

details for the process. The objective of this additional testwork will be to confirm extraction

kinetics, determine optimum percent solids MgO vs. cobalt recovery, confirm Co/Mg

selectivity, determine strip liquor impurities and confirm the overall circuit mass balance. The

cobalt and zinc stripping conditions also need to be confirmed.

13.8.3 Copper Solvent Extraction

The design of the copper solvent extraction circuit is based on the 2005 mini pilot plant test

program, the object of which was to produce cathode copper not copper sulphate crystals.

There may be a benefit of reviewing this circuit as the differences in the optimal PLS

specifications for these two applications (electrowinning vs crystallization) could result in a

simpler system and lower capital costs.

13.8.4 Crystallization

Although adequate bench scale testwork has been completed to provide a design for the cobalt

crystallizer circuit, additional detailed work needs to be completed to establish the actual

maximum recovery rate per pass and the critical impurity concentration prior to the finalized

design and procurement of the system. It is recommended that extended continuous operations

be performed using a high purity feed electrolyte to produce additional cobalt sulphate crystals

and investigate the impact of impurity buildup of the product over a more prolonged period of

operation. A process to treat the bleed stream and recycle cobalt will also need to be developed.

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Testwork needs to be completed using representative solution samples to provide detailed

design details of the magnesium sulphate crystallizer circuit.

Based on the recent copper crystallization testwork at SGS-L, it is recommended to perform

additional neutralization tests on both the feed solution and the copper raffinate with the

objective to (i) minimize cobalt and copper losses in the primary precipitate stage and (ii)

reduce the copper concentration in the feed to cobalt recovery, without losing cobalt to the

copper precipitate. This work should also include an evaluation of a two stage precipitation

process at two target pH levels for both processes.

13.8.5 Gold Recovery Circuit

Additional testwork is required to optimize the elemental sulphur flotation and the cyanide

leaching circuit circuits. Testwork also needs to be completed in order to model the CIL circuit

and gold/silver carbon loading as well as the cyanide destruction circuit.

13.8.6 CPF Pilot Plant

Much of the CPF processing circuits have been designed using batch tests or continuous pilot

tests using synthetic solutions. It is therefore recommended that the complete CPF process be

tested using a continuous pilot plant using composite samples of flotation concentrate.

During the pilot plant testwork program it is suggested that solid/liquid separation and washing

of precipitates should be evaluated using pressure filtration and/or centrifuging to develop an

industrially robust methodology for removing the precipitates produced within the process

flowsheet.

13.8.7 Process Modelling and Simulation

As part of the feasibility study process engineering completed by SLI, a MetSim model was

developed for the CPF. This model needs to be developed to a higher level of detail using the

results from the additional testwork recommended above. The more robust model will be

available to stress test the final detailed design of the CPF.

13.8.8 HAZOP Studies

During the detailed design phase it is important to complete a hazard and operability study

(HAZOP) in order to identify and evaluate potential risks to personnel or equipment so that

the design can mitigate these risks.

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14.0 MINERAL RESOURCE ESTIMATES

The estimation of the mineral resources for the ICP’s Ram deposit was conducted by MDA in

2012 and updated in 2015 (Samuel Engineering, 2015). Following the completion of data

verification as outlined in Section 12, Micon reviewed/audited the resource in 2016 as detailed

hereafter and re-categorized it to ensure compliance with the CIM Definition Standards for

Mineral Resources and Mineral Reserves.

14.1 DATABASE DESCRIPTION

The database for the Ram deposit is well structured and contains all the essential elements

required in the estimation of mineral resources; it comprises collar, survey, assay, lithology

tables, density information and a DTM. The database has 4,302 assays derived from 78

diamond drill holes all of which were utilized in the resource estimation. In addition, there are

9 metallurgical drill holes in the database that were used qualitatively for guiding geological

interpretation, but they were not used in the actual grade interpolations and are not included in

the sample statistics/variography.

Drill spacing in the largest and best understood Main Zone (domain-horizon code 3023) is on

average about 230 ft. (70 m). Excluding the poorly drilled peripheral parts of the model,

average drill spacing drops to about 200 ft. (61 m). The closest spacing of 65 ft. to 121 ft. (i.e.

20 m to 36 m) is restricted to the central part and close to surface.

Micon verified the database using GEMS mining software. Assay data, down hole interval data

and grid coordinates are expressed in imperial units. The effective date of the resource database

is March 22, 2012. Since then, no additional drilling has been conducted on the property save

for 1 metallurgical hole drilled in 2016.

14.2 OVERVIEW OF MDA’S ESTIMATION METHODOLOGY

The following summary of procedures is excerpted from the March 2015 PEA Technical

Report by MDA and Samuel Engineering Inc. Micon has included some edits where warranted.

In 2012, MDA received FCC’s database and performed checks on reasonableness. Improbable

data were sent to FCC, who resolved the issues to their satisfaction. Data verification, sample

integrity and quality assurance studies, were completed in 2015 in order to bring the 2012

resource estimate and block model to status as current and compliant with NI 43-101.

Statistics were run on the analytical data in the database to evaluate metal-grade distributions

and to support metal-domain modeling. Figure 14.1 shows the distributions of the three metals.

Deviations in distributions suggest the need for multiple domains in modeling the metals.

In 2012 MDA completed the current estimate using cross-sectional geologic interpretations to

construct geologic and metal domains that were used to control the estimation of grades into

the three-dimensional blocks. New geological sections were made and the geology was

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interpreted by MDA. FCC reviewed the geology and requested some modifications that were

made by MDA.

Figure 14.1

Quantile-Quantile Plot of Co, Cu and Au

A sectional model of cobalt mineralization and related dilutionary zones (a shell around the

metal domains, described in more detail later) was made and passed on to FCC to review. After

making changes requested by FCC, MDA used this cobalt model to guide the definition of the

copper and gold sectional models, which were also sent to FCC for review and comment.

After FCC accepted the sectional model, the planar sections were snapped to the drill holes to

more accurately reflect three-dimensional spatial locations. The geology, cobalt, copper, and

gold domains were then further refined three-dimensionally on level plans.

Assay data were evaluated statistically by domain for each metal. Capping was done on each

domain and each metal separately, and the capped assays were composited to 5 ft., honoring

the domain and dilutionary zone boundaries. Multiple estimation runs were completed.

During the entire process, an objective of MDA was to maintain continuity with the previous

estimates in most aspects of modeling.

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14.3 GLOBAL/GENERAL STATISTICS

Micon conducted statistics on the raw assay data. The results are summarized in Table 14.1

and are comparable to MDA’s statistics.

Table 14.1

Descriptive Statistics of the Assay Database

Description Cobalt (%) Copper (%) Gold (oz/ton)

Count 4302 4302 4302

Minimum 0.000 0.000 0.0000

Maximum 10.650 10.200 0.5730

Median 0.025 0.060 0.0007

Mean 0.149 0.244 0.0048

Std Dev 0.474 0.649 0.0162

CV 3.191 2.655 3.367

The histograms for all three elements (i.e. Co, Cu and Au) are lognormal with positive

skewness. Accordingly, Micon generated log-probability plots (Figure 14.5) to determine the

recommended mineralization envelope cut-off grade for modelling/wireframing.

Figure 14.2

Ram Deposit Global Log Probability Plot for Co

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Figure 14.3

Ram Deposit Global Log Probability Plot for Cu

Figure 14.4

Ram Deposit Global Log Probability Plot for Au

As demonstrated in Figure 14.2 to Figure 14.4, the suggested cut-off grades for

modelling/wireframing are 0.03%, 0.06% and 0.004 oz/t for Co, Cu and Au domains,

respectively.

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14.4 GEOLOGIC AND DOMAIN MODEL

14.4.1 MDA Modelling

The following description is excerpted from the March 2015 PEA Technical Report written by

MDA and Samuel Engineering Inc., with minor edits/additions by Micon.

Cross sections were plotted on 200-ft. intervals and oriented at an azimuth of 63.5° (looking

333.5°). FCC’s lithology codes were plotted on the drill-hole traces. The lithologic codes, core

photos, and drill logs were used to define the geologic model, which differentiated the

exhalative rocks from quartzite. About a dozen beds of exhalative rocks were defined

throughout the deposit, with many of the exhalative beds being interpreted to coalesce and

bifurcate.

The geological model formed the basis of the estimate. The model is based on logged data,

which in some rare instances contradicts itself. However, for the most part, particularly in the

Main Zone (horizons 3021, 3022 and 3023), continuity and confidence in the interpretation is

high.

Once the lithology was interpreted, the sections were used to guide the cobalt, copper, and gold

mineral domain modeling. In general, these mineral domains represent the Hanging wall,

Main, and Footwall zones of the Ram deposit discussed in Section 7.3.3, though the current

resource model contains four additional hanging wall horizons that are limited in spatial extent.

Also, included within the model are weakly-mineralized materials that are modeled as volumes

of “dilutionary” zones around the metal domains. These are the areas around the mineralized

horizons, in which partial assay data are available and mineralization is weak to non-existent.

These volumes were defined and modeled to provide dilution in the block model. They are not

natural domains, as drill-hole sampling and assaying are incomplete for defining them. This

means the dilutionary volumes cannot be good predictors of grade. In some cases, the sampling

ends in mineralized material of one or any combination of the metals. The sampling is often

done in a very small rind around the mineralized domains. Other times the sampling around

the mineral domains is extensive. MDA was compelled to use the margins of the sampling to

define the boundary of the dilutionary volumes.

The horizon domains followed the geologic model, which is based on the interpretation that

the mineralization follows stratigraphy. MDA understands that some remobilization has

occurred, particularly with copper, but the apparent remobilized metal is not a significant

portion of the deposit. Contacts were defined based on grade and geologic continuity, and in

some cases by referring to the geologic logs. The modelled domains encompass grades above

~0.03%Co, ~0.1%Cu, and ~0.01oz Au/ton for cobalt, copper, and gold, respectively. These

grade cut-offs were not used as absolutes and the grades were not contoured; they defined the

boundaries between dominantly higher-grade and dominantly lower-grade areas of the deposit.

Once the domains were defined on section and digitized, they were used to code the assay

database.

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Dilutionary volumes can have exhalative rocks within them and can be mineralized, but the

mineralization is non-continuous. If continuity exists, the definition of a unique mineral

domain would have been mandated.

The geology, cobalt, copper, and gold domains were then refined three-dimensionally on level

plans, one for each bench level of five feet, corresponding to the mid-bench of each level of

blocks in the block model. A total of 319 level plans were interpreted.

These level plans were used to define individual mineral-domain and dilutionary zone

percentages, as well as the horizon domain codes into the block model. Each block was

assigned the total percent of mineralized domain volumes in the block, and the block’s domain

code was assigned to be that of the largest volume of mineralized domain in it. This same

method was applied for the dilutionary zone percentages and horizon coding. These

percentages and codes were calculated independently for cobalt, copper, and gold.

The geologic level plans were used to code blocks within an exhalative bed, where the entire

block is exhalative if it intersects the level plan interpretation. The level plans were also used

to code blocks as being within or external to a post-mineral quartz vein, using the same method.

A list of mineralized and dilutionary horizon codes is presented in Table 14.2. They follow the

same general numbering method for location as existed in the previous string models.

Table 14.2

List of Mineralized and Dilutionary Domain Codes

Dilutionary Domain Cobalt Domain Copper Domain Gold Domain

2950 2951 - 2951

2952 - -

2960 2961 - -

2970 2971 2971 2971

2972 2972 2972

2980 2981 2981 2981

2982 2982 2982

2990 2991 2991 2991

2992 2992 2992

2993 2993 2993

3000 3001 3001 3001

3002 3002 3002

3010 3011 3011 3011

3012 3012 3012

3020 3021 3021 3021

3022 3022 3022

3023 3023 3023

3025 3025 3025

3030 3031 3031 3031

3032 3032 3032

3033 3033 3033

3034 3034 3034

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Internal to these cobalt, copper, and gold domains are distinct sample populations that probably

reflect assorted styles of mineralization, represented by intensity of sulphides. With a drill

spacing of over 200 ft., they cannot be explicitly modeled, and this estimate assumes that these

internal sample populations have similar continuity. This is a risk imparted into the estimate

by the wide-spaced drilling. Capping assays has mitigated some of this risk. In many cases,

mineralization is controlled by stratigraphy at sample-size scale (feet or less). In other words,

there is high variability in grades within a single domain and across stratigraphy. Sample

compositing masks this variability and may provide a sense of security statistically, but it also

imparts some risk in the estimate.

Figure 14.5 shows a cross section of the geologic/mineralization interpretations.

Figure 14.5

Section Showing Interpreted Mineral Domains

Source: Generated by Micon from the Ram deposit wireframes, 2016.

14.4.2 Micon Review and Wireframing

Micon reviewed the MDA modelling and established that it is sensible and acceptable;

however, it does not allow for volume checks and hence, resource tonnage verification. It is

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also evident that MDA calculated the percentage of block inside mineralized rock using the

level plans (i.e. 2D method), while Micon’s preference is to use 3D wireframes.

To overcome the shortfalls highlighted above, Micon used MDA’s 5 ft. interval cobalt level

plans to construct wireframes for each horizon. The wireframes were constructed using Surpac

and Gems mining softwares. The procedure simply involved connecting the level plans every

5 ft. with 3D triangles, making sure all details of bifurcations and/or splays were triangulated

properly. The relationship between MDA’s 5 ft. level plans and Micon’s wireframes is

demonstrated in Figure 14.6.

Figure 14.6

Ram Deposit Isometric Projection Showing Wireframes Honouring the 5 ft. Level Plans

The geometry/shapes originally interpreted by FCC and MDA were maintained in the

conversion of level plans into solids. The interpretation of the domains honours both the

geology and assays.

14.5 GRADE CAPPING, COMPOSITING AND DOMAIN STATISTICS

Sample lengths in the database vary from less than 5 in. to 6 ft. The mode of the sample lengths

is 2 ft. A detailed analysis of sample lengths versus assays reveals that the majority of the

‘super high grades’ are being influenced by short sample lengths of 1 ft. or less. Thus, grade

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capping was conducted after compositing. The composite length selected is 2 ft. and is based

on the mode of the sample lengths.

Following compositing, Micon used log-probability plots to determine grade capping values

for Co, Cu and Au. Log-probability curves were used because the histograms of the of the

composite samples show log-normal populations. The Co probability plots for the three most

important horizons are presented in Figure 14.7, Figure 14.8 and Figure 14.9. Similar log-

probability curves were generated for Cu and Au.

Figure 14.7

Cobalt Log-probability Plot for Horizon 3021

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Figure 14.8


Figure 14.9


110

The statistics of the uncapped composites and the capped composites are summarized in Table

14.3 and Table 14.4, respectively. Note that the maximum values in Table 14.4 equate to the

capping levels.

Table 14.3

Statistics of Uncapped Composites

Element Cobalt Copper Gold

Zone 3021 3022 3023 3021 3022 3023 3021 3022 3023

Number of samples 62 108 324 72 78 332 33 46 189

Minimum value 0.010 0.012 0.030 0.020 0.018 0.010 0.004 0.006 0.003

Maximum value 2.222 3.800 6.110 9.040 5.062 9.640 0.065 0.091 0.573

Mean 0.315 0.547 0.624 0.834 0.655 1.153 0.023 0.028 0.033

Median 0.245 0.263 0.349 0.475 0.468 0.691 0.019 0.023 0.018

Geometric Mean 0.216 0.315 0.377 0.465 0.389 0.727 0.018 0.022 0.020

Variance 0.105 0.436 0.555 1.906 0.540 1.659 0.000 0.000 0.003

Standard Deviation 0.325 0.661 0.745 1.381 0.735 1.288 0.016 0.018 0.053

Coefficient of variation 1.032 1.207 1.194 1.656 1.122 1.117 0.701 0.666 1.605

Table 14.4

Statistics of Capped Composites


Zone 3021 3022 3023 3021 3022 3023 3021 3022 3023

Number of samples 62 108 324 72 78 332 33 46 189

Minimum value 0.010 0.012 0.030 0.020 0.018 0.010 0.004 0.006 0.003

Maximum value 1.000 3.800 3.000 2.000 2.500 5.000 0.065 0.091 0.250

Mean 0.295 0.547 0.602 0.634 0.622 1.114 0.023 0.028 0.031

Median 0.245 0.263 0.349 0.475 0.468 0.691 0.019 0.023 0.018

Geometric Mean 0.214 0.315 0.375 0.442 0.385 0.723 0.018 0.022 0.020

Variance 0.054 0.436 0.395 0.304 0.333 1.223 0.000 0.000 0.001

Standard Deviation 0.232 0.661 0.628 0.551 0.577 1.106 0.016 0.018 0.039

Coefficient of variation 0.788 1.207 1.043 0.870 0.928 0.992 0.701 0.666 1.240

Capped Composites 1 0 4 5 1 6 0 0 1

Micon notes that MDA used a different approach by capping assays before compositing and

selected a composite length of 5 ft. The global end result is comparable as revealed in Table

14.5 but interpolated grades will vary slightly.

Table 14.5

Comparison of MDA and Micon Average Values of Composites


Zone 3021 3022 3023 3021 3022 3023 3021 3022 3023

MDA capped samples 9 2 3 8 2 5 0 2 1

MDA capping values 1 4 4 2.5 3 6 - 0.1 0.25

MDA Mean 0.27 0.541 0.614 0.630 0.600 1.100 0.023 0.026 0.031

Micon capped composites 1 0 4 5 1 6 0 0 1

Micon capping values 1 3.8 3 2 2.5 5 0.065 0.091 0.250

Micon Mean 0.295 0.547 0.602 0.634 0.622 1.114 0.023 0.028 0.031

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Based on the above results, Micon accepted the rest of the MDA capping grades applied to the

rest of the Ram deposit horizons/domains as summarized in Table 14.6.

Table 14.6

Details of Assay Capping Values by Horizon

Domain Cobalt Copper Gold

Top Cut %Co No. Capped Top Cut %Cu No. Capped Top Cut No Capped

2951 0 0

2952 0 - -

2961 0 - -

2971 2 4 2 2 0.05 1

2972 1 4 0 0.03 1

2981 0 0 0

2982 1.5 2 2 2 0

2991 1 3 3 1 0.05 2

2992 1 1 2 2 0

2993 1 1 2 1 0

3001 0 2 1 0

3002 1 1 2 2 0

3011 0 0 0

3012 0 0 0

3021 1 9 2.5 8 0

3022 4 2 3 2 0.1 2

3023 4 3 6 5 0.25 1

3025 0 1.5 1 0

3031 1 1 0 0

3032 0 1 1 0

3033 1.5 3 1 4 0

3034 1 1 1 2 0

2950 0 0 0

2960 0 0 0

2970 0 0.3 7 0

2980 0 0.3 4 0.01 2

2990 0 0.4 3 0

3000 0 0.5 1 0

3010 0 0.4 2 0

3020 0.2 4 0.4 26 0

3030 0 0.4 9 0

14.6 GEOSTATISTICS

Horizon 3023 is the largest and best drilled horizon of the Ram deposit; it carries about 70%

of the Measured and Indicated resource. The rest of the horizons/domains have inadequate data

to support variograpghy/spatial analysis. Micon conducted variography for horizon 3023 using

the 2 ft. composite samples to determine the optimum distance over which samples could be

correlated, and the parameters/dimensions for the search ellipsoid.

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Initially, a down-hole variogram was computed to establish the nugget effect; thereafter,

variograms covering all the principal geometrical directions were computed and modelled

using the nugget effect established from the down-hole variogram. The principal results are

presented in Figure 14.10 to Figure 14.12; the remainder are summarized in Table 14.7.

Figure 14.10

Co Variogram along the Major Axis (Strike Direction) for Horizon 3023

Notes: Nugget = 0.3; Anisotropy factors: Major/Semi-major = 1.773; Major/Minor = 28.924

Figure 14.11

Cu Variogram along the Major Axis (Strike Direction) for Horizon 3023

Notes: Nugget = 0.3; Anisotropy factors: Major/Semi-major = 2.540; Major/Minor = 10.854

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Figure 14.12

Au Variogram along the Major Axis (Strike Direction) For Horizon 3023

Notes: Nugget = 0; Anisotropy factors: Major/Semi-major = 1; Major/Minor = 13.093

Table 14.7

Summary of Variography Results for Horizon 3023

Description Cobalt Copper Gold

Bearing 171.466 171.466 171.466

Plunge 0 0 0

Dip -50 -50 -50

Downhole variogram range 7 8 8

Major axis range (Along strike) 700 500 670

Semi-major axis range (Down dip) 350 150 150

Minor axis range (Across width) 50 50 50

The long ranges of influence along the major axis are indicative of good geological and

mineralization continuity on a global scale as observed in drill intercepts continuously over a

strike distance of 3,500 ft. These ranges are not unusual for IOCG deposits. However, there

could also be an artefact of drill hole spacing which averages about 200 ft. and does not allow

for the determination of spatial variability on a local scale. The downhole variogram ranges

reflect limited continuity on a local scale as expected. Overall, these results support the

dimensions of the search ellipsoids that were used by MDA in the resource estimation. The

estimation parameters are discussed in Section 14.7.2.

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14.6.1 Density

Data provided by FCC in 2012 contains 625 specific-gravity samples from the Ram deposit.

The data were collected by FCC using the water immersion method.

After statistically evaluating those 625 samples in the context of the coded geology, i.e.,

exhalative (BTE) and non-exhalative material, MDA applied a tonnage factor of 10.8 ft.3/ton

for exhalative rocks and 11.2 ft.3/ton for all other material.

An additional 188 specific gravity measurements were provided to MDA by FCC in 2015.

These data were collected by FCC using the same water immersion method as for the 2012

data. FCC also re-submitted the 2012 density data after reviewing and correcting minor

portions of the data set.

The combined 2012-2015 density data were again coded to geology and evaluated. Descriptive

statistics of the complete set of specific-gravity data are presented in Table 14.8. Analyses of

these data indicate no material differences (1% or less) compared to the 2012 density results.

Accordingly, no changes were made to the density values used in the resource model.

MDA also reviewed these data in the context of depths below topography. While there are

insufficient samples to make a definitive conclusion, there are indications that weathering in

the top 100 ft. has decreased the density of the exhalative rock, possibly due to the destruction

of sulphides.

Table 14.8

Summary Statistics on Specific Gravity Samples

Description BTE Rock Other Rock

SG Tonnage Factor SG Tonnage Factor

Count 356 356 457 457

Minimum 2.30 13.93 2.24 14.30

Maximum 3.9 8.21 3.65 8.78

Median 2.98 10.75 2.82 11.36

Mean 3.01 10.64 2.84 11.28

Std Dev 0.23 - 0.16 -

CV 0.08 - 0.06 -

14.7 ESTIMATION

14.7.1 Block Model Definition

The block model definition utilized by MDA is presented in Table 14.9. The upper limit

representing surface topography is based on the digital terrain model (DTM) provided by FCC.

The parent block size is based on the envisaged selective mining unit (SMU) and the geometry

of the deposit. A ‘Percent Model’ was used to represent the mineralized volume. A volume

check of the block model versus the wireframes created by Micon revealed a good

representation of the volume of the deposit components.

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Table 14.9

Ram Deposit Block Model Attributes

Item X Y Z

Origin Coordinates (ft) 1133.383 19889.131 5800

Maximum Coordinates (ft) 3433.383 23879.131 8020

Block Size (ft) 6 2 5

Rotation (degrees) 26.5 (anti-clockwise)

The block model is rotated to reflect the orientation of the deposit.

14.7.2 Estimation/Search Parameters

The search ellipse configurations were defined using variography as a guide, combined with

the geometry of the deposit and average drill hole spacing. A two-pass estimation procedure

was used for the interpolation. For all passes, the maximum number of samples per drill hole

was set to control the number of drill holes in the interpolation. The search parameters adopted

for grade interpolation are summarized in Table 14.10.

Table 14.10

Estimation Parameters for Co, Cu and Au

Description Parameter

Pass1 (Short pass)

Composites: minimum/maximum/maximum/hole (all searches) 2 / 10 /3

Search Bearing/Plunge/Tilt (all searches) 170o / 0o / 50o

Inverse-distance power Co=3, Cu=4, Au=3

Pass Search (ft): major/semi-major/minor 300/300/150

Length weighting Yes

Pass 2 (Long pass)

Composites: minimum/maximum/maximum/hole (all searches) 1 /10 /3

Search Bearing/Plunge/Tilt (all searches) 170o / 0o / 50o

Inverse-distance power Co=3, Cu=4, Au=3

Pass Search (ft): major/semi-major/minor 500/500/500

Length weighting Yes

The search ellipse ranges (for the major and semi-major axes) are well supported by the results

of the geostatistical analysis conducted by Micon, save for the minor axis which appears too

relaxed. However, the minor axis is already constrained by the modelled narrow width and

cannot be influenced by the size of the ellipse.

14.7.3 Grade Interpolation

The mineralized horizons/domains of the Ram deposit are highly variable in width, resulting

in many composites having lengths less than 5 ft. In general, the shorter sample lengths tend

to have higher grades than those with the full 5 ft. lengths. Thus, to accommodate all sample

data in the estimate, MDA used all composite data and weighted them by their lengths during

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grade interpolation into the block model. The estimation procedure restricted composite

selection to those samples coded to the domain being estimated. Based on the exhalative origin

suggested for these deposits, the metal grades of one bed would not be related to the metal

grades of another bed. The distribution of Co grades within the Main Zone horizons is shown

in Figure 14.13 to Figure 14.15.

Figure 14.13

Long Section Distribution of Co Grades in Horizon 3021

Source: Generated by Micon from the Ram deposit block model, 2016

Figure 14.14



117

Figure 14.15



14.7.4 Block Grades Validation

14.7.4.1 Visual Inspection

Micon validated the block grades by visual inspection in plan and section to ensure that block

grade estimates reflect the grades seen in intersecting drill holes. The majority of the sections

viewed reflect a good match between drill intercepts and resource blocks. An example is shown

in Figure 14.16.

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Figure 14.16

Section Through the Ram Block Model for Horizon 3023

14.7.4.2 Swath Plots

Micon generated swath plots as part of the resource model review. Swath plots are used to

demonstrate how well the block model estimates honour the spatial trends within each domain.

A certain degree of smoothing is to be expected but the general shape of the trend line should

be similar between the drill hole data/composites and the block estimates. Figure 14.17 to

Figure 14.19 show the swath plots for the three Main zone horizons which carry the Measured

and Indicated resources of the Ram deposit.

119

Figure 14.17

Cobalt Swath Plot for the Main Domain Horizon 3021 of the Ram Deposit

Figure 14.18

Cobalt Swath Plot for Domain 3022 of the Ram Deposit

120

Figure 14.19

Cobalt Swath Plot for Domain 3023 of the Ram Deposit

Note that there is more variability in the drill hole composites but the overall trend of the drill

hole data is matched.

14.7.4.3 Parallel Estimate

Micon conducted further validation by running an independent parallel estimate for horizon

3023 only, using 2 ft. composites and the ID2 interpolation method. The cobalt resource block

grades are slightly higher than the MDA interpolated grades. This is due to one or a

combination of the following:

The better refinement offered by using higher resolution 2 ft. composites as opposed to

5 ft. composites.

The metal preserved by capping the grade after compositing.

Nonetheless, the results are comparable with an insignificant difference of between 8% and

10%.

14.7.5 Mineral Resource Parameters and Categorization

Assuming a cobalt price of USD 14.50/lb, the cut-off grade that gives the Ram deposit

reasonable prospects for economic extraction by underground methods is 0.20% Co. This cut-

off grade is considered reasonable based on similar deposits currently being exploited.

121

The copper and gold resources are those resources carried within the blocks which attain the

cobalt cut-off grade. Micon believes the contributions from copper and gold will cushion any

downward trends in the cobalt price.

MDA’s protocols for resource categorization as disclosed in the March 2015 PEA Technical

Report are as follow:

“For a block to be classified as Measured, there must also be a minimum of two samples, from

two different holes, within an average distance of 125 feet or less from the block. For a block

to be classified as Indicated, there must be two samples from two different holes whose average

distance is 200 feet or less from the block”.

Micon categorized the resource using the drill hole spacing stipulated in MDA’s protocols but

imposed an additional requirement that drill hole coverage for a Measured resource must be

supported by a minimum of four holes in adjacent cross-sections such that each section is

supported by two holes that are paired one vertically above the other. This leaves no doubt as

to the continuity in both the geometry and the grade.

14.7.6 Mineral Resource Statement

Micon has completed the mineral resource validation and categorization of the Ram Deposit.

The mineral resources are reported at a cut-off grade of 0.20% Co and are summarized in Table

14.11. As stated earlier, the copper and gold resources are those resources carried within the

resource blocks which attain the cobalt cut-off grade.

Table 14.11

Ram Deposit Mineral Resources at 0.2% Co Cut-off

Category Zone Co%

Cut-off

Resource

(Tons)

Co

(%)

Co

(000 lbs)

Au

(oz/t)

Au

(oz)

Cu

(%)

Cu

(000 lbs)

Measured All Zones 0.2 1,725,000 0.54 18,590 0.014 24,300 0.76 26,325

Indicated All Zones 0.2 1,711,000 0.64 21,988 0.017 29,900 0.71 24,111

M+I All Zones 0.2 3,436,000 0.59 40,578 0.016 54,200 0.73 50,436

Inferred All Zones 0.2 1,543,000 0.51 15,594 0.012 18,700 0.68 21,032

i. CIM Definition Standards (2014) were followed for mineral resource estimation.

ii. The effective date of this resource estimate is 27 September, 2017.

iii. The mineral resource is estimated at a cut-off grade of 0.20% Co.

iv. The Mineral Resources are estimated using an average long-term cobalt price of USD 14.50 per lb.

v. Totals may not add correctly due to rounding.

The estimated mineral resources conform to the current CIM Definition Standards for mineral

resources and mineral reserves, as required by Canadian NI 43-101. However mineral

resources, unlike mineral reserves, do not have demonstrated economic viability.

Main Zone horizons/domains (3023, 3022 and 3021) contribute about 87% of the Measured

and Indicated mineral resources of the Ram deposit. The distribution of the mineral resource

categories within horizons of the Main Zone is shown in Figure 14.20 to Figure 14.22.

122

Figure 14.20

Long Section of Horizon 3023 Showing Resource Categories

Figure 14.21


123

Figure 14.22


As can be seen in the above Figure 14.20 to Figure 14.22, there has been insufficient

exploration/drilling to define the inferred resources as an indicated or measured mineral

resource.

14.7.7 Risks/Uncertainties

Factors beyond the control of Formation Capital may affect the future status of the Idaho

Cobalt Project estimated mineral resources. Mineral prices are subject to volatile price changes

due to a variety of factors including international economic and political trends, expectations

of inflation, global and regional demand, currency exchange fluctuations, interest rates and

global or regional consumption patterns, speculative activities and increased production due to

improved mining and production methods.

Thus, the mineral resource estimated will always be sensitive and vulnerable to fluctuations in

the price of cobalt, copper and gold and other related factors mentioned above. Other than this,

Micon believes that at present there are no known environmental, permitting, legal, title,

taxation, socio-economic, marketing or political issues which could adversely affect the

mineral resource estimated above. However, mineral resources unlike mineral reserves, do not

have demonstrated economic viability.

124

15.0 MINERAL RESERVE ESTIMATES

For the ICP, the Measured and Indicated mineral resource from horizons 3021, 3022 and 3023

were considered in the mine plan to be converted into the mineral reserve. The considered

Measured and Indicated mineral resource from these horizons comprise approximately 87% of

the total mineral resource at ICP.

Conversion of the mineral resource estimates to mineral reserve was inclusive of the modifying

factors, diluting material and allowances for losses which are to be expected when the material

is mined or extracted.

The total proven and probable mineral reserve for the project is 3.66M short tons of material,

with an average grade of 0.47% Co, 0.68% Cu and 0.015 oz/t of Au, as shown in Table 15.2.

15.1 RESOURCE MODEL

The resource model described in Section 14.0 was used to determine the mineral resource

considered in the mine plan. This resource was then converted to the mineral reserve.

Cobalt grades for the whole or complete parent block meeting the Cut-off Grade (CoG) and

the criteria listed below, were used to determine the mineral reserve. The parent block

dimensions are defined as 2 ft. by 6 ft. by 5 ft. in the X, Y and Z directions.

There was no additional or subsequent sub-blocking performed on these parent blocks of the

resource model. This is because the parent blocks have been deemed to provide sufficient

resolution for mine design and planning and the necessary resolution to identify the interface

between mineralized and waste material. The dimensions of the parent blocks also enable the

definition of the stope outlines, and the estimation of dilution and material losses.

15.2 CUT-OFF GRADE (COG) CRITERIA AND ESTIMATE

The mineral reserve was based on the mineral resource model’s tonnages and grades, reported

from with stope outlines defined to capture, to the extent practicable, those blocks meeting or

exceeding the CoG of cobalt. The CoG value used in the mine design was based on the

operating cost and metallurgical recovery estimates that resulted from the 2015 PEA (Samuel,

2015), together with recent commodity price values.

The stope outlines and mineable tonnages and grades for longhole stoping and cut and fill

mining methods were defined based on a CoG of 0.25% Co.

Parameters and values used to determine the CoG value are presented in Table 15.1.

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Table 15.1

Cut-off Grade Criteria

Description Unit Values

Metal Prices

Cobalt Price $/lb 14.74

Copper Price $/lb 2.49

Gold $/oz 1,150

Recoveries

Cobalt – Overall % 91.08

Copper – Overall % 92.8

Gold – Overall % 78.5

Costs

Mining $/st 63.57

Milling $/st 21.93

CPF $/st 26.49

15.3 STOPE OUTLINE

The stope outlines were prepared in such a manner as to represent the planned extraction of

the mineralized zones together with any internal or adjoining waste rock which cannot be left

in-situ. The tonnage and grade contained within these stope outlines are reported as whole

blocks only, on the basis that if 50% or more of a block is within the stope outline, then the

whole block is counted as part of the mineral resource being considered in the mine plan for

conversion into the mineral reserve.

Blocks having less than 50% of their volume within the stope outlines, are excluded. In so

doing, the reported tonnage and grade for the mine plan already considers some material loss

and dilution of the mineral resource, even before any further additional modifying factors are

applied.

The stope outlines were generated from 10 ft vertical level interval shells, honoring the cobalt

CoG of 0.25%. These shells were transformed into solids and sectioned into individual stopes

of approximately 70 ft H by 300 ft L.

15.4 DILUTION AND LOSSES

Two types of dilution values were applied in determining the mineral reserve, depending on

the dip angle of the deposit, configuration of the minimum mining width and the mining

methods:

Planned or internal dilution: including all the mineralized, low grade and waste material

contained in the whole block and the stope outline.

Unplanned or external dilution: accounting for additional zero grade waste material

being included for the proposed mining methods due to the physical configuration of

the horizons and mining widths for the proposed mining methods.

126

The total planned dilution is approximately 34.6%, based on the conversion of grades from

partial to whole blocks in the mineral resource model and stope boundaries.

The unplanned dilution sources are:

Longhole stoping: One ft. drilling deviation at the top of the stope is accounted for in

the estimate as external dilution, which results in an average of 0.5 ft. overbreak on the

walls into the stope, or on average a value of 4.28 %.

Cut and fill: Source of dilution is from the waste mined in the footwall toe and top

corner of the hangingwall, depending on the width of the deposit. On average, this

accounts for an additional 9.0 % of external dilution.

The weighted average unplanned dilution is calculated to be approximately 6.1% for a

combined mining methods.

15.5 MINING RECOVERY

Mineralized material losses arise because of the difficulty of loading and mining mineralized

material from the excavated stopes. This includes losses due to fines and pillars left behind

during mining.

Sill mats are constructed and high strength paste backfill material poured into lead stopes for

both mining methods to minimize the amount of mineralized material lost as pillars or sill

pillars. Much consideration was done during the mine sequencing for the placement and

location of the high strength backfill, to reduce the amount of mineralized material abandoned

in the mine or left for extraction towards the latter years of the mine life.

The following bullets summarize the basis of the estimate for the material loss and recovery

for the mining methods:

Longhole stoping: A mineralized material recovery of 75% from the hangingwall toe

area is allowed for (i.e. 25% losses in this area) in the lead stopes. Losses accounted

for also include a 1-inch layer of fines in the sill drive of a longhole stope, as well as a

1 ft. skin pillar at the top of each column of stopes in a bottom-up sequence.

Cut and fill: An allowance of 25% of material being considered unrecoverable is made

from the hangingwall toe of all the lead stopes. The losses also include a 1-inch layer

of fines in the first (lowest) cut of the cut and fill stope, as well as a 1 ft. skin pillar at

the top most cut of each column of stopes in a bottom-up mining sequence.

The average recoveries for the longhole stoping and the cut and fill mining methods are 98.7

% and 99.8 % respectively.

127

15.6 MINERAL RESERVE ESTIMATE

The parameters discussed above were applied on a stope-by-stope basis to the designed stopes,

with the key variables being the minimum mining widths on the planned stope, the dip angle

and the mining methods selected.

Table 15.2 categorized the total mineral reserve from horizons 3021, 3022, and 3023 by class.

Table 15.2

Mineral Reserve for ICP at 0.25% Co Cut-off Grade























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16.0 MINING METHODS

16.1 MINING METHODS

The mining methods proposed for the Ram deposit are longitudinal longhole stoping and

overhand cut and fill.

The selection of these mining methods for the deposit was determined primarily by the

geometry of the mineralized horizons, including factors such as its continuity, dip and width,

and the geotechnical parameters of the rock mass.

The Ram deposit is composed of numerous parallel mineralized horizons, with thickness

ranging from one foot to more than 20 ft., at an average dip of 55° (Samuel, 2015). Currently,

only three horizons within the main zone containing the majority of the mineralization are

considered in the mine design, plan and mineral reserve. These are horizons 3021, 3022 and

3023.

Cut and fill mining will be applied to areas dipping less than 50°, or in stopes having widths

ranging from 6 to 10 ft. Conventional cut and fill mining, using hand-held pneumatic drills,

will be carried out in areas having economic mineralized width ranging from 6 to 8 ft. Areas

having widths ranging from 8 to 10 ft. will be mined by mechanized cut and fill. Horizons

wider and steeper than 10 ft. and 50° will be mined with longitudinal longhole stoping. Small

and mid-sized mining equipment was selected to provide a higher selectivity for the proposed

mining methods.

The ratio of mineral reserve that will be extracted through longhole stoping and cut and fill

mining methods is 70% and 30% respectively. In combination, these two mining methods

provide a production capacity in the underground mine that is higher than the nominal mill

capacity (800 t/d). The mine has the capacity to supply approximately 323,000 t/y of ore to the

mill during steady state operation, equivalent to approximately 880 t/d production rate for 365

d/y. The proposed mine working schedule is two 12 hours shifts for 7 days a week.

Conservatively, the mine operating cost estimates have been based on achieving this higher

rate of production, whereas in practice it is anticipated that production would keep pace with

milling, and so mining expenditures that are projected over a period of 12 years would in fact

be spread over the whole of the mill operating life of approximately 12.5 years.

Paste prepared from mill tailings will be utilised as backfill material in combination with waste

rock fill arising from mine development.

Excavated material will be hauled by 22-t payload low profile trucks to the tram loading area

located at the mine portal, and then loaded into an aerial tramway for final transportation to

the processing plant. Ore will be directed in the crushing circuits at the processing plant, while

waste will be transported to the tailings and waste storage facility.

129

16.2 MINE DESIGN PARAMETERS

The following bullet points summarize the mine design parameters and criteria for the Ram

deposit.

Cut-off Grade of 0.25% Cobalt.

Cut and Fill mining in areas with dip angle less than 50°, or in stopes less than 10 ft.

wide.

Longholes stope mining will be carried out in areas with dip angle of greater than 50°,

and stopes wider than 10 ft.

Stope vertical level intervals set at 70 ft., and stope layout generated at 300 ft. along

strike or on the YZ plane.

Only Measured and Indicated mineral resource from horizons 3021, 3022, 3023 are

considered in the mineral reserve estimate.

16.3 GEOTECHNICAL CONSIDERATIONS

The report “Underground Geotechnical Design Parameters Ram/Sunshine Deposits”, prepared

by Minefill in January 2006, presents the final recommendations for underground geotechnical

design of parameters such as safe spans, stand-up times and estimate ground support

requirements for the Ram and Sunshine deposits. That report brought together all the existing

data collected in previous years together with some additional drilling data from the Ram

deposit’s 2004 drilling campaign.

Having visited accessible areas of the Blackbird Mine on 13 July 2017, Micon is in agreement

with the technical information and findings of the 2006 Minefill report, which provides the

basis of the geotechnical engineering estimate applied in the underground mine design for the

current feasibility study.

16.3.1 Principal Rock Type

A geotechnical structural mapping campaign was carried out on the Ram deposit in 2005, with

14 mapping cells in the nearby 6930 Level adit which was developed by Noranda Inc. in

1995/1996. Geotechnical information from 295 discontinuities were recorded during this

mapping campaign. A summary of the mapping results are presented in the Table 16.1.

130

Table 16.1

Principal Structural Trends from 6930 Level Adit

Rock Type Joint Set (Dip/Dip Direction)

(J1) (J2) (J3) (J4)

All Data 49/062 39/038 82/311 58/343

Thin Bedded Siltite (TBS) 37/42 85/169

Coarse Grained Quartzite (CGQ) 61/089 71/312

Medium to Fine Grained Quartzite (MFQ) 48/062 39/039 57/344 Source: Minefill, 2006.

FCC geologists believe that the information gather from this adit is representative of the

structural regimes of the Ram deposit, because of its proximity to the proposed mine workings

(Minefill, 2006).

A majority of the Ram deposit’s data of structural discontinuities was recorded from the MFQ

rock type, accounting for 233 out of 295 data points. The joint sets at the Ram deposit

comprised of one major unit J1 at 49/062 dip to dip direction, with two other minor unit sets J3

(82/311) and J4 (58/343). Minefill (2006) reported that the structure in the TBS unit appeared

to be very similar to MFQ at the Ram deposit.

Minefill indicated that it could be expected that Joint Set 1 (J1) be a persistent problem in the

back (or roof) of the upper levels of the decline because of the structure trends parallel to the

adit based on the previous mine plan. Depending on the actual persistence and spacing of these

joints, the joint sets can be expected to form large wedges in the back of the adit, and hence

continuous ground support will be required. Joint Set 2 (J2) is also parallel to some of the major

cross cuts envisioned in the upper levels of the previous mine, particularly in stope

development cross cuts. These structures will form wedges in the back of the cross cuts and

will need artificial support to create a safe opening.

The Uniaxial Compressive Strength (UCS) for the Ram deposit varies on average from 35 to

85 MPa, and at its peak strength 85 to 285 MPa. These values were estimated from Point Load

Testing on drill cores available during year 2000 drilling campaign. An estimated 1,713 core

samples were tested. A conversion factor of value 16 was used to convert the Point Load Index

to UCS. Table 16.2 presents a summary of rock strength testing results.

Table 16.2

Rock Strength for the Ram Deposit

Note: *Grade according to ISRM (1981).

Rock Type (Code) No. of

Tests

Percentage

Sample

Avg. UCS

(MPa)

Max. UCS

(MPa) Avg. Rating*

Biotitic Tuffaceous Exhalative (BTE) 23 1.3% 40.4 121.7 Medium Strong

Coarse Grained Quartzite (CGQ) 27 1.6% 82.3 201.9 Strong

Mafic Dykes and Sills (MDS) 47 2.7% 35.2 196.6 Medium Strong

Medium to Fine Grained Quartzite (MFQ) 1421 83.0% 57.4 275.8 Strong

Quartz Vein (QTV) 25 1.5% 75.0 285.7 Strong

Thin Bedded Quartzite (TBQ) 169 9.9% 39.0 150.6 Medium Strong

Total Samples Tested 1,713 100%

131

The quartzite rock types such as MFQ and CGQ have the highest strengths, with maximum

values generally over 200 MPa and an average value ranging from 50 to 75 MPa. This is

considered as very strong rock. Some of the other rock types, such as the exhalates, appeared

to be generally of medium strength with an average UCS ranging from 30 to 50 MPa.

Minefill (2006) indicated however that there has been no laboratory uniaxial compression tests

to verify and correlate the conversion factor of 16 used in the conversion of Point Load Test

results to the UCS. The conversion factor was assumed based on Minefill’s experience at other

projects with similar weak rocks. Minefill recommends a full laboratory rock test program to

confirm the correlation factor of the point load values versus UCS.

16.3.2 Rock Quality Designation (RQD)

Most of the rock units at the Ram deposit, as presented in Table 16.3, were reported as having

low RQD values by Minefill (2006). Formation (2016) indicated that one of the reasons for the

lower RQD measurement was contributed to by mechanical breaks induced during the

extraction of the cores from the drill steels by hammering and handling. The induced

mechanical breaks could potentially be recorded as part of the RQD.

Table 16.3 presents the RQD results from year 2000 drilling, as well as the more recent data

collected during year 2004 drilling program.

Table 16.3

Average RQD Data from 2000 and 2004 for Ram Deposit

Rock Code 2004 2000 Avg. Rating

Description Max. (%) Min. (%) Avg. (%) Avg. (%)

BTE 78 30.5 37 Poor

CSD 48 10.5 Very Poor

FLT 40 5.5 Very Poor

MDS 66 22.2 35 Very Poor

MFQ 60 12.7 36 Very Poor

MUD 26 8 16.7 Very Poor

QTV 80 17.6 25 Very Poor

QTZ 100 27.6 Poor

QTZ/BTE 93 36.2 Poor

TBS 56 7.1 Very Poor

STE 77 37 58.0 Fair

Overall Avg. 65.8 22.5 25.7

Average 64.7 22.8 (Avg. excluding STE)

16.3.3 Joint Data

The longest stick measurement data recorded during year 2004 drilling program reflects poor

average RQD values, with an overall measurement of 8.2 cm. The QTZ and QTZ/BTE rock

types exhibit recorded maximum value of 72 and 40.5 cm respectively (Table 16.4).

132

The longest stick measurement or the longest intact core in a core run is a good indicator of

the average joint spacing in a rock mass, and is a good indicator to compare to the measured

RQD values. Such data has been collected by FCC since year 2000.

Table 16.4

2004 Drilling Longest Stick Measurement

Rock Code Max. (cm) Min. (cm) Avg. (cm)

BTE 26.5 9.9

CSD 3.9 2.0

FLT 9.5 5.5 7.3

MDS 11.25 3.25 7.3

MFQ 31.5 5.1

QTV 12 4.5 6.8

QTZ 72 10.8

QTZ/BTE 40.5 0.1 12.5

TBS 17 0.75 6.0

STE 18 11 14.5

Overall Average 24.2 4.2 8.2

The lower than expected measurements of stick length may have been impacted in a similar

manner to the RQD, especially if the cores were not handled properly during extraction.

The results in Table 16.4 indicate that most of the rock types have average measurement ranges

from 10 to 14.5 cm. This suggests that the cores may potentially be more competent than

initially reported. This is because RQD is estimated based on the total lengths of “sound” and

intact rock core pieces over 10.0 cm in length divided by the length of the measured core run.

16.3.4 Rock Mass Rating

The average Rock Mass Rating (RMR) for the Ram deposit ranges from 50 to 59, which is

equivalent to “Fair” rock quality. The RMR was determined by Minefill, with the factors and

parameters indicated above and assuming “wet” ground and similar joint set conditions. These

joint set conditions are planar, with rough breaks and little or no clay infill from the year 2000

drilling program (Holes R99-01 to R99-11). It included a point load test on each core run

(Minefill, 2006). Table 16.5 illustrates the Rock Mass Rating of the Ram Deposit.

Table 16.5

Rock Mass Rating Estimate for Ram Deposit from year 2000 drilling

Rock Type RMR RQD Longest Stick Count Rating Description

BTE 57 37 0.099 17 Fair

CGQ 57 44 27 Fair

MDS 59 35 0.073 16 Fair

MFQ 56 36 0.51 477 Fair

QTV 52 25 0.068 11 Fair

TBQ 50 21 26 Fair

133

Minefill reported that most of the drill holes show no correlations between the RMR and depth,

except for possibly Hole R99-01. This hole appears to show increased RMR value with depth.

This suggests that the RMR will generally be a function of rock type instead of depth.

16.4 GROUND SUPPORT RECOMMENDATIONS

16.4.1 Background

Minefill summarized two versions of ground support recommendations for ICP, based on the

results from two reports:

Underground Geotechnical Design Parameters Ram/Sunshine Deposits – Idaho Cobalt

Project dated in January 10, 2006 (Minefill, 2006).

Idaho Cobalt Project – Updated ‘Preliminary’ Ground Support Recommendations

dated in January 30th, 2006 (Minefill, 2006 b).

16.4.2 Underground Geotechnical Design Parameters Ram/Sunshine Deposits

(Minefill, 2006)

Minefill (2006) recommended that the ground support requirements for the permanent opening

(haulage, and decline) be based on the MFQ rock type unit, because it had the most data

available of 477 data points.

“Based on an average RMR of 56, and a corresponding Q value of 4.14, the MFQ unit is

predicted to have a safe unsupported span of 18 ft. in temporary mine openings. At the best

RMR of 63, the Q increases to 9.18 and the safe span increases to 25 ft.” (Minefill, 2006).

For this feasibility study, the stability of the unsupported span for man and non-man entry

proposed by Minefill are verified with the Critical Span Curve (Ouchi and Brady, 2004) and

the Stability Graph (Potvin ,1988 and Nickson, 1992). The stability of the excavations, with

respect to their rock mass quality, is concluded at the end this section.

For the temporary mine entries, such as cross-cuts and open stopes, Minefill (2006) suggested

the use of light pattern bolting provided the open spans were less than the safe spans quoted

above (e.g. 6-ft split sets on 5-ft centers). Table 16.6 presents a summary of the recommended

ground support requirement for the Ram deposit as proposed by Minefill.

134

Table 16.6

Recommended Ground Support Requirements for Ram Deposit (Minefill, 2006)

RMR – Rating Entry Type Recommended Support

Average Case – 56 Permanent – 15 ft. (4.6 m) Pattern bolting with 5-ft bolts on 3-ft centers

Mesh on backs down to shoulders

Local shotcrete where required – to 1.5-inches

Best Case – 63 Permanent – 15 ft. (4.6 m) Pattern bolting with 6-ft bolts on 5-ft centers

Spot meshing may be required

Local shotcrete where required – to 1.5-inches Source: Minefill, 2006

16.4.3 Updated ‘Preliminary’ Ground Support Recommendations

The ground support recommendation proposed by Minefill was updated on January 30th, 2006

to include re-evaluation of the preliminary work from a ‘value engineering’ perspective. This

included data collected from previous preliminary ground support analysis, as well as the data

retrieved during Minefill’s site visit from January 10 to 15, 2006 (Minefill, 2006 b).

Nevertheless, Minefill stated that even though there had been a number of earlier geotechnical

reviews conducted on the project based primarily on information collected during exploration

drilling programs, a limited amount of geotechnical data was contained in the geotechnical

data set having limited supplementary adit mapping data and performed in 2005. The current

geotechnical data set therefore continues to “preclude the provision of detailed ground support

guidelines” (Minefill, 2006 b).

16.4.3.1 Background

The additional supplementary data set and analysis included in this updated ‘preliminary’

ground support recommendations include:

Detailed geotechnical logging carried out in the 2004 drill core runs which intercepted

high grade mineralization (Co > 1.0%) to characterize the HW and FW rock masses.

Data collected included RQD, lithology, rock strength, and discontinuity conditions

and joint spacing.

Historical data from various drill logs was also reviewed and statistically analyzed, and

core photos were also reviewed.

Geotechnical data collected from diamond drilling completed after the provision of the

previous support recommendations was also reviewed and analyzed. This included data

from the 2005 exploration program.

The geotechnical data and parameters collected in the 2004 drilling conformed to the Rock

Mass Rating (RMR) and the Rock Tunneling Quality Index (Q) classification systems as

proposed by Bieniawski and Barton (although it was not clear in Minefill’s reports which

referencing year the RMR rating was based on). Both these classification systems are standard

industry rock mass classification systems for mining and civil engineering. Minefill indicated

that the Q system has not been used directly in the past to collect geotechnical data for the

135

project. The Q values from the previous work was estimated by Minefill (2006 b) based the

following correlation between RMR and Q as proposed by Bieniwaski (1993):

RMR76 = 9 Loge Q + 44

The Q values for this updated ‘preliminary’ ground support pertains only to the mineralized

zones (of varying lithology, but primarily Quartzite), while the RMR values pertain to the MFQ

(Quartzite) lithology. Minefill indicated that despite these data sets being not completely

comparable, the analysis is nonetheless considered valid given the pervasiveness of this rock

type. For comparison purposes, the Equivalent RMR (Eq. RMR) values determined from the

Q values recorded in 2004 are presented in the Table 16.7 and compared with the RMR values

collected in year 2000.

Table 16.7

Summary of Rock Mass Classification for Mineralized Zones and MFQ at ICP

Description Lower Bound Average Upper Bound

Qavg. 1.7 (Poor) 6.6 (Fair) 14 (Good)

Eq. RMR 49 (Fair) 61 (Good) 68 (Good)

RMR (Previously Reported in year 2000) 56 (Fair) 63 (Good)

Minefill reported that the Q values collected from the 2004 drilling are explicitly determined

and are higher than those previously determined and reported using the ‘Eq. RMR’ calculation

method.

16.4.3.2 Updated ‘Preliminary” Ground Support

Minefill utilized a statistical analysis approach to determine the percentile distribution for the

Q values and concluded that:

20% of the mineralized zones having a Q value in the “Lower Bound” range (Poor

rock).

40% of mineralized zones have an “Average” Q value (Fair rock).

40% of mineralized zones have Q values in the “Upper Bound” range (Good rock).

Minefill (2006 b) also indicated that these percentages are approximations, and needed to be

confirmed through further geotechnical investigation and analysis.

Table 16.8 and Figure 16.1 summarize the updated ‘preliminary’ ground support for Permanent

Mine Openings recommendation made by Minefill (2006 b) for ICP, based on Q rock mass

classification system proposed by Grimstad and Barton (1973).

Minefill (2006 b) recommended that for temporary mine openings, such as cross cuts and open

stopes, light patterned bolting be carried out (e.g. 6-8 ft. split sets on 5-7 ft. centers, with mesh

as required), provided the open spans remain less than 20 to 25 ft.

136

Table 16.8

Updated ‘Preliminary’ Ground Support Recommendations for Permanent Openings for ICP

Q Rating Entry Type Recommended Ground Support Remarks

(Micon)

Lower Bound

Qavg. = 1.7

(20%)

Permanent

15 ft. Span

(De = 2.85)

Pattern bolting with 6-8 ft. bolts on 5 ft.

centres.

Mesh on back down to shoulders.

Local shotcrete where required – to 1.5

inches.

Poor Rock

Quality

Average Conditions

Qavg. = 6.6

(40%)

Permanent

15 ft. Span

(De = 2.85)

Pattern bolting with 6-8 ft. bolts on 6 ft.

centres.

Mesh on back down to shoulders, as required.

Fair Rock

Quality

Upper Bound

Qavg. = 14

(40%)

Permanent

15 ft. Span

(De = 2.85)

Spot/localized pattern bolting with 6-8 ft.

bolts on approx. 7 ft. centres.

Mesh on back down to shoulders, as required.

Good Rock

Quality

Source: Minefill, 2006 b

Figure 16.1

Estimated Ground Support for ICP

16.4.4 Conclusion – Geotechnical Consideration

A simplified and standardize ground support requirement was formulated in this feasibility

study, based on the information documented by Minefill. This standardization is a bolting

pattern of 5 ft. x 5 ft., with five 6 ft. length frictional bolts pattern supported with welded wire

137

mesh and installed down to approximately 5 ft. off the ground on the side walls for all the

underground excavations.

Additional provision of 10% for shotcreting of the main decline and ramp was accounted for

in the mining cost, along with additional cost for rehabilitation of one intersection per year

with 12 ft. L connectable Swellex bolts and shotcreting accounting for the very poor ground

conditions during mine development. The underground mechanics shops will be supported by

8 ft. L Swellex with welded wire as primary support, 12 ft. L Swellex as secondary support,

and 4-inches of shotcrete. Further optimization of this support requirement can be evaluated in

the detailing engineering and construction stage of the project.

The stability of the span or excavation for man-entry working areas or in cut and fill stopes for

the proposed widths and RMR rating is suggested by Minefill, and the stope generated in this

feasibility study is verified by the Critical Span Curve (Ouchi and Brady, 2004). Results of this

preliminary analysis is presented below and in Figure 16.2.

Excavations having a width of 15 ft. (4.5 m) and an average RMR of 56 are at the limit of

being in the “Stable” zone, while excavation spanning 25 ft. (7.6 m) with the same RMR value

is considered “Potentially Unstable”.

Currently, the minimum and maximum cut and fill stope width are 6.3 and 21.8 ft. (1.86 and

6.6 m). Excavation at 6.3 ft. is “Stable” while excavations having width of 21.8 ft. is at the

limit between a “Stable” to “Potentially Unstable” zone for the average RMR of 56.

Figure 16.2

Weak Rock Mass Design Span Curve for Man Entry (Ouchi and Brady, 2004)

The stability of the open stopes for longitudinal longhole stopping was also evaluated with the

Stability Graph for unsupported stopes proposed by Potvin (1988) and Nickson (1992).

Unsupported stopes of 70 ft. H with 150 ft. L are classified as being “Stable” and stopes having

the same height with a strike length of 300 ft. lies close to the rim of “Transition” to “Stable”

zones (Figure 16.3), based on this preliminary analysis.

138

Critical Span Curves and Stability Graphs are empirical approaches used to determine

stabilities of excavations in the Ram deposit. Currently, the maximum span and strike length

for the average RMR locates the excavation at the limit of stable zone. This however, does not

mean that stabilities of the excavations are jeopardize, especially when they are located at or

close to the boundary of stable to potentially unstable zones. The assessment above was made

based on available information. Additional geotechnical investigation will assist in determined

the maximum span of the excavations.

Figure 16.3

Stability Graph (Unsupported Stopes)

16.5 MINING CUT-OFF GRADE AND SPECIFICATIONS

The mining stopes are generated based on a CoG of 0.25% cobalt, which takes into account

recoveries and cost estimates values from the 2015 PEA, and recent cobalt spot prices (see

Table 15.1.

Stope outlines were generated at 10 ft. vertical interval, transformed into solid and sectioned

by 70 ft. H by 300 ft. L generating individual mining stopes. Details on the CoG criteria is

presented in Section 15.0.

139

16.6 SELECTIVITY, DILUTION AND RECOVERY

16.6.1 Mining Selectivity

Cut and fill mining will be carried out in areas with a dip angle of less than 50°, or in stopes

with width ranging from 6 to 10 ft. Areas having widths of less than 8 ft. will be mined with

pneumatic handheld drills, while stopes greater than 8 ft. wide will be mined with single boom

jumbos. Longholes stope mining will be carried out in areas steeper than 50° and with widths

greater than 10 ft.

These proposed mining methods will provide the mine with the flexibility to mine from

moderately dipping and narrow horizons to steeper and wider horizons.

16.6.2 Dilution

Planned dilution accounts for all the material which is contained within blocks having centroids

that lie within the design stope boundaries, which are determined by the selectivity of the

mining methods and the continuity of the orebody. The total value of planned dilution is

approximately 34.3%. This value was estimated from a comparison of the undiluted grades

with diluted block grades for those blocks lying within the stope boundaries.

The unplanned dilution, however, arises primarily due to imprecision of the mining operation.

One foot of drill-hole deviation was incorporated in the estimate for dilution values in longhole

stoping mining method. This equates to an average of 4.3% dilution related to this mining

method. The sources of unplanned dilution from the cut and fill include waste rock extracted

from the walls of the cut, the percentage depending on the dip and width of the stope. This

accounts for a total of 9.0% of all the cut and fill stopes.

The average unplanned dilution for the ICP project is approximately 6.1% for the combination

of mining methods.

16.6.3 Mining Recovery

The mining recovery was estimated based on the difficulty of mining, loading or recovery of

the blasted material from the mining stopes.

The average recoveries from the longhole stoping, and cut and fill mining methods are 98.7%

and 99.8% respectively. These values were determined on the basis that 25% of material is

unrecoverable from the hangingwall toes of the mining method in the lead stopes.

The losses also include a 1-inch of unrecoverable fines on the sill drives of the lead longhole

stopes and the initial lift of the cut and fill stopes. Skin pillars of 1 ft. thick are also considered

to be unrecoverable at the top most cut of each stope in the bottom-up mining sequence.

140

16.7 MINE DESIGN

The mine design was developed to support a mine production rate of approximately 800 t/d for

the proposed mining methods. High grades stopes from horizon 3023 were given priority

during the mine production scheduling, and the remaining horizons were accordingly

scheduled into the sequence based on their grades.

16.7.1 Underground Excavation Dimensions

All the main underground development was designed to a cross section area of 14 ft. H x 12

ft. W, except for the main decline, ramp, safety bays situated in these excavations, explosive

bays and the mechanics shop. Main ventilation shafts will be excavated with a raise bore and

the remaining ventilation raises in between level to level will be excavated by drilling and

blasting.

Table 16.9 summarizes the estimated Life of Mine (LoM) development length and the

proposed excavation dimensions.

Table 16.9

Estimated Mine Development Distance

Estimated Development Footage Dimension

(H ft x W ft) Total LoM (ft.)

Ramp, Decline & Safety Bays 15 x 13 15,383.7

Access Drive 14 x 12 3,574.1

Haulage Drift 14 x 12 15,459.8

Safety Bay on Haulage 14 x 12 585.0

Cross Cuts 14 x 12 11,058.5

Remuck 14 x 12 2,070.0

Attack Ramps 14 x 12 24,669.7

Ventilation Drift 14 x 12 2,410.6

Explosive Bay 15 x 13 85.1

Mechanical Shop 20 x 16 212.8

Main Sump 14 x 12 42.6

Secondary Sumps 14 x 12 375.0

Backfill Sumps 14 x 12 270.0

Ventilation Raise 9 ft dia. 2,294.6

Sub-Total Hor. Dev. 76,196.8

Sub-Total Vert. Dev. 2,294.6

Total 78,491.4

16.7.2 Mine Access

The main decline and a system of ramps provide access to the underground workings and

production areas. There are two portals into the mine: one as the main decline providing access

into the mine production heading, and the other acting as the service tunnel.

141

The service tunnel provides access to main underground services and storage areas such as the

mechanics shops, ventilation exhaust shaft, main sump, explosive storage and where the paste

backfill boreholes breaks through from the surface. The explosive storage areas are located in

a lateral drift connecting the main decline to the service tunnel. Ventilation and fire doors or

bulkheads are placed to prevent short-circuiting of the ventilation system and for fire control.

Both portals are located at approximately 7080 ft. elevation.

The decline is approximately 513 ft. L and designed at -12.5% grade. The remaining ramp in

between levels is designed at -15% grade. Muck bays of 15 ft. L are located at approximately

the mid-point of the level to level ramp system. These bays will be converted into safety bays

or vehicle passing bays during operation. Figure 16.4 shows the mine development layout.

Figure 16.4

ICP Mine Development Layout

16.7.3 Underground Mine Layout

Currently, all the underground development and access into the stope is located in the

hangingwall, enabling a better location for additional exploration drilling to be carried out and

facilitating longhole stope definition drilling.

Access into the production stopes will be from hangingwall and will be through a series of

lateral development with access drives connecting the decline or the ramp to the haulage drifts,

cross cuts and attack ramps into the longhole and cut and fill stopes. Figure 16.5 shows the

stope layout.

142

Figure 16.5

ICP Stope Layout

16.8 MINE DEVELOPMENT AND PRODUCTION SCHEDULE

As noted above, the mine development and production schedules were generated for an overall

rate of ore production approximately 10% higher that the nominal mill capacity of 800 t/d.

This assumption is conservative in that forecast mining expenditures are incurred earlier than

would be required in order to achieve the milling schedule. In the project cash flow model, the

extent of this advanced mining is reflected as a notional mill-feed stockpile, whereas in practice

the rate of underground ore production would be scaled back to keep pace with milling, and so

mining expenditures that are projected over a period of 12 years would in fact be spread over

the whole of the mill operating life of approximately 12.5 years.

The mine development and production from pre-production period year -1 (Y-1) into the year

2 of operation will be carried out by Small Mine Development (SMD), an underground mining

contractor. The remaining mine excavations and production following this period will be

performed by the owner’s mining crew.

The mining sequence commences with the extraction of lead stopes from the bottom of the

sequence. There will be two sills in the lead stopes for longhole stoping mining method. The

initial sill acts as a sill pillar where a sill mat and high strength pastefill will be constructed and

placed. The second sill provides the working area to mine the stope and consecutive stopes

above. Lead stopes in the cut and fill stopes will only have one lead stope where a sill mat and

high strength pastefill will be placed in the initial cut.

The objective for incorporating the sill mat and high strength pastefill into the mining methods

is to enable higher recovery of the mineralized material in horizons.

143

16.8.1 Mine Development

The mine development commences during the pre-production Y -1, focusing on the excavation

of the primary access into an initial cut and fill mining stope located on elevation 7156 ft.

elevation, and accessing the longhole stope on elevation 6876 ft. to initiate the initial series of

bottom-up mining. The mine ventilation shafts will serve as secondary escape-ways when

production commences, and will be connected via ventilation drifts to the haulage drives.

The advance rate for single development heading with mechanized mining equipment is

approximately 10.0 ft./round for single headings and 15 ft./round for multiple headings. The

advance rate for manual mining is estimated to be approximately 6.0 ft./round.

Mucking bays will be situated at approximately 150 to 250 ft. along the haulage drive. The

distance in between these bays is determined by the location of other underground excavations

along the haulage drive during development. For instance, a cut for cross cuts or ventilation

drives can be used as a temporary muck bay during the level development prior to production,

until the next muck bay is excavated. These bays can be turned into safety bays, vehicle passing

area or even as temporary storage areas during operations. The mine development and layout

includes a total of 25 dewatering and 15 backfill sumps.

Access into the stopes from the haulage drive includes a lateral development of 50 ft. followed

by approximately 160 ft. of cross cuts before reaching the stope. Remucks are located at the

mid-point of the lateral development before the cross cuts. In cut and fill stopes, the initial

cross cut will be the initial stope access to initiate the attack ramps for the subsequent lift to

mining the stope. The initial cross cut in this case will be excavated at -15% grade, followed

by subsequent attack ramps development upwards to a maximum grade of +17%. Cross cuts

into longhole stopes will have minor downward grade towards the haulage drive facilitating

drainage out of the stoping areas.

The LoM annual mine development footages and tonnage summaries are presented in Table

16.10 and Table 16.11.

144

Table 16.10

Estimated Mine Development Summary (Footage)

Development Footage Dimension

(H ft x W ft)

Period/

Unit Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12

Ramp, Decline &

Safety Bays

15 x 13 ft. 5,429 4,160 3,986 1,788 20 - - - - - - - -

Access Drive 14 x 12 ft. 838 1,294 1,126 316 - - - - - - - - -

Haulage Drift 14 x 12 ft. 2,497 6,908 5,136 918 - - - - - - - - -

Safety Bay on Haulage 14 x 12 ft. 140 249 137 60 - - - - - - - - -

Cross Cuts 14 x 12 ft. 160 961 1,601 1,441 1,566 801 801 801 629 320 870 949 160

Remuck 14 x 12 ft. 30 180 300 270 293 150 150 150 118 60 163 178 30

Attack Ramps 14 x 12 ft. - 2,295 1,072 2,370 1,772 1,200 3,261 2,106 2,171 2,467 1,940 2,290 1,725

Ventilation Drift 14 x 12 ft. 84 1,949 251 128 - - - - - - - - -

Explosive Bay 15 x 13 ft. 85 - - - - - - - - - - - -

Mechanical Shop 20 x 16 ft. 213 - - - - - - - - - - - -

Main Sump 14 x 12 ft. 43 - - - - - - - - - - - -

Secondary Sumps 14 x 12 ft. 61 168 125 22 - - - - - - - - -

Backfill Sumps 14 x 12 ft. - 75 105 60 15 - - - - - 15 - -

Ventilation Shaft and

Raise

9 ft. dia ft. - 1,038 852 343 61 - - - - - - - -

Sub-Total Hor. Dev. ft. 9,578 18,238 13,839 7,373 3,667 2,150 4,212 3,056 2,918 2,847 2,988 3,417 1,915

Sub-Total Vert. Dev. ft. - 1,038 852 343 61 - - - - - - - -

Total ft. 9,578 19,276 14,691 7,716 3,728 2,150 4,212 3,056 2,918 2,847 2,988 3,417 1,915

145

Table 16.11

Estimated Mine Development Summary (Tonnage)

Development

Tonnage

Total or

Average

(LoM)

Period

or Unit Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12

Tonnage

(>=0.25% Co) 9,904 t 233 799 1,346 1,681 2,048 813 786 501 885 44 282 354 132

Cobalt Grade 0.586 % 0.411 0.850 0.842 0.641 0.546 0.547 0.481 0.404 0.366 0.331 0.322 0.459 0.623

Copper Grade 0.917 % 0.043 1.124 1.588 0.935 0.566 1.039 0.902 0.304 1.352 0.659 0.515 0.532 0.324

Gold Grade 0.020 oz/t - 0.031 0.035 0.021 0.014 0.022 0.015 0.009 0.013 0.021 0.011 0.017 0.029

Cobalt Metal 116,088 lb 1,919 13,586 22,658 21,543 22,378 8,900 7,569 4,047 6,487 294 1,816 3,245 1,644

Copper Metal 181,649 lb 201 17,958 42,736 31,411 23,177 16,904 14,182 3,043 23,933 586 2,901 3,762 855

Gold Metal 196 oz - 25 48 36 28 18 12 4 12 1 3 6 4

Tonnage

(<0.25% Co) 398,187 t 56,670 131,153 104,041 30,040 20,402 10,269 7,815 11,185 4,515 1,797 8,055 9,934 2,310

Cobalt Grade 0.029 % 0.018 0.030 0.025 0.036 0.031 0.037 0.037 0.042 0.052 0.028 0.033 0.039 0.068

Copper Grade 0.077 % 0.058 0.089 0.049 0.066 0.070 0.137 0.170 0.132 0.102 0.133 0.090 0.136 0.173

Gold Grade 0.001 oz/t 0.001 0.001 0.001 0.001 0.001 0.002 0.002 0.002 0.002 0.002 0.002 0.003 0.006

Cobalt Metal 231,255 lb 20,334 79,559 52,639 21,512 12,464 7,515 5,841 9,444 4,706 1,008 5,293 7,793 3,147

Copper Metal 617,114 lb 66,021 233,766 101,360 39,703 28,536 28,096 26,610 29,576 9,182 4,794 14,557 26,939 7,974

Gold Metal 559 oz 80 161 118 41 27 21 14 23 9 3 19 29 13

Waste 485,988 t 163,409 229,928 84,633 8,018 - - - - - - - - -

Total Dev.

Material 894,080 t 220,313 361,880 190,020 39,738 22,450 11,083 8,601 11,686 5,401 1,842 8,337 10,288 2,442

146

Mine development material from the proposed mining horizons and other horizons having

average grade greater and equal to 0.25% Co and that are generated from the Measured and

Indicated mineral resource are also considered in the mineral reserve estimate (i.e., 9,904 t at

0.58% Co).

Over the life of the mine, underground development generates approximately 694,000 t of

waste, and part of this (around 208,000 t) will be used as backfill, to reduce transportation and

pastefill costs. The remaining 486,000 t of waste material will be transported to the tailings

and waste management facility.

16.8.2 Production Schedule

Mining will commence with an initial extraction of a cut and fill stope at elevation 7156, closest

to the main decline, followed by the mining of longhole stopes in a bottom up sequence on

elevation 6876.

Sill mat and high strength pastefill placed and poured into the lead stopes enable higher

recovery of the mineralized horizons and safer working area during the extraction of stopes

beneath the backfilled stopes. An allowance of 14 and 28 days backfill curing days for cut and

fill, and longhole stoping were incorporated during the mine sequencing.

Stopes having dip angles of less than 50° will be mined by cut and fill mining methods.

Mechanized cut and fill mining methods will be applied in areas with widths ranging from 8

to 10 ft, and conventional cut and fill will be performed by handled pneumatic drills and

mucked with small LHDs. Longitudinal longhole stoping will be applied to locations with dip

angles greater than 50° and widths greater than 10 ft.

The stopes are designed at 70 ft. H by 300 ft. L. Cut and fill stopes will be mined by horizontal

lifts of 14 ft. H, with advance rate of either 6 or 10 ft. per round depending on the width of the

deposit. On average, there are 5 lifts per cut and fill stope. Longhole drilling, production

blasting and mucking for longhole stoping are carried out from the sills. Longhole stope

production commence with the blasting of the slot raise and mining of the remaining

mineralized material in vertical slices.

The mining sequence begins with the extraction of the high-grade material. Priority is given to

material in horizon 3023 where possible, and retreating into horizon 3022 followed by the

mining of horizon 3021.

Mine production starts in year 1 (Y1) at a ramp-up production rate of approximately

209,000 t/y, followed by a steady state production rate averaging 323,000 t/y and a ramp-down

rate in the final year of 220,000 t/y.

Table 16.12 present the mine production schedule along with development tonnages having

grades above 0.25% Co. Figure 16.6, presents a schematic of the mine and stope layout.

147

Table 16.12

Mining Production Schedule

Tonnage Total/Ave.

(LoM)

Period

/Unit Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12

Cut & Fill 1,080,120 t - 100,471 47,416 103,746 77,664 52,391 142,522 92,131 94,735 107,768 84,959 100,749 75,567

Cobalt Grade 0.423 % - 0.717 0.645 0.606 0.630 0.445 0.409 0.396 0.329 0.299 0.237 0.223 0.237

Copper Grade 0.367 % - 0.759 0.152 0.674 0.295 0.369 0.422 0.266 0.167 0.234 0.220 0.387 0.228

Gold Grade 0.013 oz/t - 0.031 0.023 0.011 0.018 0.009 0.009 0.009 0.013 0.008 0.006 0.006 0.012

Cobalt Metal 9,132,819 lb - 1,441,355 611,874 1,258,395 978,760 466,788 1,167,146 729,938 622,995 645,357 402,774 449,624 357,813

Copper Metal 7,925,692 lb - 1,525,301 144,173 1,399,102 458,905 387,010 1,203,130 489,915 315,653 505,060 373,582 779,164 344,699

Gold Metal 13,686 oz - 3,110 1,069 1,193 1,415 490 1,318 874 1,239 871 547 646 914

Long Hole Stoping 2,571,870 t - 107,464 276,733 227,011 259,861 263,367 175,669 225,972 223,716 219,508 230,644 217,968 143,957

Cobalt Grade 0.491 % - 0.850 0.794 0.750 0.598 0.455 0.512 0.386 0.374 0.301 0.288 0.333 0.282

Copper Grade 0.810 % - 0.833 1.106 1.176 0.407 0.725 0.920 0.541 0.871 0.941 0.932 0.701 0.486

Gold Grade 0.015 oz/t - 0.022 0.024 0.022 0.011 0.011 0.018 0.009 0.009 0.013 0.015 0.018 0.015

Cobalt Metal 25,268,018 lb - 1,826,390 4,395,969 3,404,142 3,108,386 2,398,320 1,798,467 1,744,033 1,675,587 1,322,322 1,328,191 1,453,220 812,990

Copper Metal 41,648,204 lb - 1,791,303 6,120,952 5,339,716 2,115,888 3,821,032 3,233,121 2,446,429 3,895,676 4,129,719 4,298,377 3,055,952 1,400,039

Gold Metal 39,404 oz - 2,363 6,561 5,087 2,828 2,969 3,090 2,024 2,078 2,844 3,500 3,973 2,088

Dev. Tonnage 9,904 t 233 799 1,346 1,681 2,048 813 786 501 885 44 282 354 132

Cobalt Grade 0.586 % 0.411 0.850 0.842 0.641 0.546 0.547 0.481 0.404 0.366 0.331 0.322 0.459 0.623

Copper Grade 0.917 % 0.043 1.124 1.588 0.935 0.566 1.039 0.902 0.304 1.352 0.659 0.515 0.532 0.324

Gold Grade 0.020 oz/t - 0.031 0.035 0.021 0.014 0.022 0.015 0.009 0.013 0.021 0.011 0.017 0.029

Cobalt Metal 116,088 lb 1,919 13,586 22,658 21,543 22,378 8,900 7,569 4,047 6,487 294 1,816 3,245 1,644

Copper Metal 181,649 lb 201 17,958 42,736 31,411 23,177 16,904 14,182 3,043 23,933 586 2,901 3,762 855

Gold Metal 196 oz - 25 48 36 28 18 12 4 12 1 3 6 4

Total Tonnage 3,661,894 t 233 208,734 325,495 332,438 339,573 316,571 318,978 318,604 319,336 327,320 315,885 319,071 219,656

Cobalt Grade 0.471 % 0.411 0.786 0.773 0.705 0.605 0.454 0.466 0.389 0.361 0.301 0.274 0.299 0.267

Copper Grade 0.679 % 0.043 0.799 0.969 1.018 0.383 0.667 0.698 0.461 0.663 0.708 0.740 0.602 0.397

Gold Grade 0.015 oz/t - 0.026 0.024 0.019 0.013 0.011 0.014 0.009 0.010 0.011 0.013 0.014 0.014

Cobalt Metal 34,516,925 lb 1,919 3,281,331 5,030,501 4,684,081 4,109,524 2,874,008 2,973,182 2,478,018 2,305,070 1,967,973 1,732,780 1,906,089 1,172,447

Copper Metal 49,755,545 lb 201 3,334,561 6,307,861 6,770,229 2,597,970 4,224,945 4,450,434 2,939,387 4,235,261 4,635,365 4,674,860 3,838,878 1,745,593

Gold Metal 53,286 oz - 5,498 7,678 6,315 4,271 3,477 4,419 2,902 3,329 3,716 4,050 4,624 3,006

148

Figure 16.6

Schematic of the Mine and Stope Layout

149

The number of active headings and stopes vary depending on the mining horizons, widths of

the horizons and mining methods. A minimum of 3-5 active production faces are required for

cut and fill mining while only one active stope is necessary to meet the daily production target.

In minimum, it is envisioned that similar amount of production areas or stopes should be in

preparation in order to meet the production cycle and to provide mining grade selectivity.

16.9 MANPOWER REQUIREMENTS

The manpower and mine labour requirement are supplied by SMD during the mine

development and production up to year 2. ICP will also have their mine staff during this period

working with and supervising the work carried out by the contractor. The owner will mobilize

its mining crew in year 3. The manpower requirements were estimated based on productivities,

capacities and availabilities of the equipment.

The mine staff and labour for ICP is listed in Table 16.13.

150

Table 16.13

Mine Staff

Mine Staff Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12

Mine Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1

Chief Engineer 1 1 1 1 1 1 1 1 1 1 1 1 1

Safety Foreman 1 1 1 1 1 1 1 1 1 1 1 1 1

Surveyor/Rodman 2 2 2 2 2 2 2 2 2 2 2 2 2

Mine Engineer 1 1 1 1 1 1 1 1 1 1 1 1 1

Chief Geologist 1 1 1 1 1 1 1 1 1 1 1 1 1

Geologist 1 1 1 2 2 2 2 2 2 2 2 2 2

Technicians 1 1 1 1 1 1 1 1 1 1 1 1 1

Shift Foreman 1 1 1 2 2 2 2 2 2 2 2 2 2

Clerk 1 1 1 1 1 1 1 1 1 1 1 1 1

Maintenance Coordinator 1 1 1 1 1 1 1 1 1 1 1 1 1

Total Mine Staff 12 12 12 14 14 14 14 14 14 14 14 14 14

151

Table 16.14

Underground Mine Labour

Mine Labour Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12

Anfo Basket - - - 2 2 2 2 2 2 2 2 2 2

FEL Operator - - - 2 2 2 2 2 2 2 2 2 2

2 Boom Jumbo - - - 1 2 2 2 2 - 2 2 2 2

1 Boom Jumbo - - - 1 2 2 2 2 - 2 2 2 2

Longhole Drill - - - 1 1 1 1 1 1 1 1 1 1

Explosives Truck - - - 1 1 1 1 1 1 1 1 1 1

Forklift - - - 2 2 2 2 2 2 2 2 2 2

SkyTrak - - - 6 6 6 6 6 6 6 6 6 6

LHD 2.5 cu. yd. - - - 4 3 2 4 2 2 2 2 2 2

LHD 4.0 cu. yd. - - - 6 4 4 4 4 4 4 4 4 2

LHD 1.5 cu. yd. - - - - - - - 1 - - 1 1 1

Small Bolter - - - 1 1 1 1 1 - 1 1 1 1

1 Boom Bolter - - - 4 2 2 2 2 2 2 2 2 2

Grader - - - 1 1 1 1 1 1 1 1 1 1

Scissorlift - - - 2 2 2 2 2 2 2 2 2 2

Service truck - Lube - - - 2 2 2 2 2 2 2 2 2 2

Transmixer - - - 1 1 1 1 1 1 1 1 1 1

UG truck - - - 4 4 4 4 4 4 4 4 4 4

Shotcrete sprayer - - - 1 1 1 1 1 1 1 1 1 1

Articulated Truck 25 yd3. - - - 1 1 1 1 1 1 1 1 1 1

Pneumatic Handheld Drill - - - - - - - 2 2 2 2 2 2

Backfill & Gen. Service Crew - - - 2 2 2 2 2 2 2 2 2 2

Mechanic - - - 6 6 6 6 6 6 6 6 6 6

Electrician - - - 4 4 4 4 4 4 4 4 4 4

Total Mine Labour - - - 55 52 51 53 54 48 53 54 54 52

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16.10 EQUIPMENT SELECTION

The mining contractor will supply all the required mining equipment and manpower during

the mine development, pre-production and up to year 2 of operation. The estimated mining

equipment proposed SMD and the purchasing of these equipment from the contractor by the

year 3 is summarized in Table 16.15.

The list of equipment is not divided into years of operation because this is the complete fleet

of equipment purchased from the contractor and available on site. The operating cost of each

individual mining equipment, however are broken down and estimated based on their working

hours. This is presented in operating cost section.

Table 16.15

Mining Equipment List

Description Quantity

Anfo Basket 1

Buggy / Tractor 4

FEL 1

Dozer 1

2 Boom Jumbo 1

1 Boom Jumbo 1

Longhole Drill 1

Explosives Truck 1

Forklift 1

Skytrak 3

LHD 2.5 cu. yd. 2

LHD 4.0 cu. yd. 2

LHD 1.5 cu. yd. 1

Small Bolter 1

1 Boom Bolter 1

Grader 1

Scissorlift 1

Service truck - Lube 1

Transmixer 1

UG truck - 22st 4

Shotcrete sprayer 1

Articulated Truck 25 cu. yd. 1

Pneumatic Handheld Drill 4

Air Compressor 1

Vent. Fans 3

Dewatering Pumps 11

16.11 UTILITIES, SERVICES FOR UNDERGROUND

16.11.1 Temporary Mine Area Building

SMD will set up a temporary surface structure on the portal area during the pre-production

stage of the mine development to provide equipment maintenance and service until the

153

underground mechanic shop and infrastructure are completed. The underground mine office

can be located in one of the five underground mechanical bay.

16.11.2 Explosive Storage

The main explosive storage is located on the surface with temporary underground storage

located in between the connecting drift between the main decline and service tunnel. Blasting

supplies and explosives will be transported from surface to the underground storage facility

and distributed to the underground development and production area.

16.11.3 Underground Communication System

The communication system will be via telephone and leaky feeder system. The leaky feeder

system can also be used to control the fans and pumps. All mobile mining equipment are

equipped with two-way radio system.

16.12 VENTILATION

The main decline and ramp will mainly supply fresh air into the development and production

areas. Ram deposit ventilation network will be composed of a series of ventilation drifts

connecting underground development to the ventilation shafts and raises. Ventilation along

lateral development and in production areas will be supplied and controlled by a combination

of regulators, ducting and auxiliary fans.

There are three main ventilation shafts into the Ram deposit: the main shaft located in the south

zone, a secondary shaft situated in between the south and north deposit, and the final shaft

positioned in the north zone. The mid and north shafts daylight to the surface while the main

shaft breaks through into the service tunnel at elevation 7040.

The ventilation shafts will be excavated by the contractor with a raise borer to 9 ft diameter.

Ventilation raises in between levels will be excavated with the conventional drilling and

blasting. All the mine ventilation shafts and raises at the Ram deposit also serves a secondary

emergency escapeways.

Fresh air enters the mine through the portal, travels down the main decline and ramp system

before splitting between the south or north mining areas. Bulkheads and ventilation doors

situated between the ventilation raise and drifts prevent recirculation of exhaust air. The

ventilation shaft and raise in the middle of the deposit can be utilized as fresh intake or exhaust

during operations depending on the demand for ventilation during mining.

The mine ventilation demand for the RAM deposit was estimated based on the ventilation

design prepared by Mine Ventilation Services, Inc. in 2005 (MVS). This is considered

reasonable because the extent of the proposed mine layout is very similar to the earlier design

proposed by MDA in December, 2006.

154

The development areas of the mine are estimated to require a minimum 66.2 kcfm while

production and stope preparation areas require approximately 49.5 kcfm. These estimates were

based on the ventilation demand specified based on the break horsepower of the operating

mining equipment and assumptions stated above. However, MVS stated that the estimated

amount of air required for active areas of the mine should be 70 kcfm and 50 kcfm for mine

development and production areas, to account for leakages and other design factors.

The mine decline and ramp ventilation demand was designed to support a single underground

haul truck, a 3.5 yd3 LHD and a utility vehicle. Production area and stope preparation were

anticipated to contain one haul truck and a 2.0 yd3 LHD scoop in operation. The airflow

requirement for the vehicles were based on their break horsepower and assumed utilization

factors. It is assumed that all diesel powered vehicles will be turned off when they are not in

use for extended of period of time. SMD has indicated that its mining equipment is well

maintained and meets diesel particulate matter emissions limits and it regularly conducts

emission audits on its mining equipment fleet. The manufacturing years of the mining

equipment fleet currently proposed by SMD range from 1983-2012 for LHDs, between 2002-

2006 for underground trucks, and 2015-2016 for the buggy and tractors. Consequently, there

may be opportunities to optimize the demand for ventilation underground during further

development stages of the project.

Currently, there are five fans proposed: three located underground and two exhaust fans located

on the surface. Figure 16.7 shows the schematic of ICP mine ventilation layout.

Figure 16.7

Schematic of ICP Ventilation Layout

16.13 BACKFILL SYSTEM

The principal method of backfill at ICP is pastefill with a combination of waste material

generated from the mine development. Summary of the annual backfill placement is shown in

Figure 16.8.

155

At the present stage, the incorporation of waste as backfill material contributes to

approximately $2.68 M saving in binder and cement costs excluding the accounting of waste

material transportation costs. These material, however will have to be place in conjunction

during the pastefill pours to ensure homogenous matrix between the pastefill to waste material.

Figure 16.8

Backfill Schedule and Material Source.

The pastefill prepared in the backfill plant located at the processing plant is routed through an

overland pipeline along or within the vicinity of the tramline alignment corridor to

approximately elevation 7445 ft. From there, the backfill material is directed with cased

boreholes into the service tunnel. Pastefill from the service tunnel is routed into the mines

through a series of inter-level boreholes and along haulage levels to final discharge points in

the stopes. Currently, two main delivery lines and two cased boreholes are proposed for the

backfill system at ICP: one operating and the other on-standby.

The paste backfill design criteria, hydraulic design, pump recommendations and control

philosophy developed by Paterson and Cooke (P&C) in 2012 for the construction of the mine

at the time is incorporated to the current study. This is because the pastefill pump, mixer,

filtering system and a majority of the backfill plant system had been purchased, and there is

close similarity of the current mine layout to the layout by P&C in 2012.

However, additional pastefill strength testing with two types of binders and slag were carried

out in 2017 by P&C complementing the initial test works performed by in 2009.

16.13.1 Backfill Reticulation and Pumping System

The backfill reticulation and distribution philosophy for the current study is similar to the

proposed route proposed in the previous approach where pipeline is routed overland from the

156

paste plant to a borehole located near the tramway and access road with entry boreholes into

the mine workings.

The following summarizes the proposed pastefill distribution specified by P&C in 2012, with

slight modification to match with the current mine design:

Schedule 120, 4-inch pipes from the surface to the borehole, and down to

approximately elevation 6740 ft. The higher pressure rating and thicker pipe walls

provide additional “safety margin against wear considering the critical location,

difficulty of replacement and potential slack flow. This pipe rating was incorporated

into the design, included in all the haulage levels up to elevation 6700 as per

recommendation. The current mine design also includes this pipe rating for interlevel

boreholes.

Schedule 80, 4-inch pipes are required for the majority of the backfill reticulation

system. This pipe rating is currently used in levels 6700 to 6386.

Schedule 40, 4-inch pipes will be installed throughout the mine below level 6386

instead of cross cuts.

HDPE 4-inch DR9 on XC and AR is used for in-stope piping as well as in the cross

cuts and attack ramps.

P&C “recommends installing a pump with at least 12 to 13 MPa discharge capacity” and

operating the pump at 10 MPa enabling the delivery of high concentration paste to minimize

the potential segregation of paste within the line (P&C, 2012) with the pouring rate of 911ft3/h

of pastefill (P&C, 2011). A standby high pressure plunger type water pump is also connected

to the pastefill line so that the system can be flushed in the event of the pastefill pump breaks

down (P&C 2011).

The backfill plant has been designed to operate for two 10 hours shift, with 4 hours interval for

blasting and maintenance in between shifts. Pouring of pastefill over the period of shift chance

can be achieved by using remote camera if necessary but currently this is not included into the

design criteria (P&C, 2011).

16.13.2 Backfill Material Testing

There are two backfill material testing campaign being performed on the tailings from ICP:

2008 material was tested with Holcim Type I Ordinary Portland Cement.

2017 material was tested with Ash Grove Type I-II, II-IV cement and also blended with

DuraSlag.

The backfill material testing includes material characterization and determination of the

rheology of the tailings. The following sections present conclusions and observations made on

the results from both of these material testing campaigns.

157

16.13.2.1 Strength Testing Results

Table 16.16, summarizes the strength testing results from 2008 where material was tested

based on Holcim Type I Ordinary Portland Cement at 28 curing days. The test matrix in 2017

was more comprehensive with the testing of material with a blend of binder and up to 28 curing

days. See Table 16.17.

Table 16.16

2008 UCS Testing Results

Mix # Slump

(inches)

Est. Mass

Concentration (%m)

Binder Content

(%)

W:C

Ratio

28 Day UCS

(kPa)

1 5 70.7 4 10.4 160

2 6 69.8 2 21.7 82

3 6 69.5 4 11.0 150

4 6 70.0 8 5.4 368

5 7 68.4 4 11. 6 117

6 8 66.5 2 25.2 81

7 8 66.8 4 12.4 125

8 8 67.9 8 5.9 387

9 9.25 65.9 4 12.9 154

10 10.25 65.3 4 13.3 104

11 - 66.0 10 5.2 551

12 - 68.0 4 11.7 160

13 - 68.0 10 4.7 723

14 - 68.0 12 3.9 1010

15 - 70.0 10 4.2 808

Table 16.17

2017 UCS Testing Results

Mix # Tailings

Content (%)

Binder

Content (%) Binder Type

W:C

Ratio

As Cast

%m Solids

UCS (kPa)

7 days 28 days 120 days

1 100 9 50% Type I-II,

50% DuraSlag

4.64 70.5 787 1794 -

2 100 6 50% Type I-II,

50% DuraSlag

5.86 74.0 478 1093 1282

3 100 8 50% Type I-II,

50% DuraSlag

4.33 74.3 820 2050 -

4 100 4 50% Type I-II,

50% DuraSlag

8.59 74.4 293 597 -

5 100 6 Type I-II 5.73 74.4 402 504 494

6 100 6 Type II-V 5.76 74.3 449 547 506

7 100 4 50% Type I-II,

50% DuraSlag

10.03 71.4 219 408 -

8 100 3 50% Type I-II,

50% DuraSlag

13.43 71.3 143 247 -

158

16.13.2.2 Backfill Material Testing Conclusion and Observations.

The following are comparisons, conclusions and observations were made by P&C (2017)

during the material stages:

Material Characterization: During tailings preparation both tailings samples settled

quickly. Water was observed to accumulate on the top of the tailings shortly after

mixing.

Rheology: Yield stress was determined for cemented tailings with mass concentration

ranges from 69.3%m to 72.4%m with a viscometer using a vane spindle. At 71%m

solids and 6% binder, the yield stress is 200 Pa.

Strength:

o Water was observed at the bottom of the cylinders during casting.

o There is very good correlation between the various data sets that suggests that

the early strength is not fully dependent on the different type of binders but

rather on the water to cement ratio.

o Long-term strength gain is only attainable with a 50% Type I-II, 50% DuraSlag

blend and higher solids concentration, based on 120 day strength results.

o The difference in UCS between 28 and 120 days cured was 8% with Type II-V

cement and less than 2% with Type I-II cement (i.e. strength loss).

o A w:c ratio of approximately 11.6 and 16.3 is required to achieve the minimum

strength of 170 kPa after 7 and 28 days, respectively.

16.13.3 Design Criteria

The strength of the backfill required for the mining methods proposed at ICP were estimated

based on stability of free standing backfill formula by Mitchell (1983).

However, assumptions were made for the binder or cement additions for the current study

because of the availability of the latest test results. These assumptions were made based on 28

days curing time binder blends and the information from 2008.

The assumptions made are presented in Table 16.18, where the lead stopes sill mats’ have the

highest cement or binder content to support the extraction of the stope underneath it during the

subsequent mining cycles. High strength caps are prepared with additional binder enabling

transiting of mining equipment while working in the stope and the cores of the stopes have the

lowest cement content to achieve the minimum free standing target strength.

159

Table 16.18

Summary of Estimate Binder Addition

Description Estimated Binder

Addition

Lead Stope

Cut and Fill Sill Mat 10.0%

Core 3.5%

High Strength Cap 10.0%

Longhole Stoping Sill Mat 8.0%

Core 3.50%


Normal stope Cut and Fill

Core 4.0%


Longhole Stoping Core 4.0%


16.14 MINE DEWATERING

The ground water inflow estimate was based on a preliminary estimate documented by Telesto

Solutions, Inc. (Telesto) in 2006 for the development of the Ram deposit. The estimate ranges

from 33 to 66 gpm which Telesto considered to be over-estimated and in the opinion that a

flow rate of 43 gpm is more accurate estimate for the Ram deposit at full excavation.

Based on ICP, the mine could be dry during the development and pre-production stages where

water will be recycled and reuse for the initial development until water wells are established

for the mine. ICP plans to drill 2 wells for potable water. These wells are expected to be in

operation by year of the mine life supplying water for the mining and milling operation.

A series of submersible pumps located on the haulage levels will dewater to the main sump

pump located at the service tunnel. The water will then be pumped to the surface water

treatment by a 200 hp pump. Currently, the mine design includes two dewatering sumps on the

main levels, and one on sublevels located in the north portion of the Ram deposit and at depth.

Dewatering from the mine development, production areas and operating levels will be

accomplished by a series of ten 6 hp submersible pumps each having the capacity to deliver

approximately 150 gpm of water through 4-inch HDPE diameter pipes boreholes during steady

state operation.

160

16.15 COMPRESSED AIR

Compressed air will be supplied 200 hp rotary screw, air cooled air compressor capable of

delivering 1,075 acfm @ 125 psig maximum discharge pressure. Compressed air will be

distributed via 6-inch HDPE lines.

16.16 POWER REQUIREMENTS AND DISTRIBUTION

The mine electrical power demand is approximately 0.9 MW to 1.26 MW with approximately

half of the power demand is from mine dewatering, ventilation and air compressor. Variable

frequency drives installed on the fans, strategic location of the dewatering pumps and

regulating the air compressor to on demand basis can reduce the power consumption.

161

17.0 RECOVERY METHODS

The recovery of all products is completed in a two-step process in separate locations. Initially,

a flotation concentrate containing cobalt, copper and gold is produced at the mine site near

Salmon Idaho, which is transported for processing at the Cobalt Processing Facility (CPF)

located near Blackfoot, Idaho. The valuable products from the CPF include cobalt sulphate,

copper sulphate, gold on activated carbon and magnesium sulphate.

17.1 MINE SITE PROCESS PLANT DESIGN

The process engineering, including flowsheets, process design criteria, mass balance and

equipment selection, for the mine site concentrator was developed during an earlier project

development phase by eCobalt (previously known as Formation) and the then engineering

consultant. Many of the main equipment items have been purchased and are being stored by

eCobalt near to the project mine site.

17.1.1 Process Description

The primary facilities at the mine site include the concentrator, paste backfill plant and the

water treatment plant. A block flow diagram of the mill and concentrator is shown in Figure

17.1.

Figure 17.1

Mine Site Block Flow Diagram

17.1.1.1 Ore Transport

Ore and waste from the mine are hauled to the mine portal using rubber tired underground haul

trucks which are dumped into ore or waste hoppers at the mine portal. Material is transported

from the mine portal to the plant area using an overhead tram where it is discharged onto

UNDERGROUND

MINING

CRUSHING AND

SCREENING

GRINDING FLOTATION CONCENTRATE

THICKENING & FILTERING

TAILS THICKENING & FILTERING

MINE BACKFILL SYSTEM

MINE

PROCESS

WATER TANK

TO REFINERY

CONCENTRATE SHIPPING BY

TRUCK

TAILINGS & WASTE ROCK

STORAGE

162

separate ore and waste stockpiles located near the concentrator building. Each stockpile has a

live capacity of 800 tons, which is sufficient for one day’s operation in the event of a shutdown

for maintenance or repair of the tram. A front end loader will transfer material from the

stockpiles to the hopper of the primary crusher.

17.1.1.2 Ore Crushing Screening and Storage

Primary crushing is via a jaw crusher (22” x 36”) with secondary crushing via a cone crusher

(4 foot). The product from the primary crushing circuit has a nominal 80% passing (P80) size

of 2.5 inches. The secondary crusher operates in closed circuit with a double deck sizing screen.

The product from the secondary crushing circuit has a nominal P80 of ⅜-inch which is stored

in the 400 capacity fine ore storage bin. Ore from the storage bin is transported at a controlled

rate to a grinding ball mill via a conveyor.

Dust collectors are provided in the crusher area and the ore storage area in order to control dust

emissions.

17.1.1.3 Grinding

Ore from the crushing circuit is transported via the mill feed conveyor into the ball mill feed

chute at a nominal rate of 36.2 T/h. Ore, process water and potassium amyl xanthate (PAX)

are fed to the mill, which is operated in closed circuit with two parallel hydrocyclones. The

circuit is designed with a circulating load of 300 % of the feed. The cyclone overflow product

with a target P80 of 70 µm gravitates to the flotation conditioner.

17.1.1.4 Flotation

The flotation circuit consists of a conditioning tank, rougher flotation cells and cleaner

flotation. Frother and additional PAX is added to the slurry in the agitated conditioning tank.

The slurry from the conditioner feeds the rougher flotation bank which comprises two banks

of four flotation cells. The rougher flotation tailings are the final tailings from the concentrator.

The concentrate from the rougher flotation circuit is collected and pumped to the distribution

box in the cleaner flotation circuit. The distribution box feeds the cleaner flotation bank, which

comprises two banks of six cleaner flotation cells. The tailings from the cleaner flotation circuit

are recycled to feed the rougher flotation circuit. If the metal concentrations in the cleaner

tailings are sufficiently low, they can be pumped directly to the tailings thickener.

The cleaner concentrate is the final product of the flotation circuit. The mass of the concentrate

is typically reduced to about 9% of the mass of the feed to the circuit. The cleaner concentrate

is pumped to the concentrate thickener.

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17.1.1.5 Concentrate Dewatering

The final flotation concentrate feeds the concentrate stock tank for storage prior to dewatering

in the concentrate filter press. The concentrate is trucked to the CPF refinery for final

processing.

17.1.1.6 Tailings Dewatering

The tailings from the flotation circuit are pumped to the tailings thickener. Flocculant is added

to the thickener in order to enhance the settling of the solids. The water recovered in the

thickener overflow is stored in the process water tank for reuse in the circuit. The thickener

underflow slurry is pumped to the mine backfill system.

The thickened tailings are filtered using a vacuum disc filter and the filter cake is either

conveyed to a storage area from which it is loaded into trucks to be transported to the tailings

storage facility or it feeds the backfill mixer where it is blended with cement and water then

pumped underground to be used as paste backfill.

17.1.1.7 Concentrator Reagent Feed Systems

Reagent mix systems are provided for flocculant, PAX, frother and lime.

Two flocculant mix systems are included. One of the flocculant mix systems is used to prepare

flocculant for use in the tailings thickener; the second one is used to prepare flocculant for use

in the water treatment plant.

The frother system includes a frother stock tank and metering pumps.

Sodium bisulphite, antiscalent and hydrochloric acid are pumped directly from drums to the

water treatment system.

17.2 COBALT PROCESSING FACILITY (CPF)

17.2.1 Process Description

The CPF design is a hydrometallurgical processing facility located near Blackfoot, Idaho. It is

a sophisticated process that uses a complex series of unit operations to produce a number of

products. The processes include pressure leaching in autoclaves, solvent extraction,

crystallization, precipitation, thickening and filtration. Flotation concentrate from the mine site

is shipped to the CPF for processing into saleable products.

The residue produced in the refinery is shipped via rail using side dump cars to an offsite waste

facility. Waste effluent from the site is disposed of in the municipal sewer line for treatment in

Blackfoot, Idaho.

Block flow diagrams of the CPF facility are outlined in Figure 17.2 and Figure 17.3

164

Figure 17.2

CPF Refinery Block Flow Diagram – Concentrate Feed Circuit

165

Figure 17.3

CPF Refinery Block Flow Diagram – Cobalt Simplified Circuit

166

17.2.2 Acidulation

The cobalt-copper-gold flotation concentrate is reground then repulped with Cu solvent

extraction (SX) raffinate and dilute sulphuric acid in order to be conditioned in a series of

acidulation tanks prior to being fed to the leach circuit.

17.2.3 Pressure Oxidation Acid Leaching

The pressure oxidation leaching is the heart of the whole hydrometallurgical process. The

cobalt-copper-gold concentrate is leached with acid in a continuous autoclave circuit at high

temperature (155ºC) and pressure (5 bar) via oxidation of sulphide host minerals with oxygen.

At this condition, more than 99% of the valuable metals such as copper, cobalt and some of

the gold are leached from the concentrate, while most of the iron and arsenic are hydrolysed

as scorodite and hydronium jarosite in a stable leach residue.

The feed to the autoclave is comprised of two main streams:

Repulped concentrate, which consists of the concentrate slurry from the acidulation

circuit and is mixed with other recycle streams such as the End Flash primary filtrate

and loaded Zn scrub solution.

Recycle streams, which are composed of Cu/Fe removal thickener underflow,

autoclave discharge thickener U/F and repulped traced metal filter cake for further

cobalt and copper recovery.

In the autoclave, steam is added during the start-up to initiate heating. Oxygen is injected to

further oxidise the surface of chalcopyrite and cobaltite for acid leaching. The pressure

oxidative leaching of cobaltite, chalcopyrite and other minor sulphide minerals such as pyrite,

millerite and sphalerite (contained in the repulped concentrate) are highly exothermic as per

following reactions:

4CoAsS(s) + 13O2(g) + 6H2O(a) → 4CoSO4(a) + 4H3AsO4(a)

2HNO3(a) + CuFeS2(s) + 3O2(g) → H2O(a) + CuSO4 (a) + FeSO4(a) + NO(g) + NO2(g)

CuFeS2(s) + 2Fe(SO4)3(a) → CuSO4(a) + 5FeSO4(a) + S(s)

4FeS2(s) + 15O2(g) + H2O(a) → 2Fe(SO4)3(a) + 2H2SO4(a)

In a continuous operation, the combined heat generated from these reactions is enough to

sustain the autoclave reaction temperature at 155ºC without the need for a live steam make-up.

It is therefore critical to control the density of the of the repulped concentrate feed to maintain

the heat requirement inside the autoclave. The operating costs have accounted for five cold

starts in the first year followed by three in subsequent years.

Sulphuric acid is added at typical rate of 200 kg/t to leach the oxide component of the

concentrate (talc, clinochlore, biotite, goethite and calcite) and hydroxide precipitates (copper

and cobalt hydroxide) from the recycle streams.

167

At high temperature and pressure, ferric iron in the solution subsequently hydrolyze with

arsenic acid as stable scorodite, hematite and hydronium jarosite (with localised acid gradient)

according to the following reactions:

Fe2(SO4)3(a) + H2O(a) → Fe2O3(s) + 3H2SO4 (a)

Fe2(SO4)3(a) + 2H3AsO4(a) + 4H2O(a) → 2FeAsO4.2H2O(s) + 3H2SO4(a)

3Fe2(SO4)3(a) + 14H2O(a) → 2(H3O)Fe3(OH)6(SO4)2 (s) + 5H2SO4 (a)

These three mechanisms re-generate sulphuric acid, thus contributing to a large reduction in

acid consumption rendering the process economically feasible. Scorodite, hematite and jarosite

report to the leach residue and have good filtering characteristics.

Nitric acid is employed as the catalyst in the autoclave at concentrations of approximately 2

g/L. In the acid system at the temperature of 155ºC some of the nitrate reacts with sulphides

and releases nitrogen oxides in the vapor phase.

MS(s) + 4HNO3(a) → MSO4(a) + 2H2O(a) + 2NO(g) + 2NO2(g)

The nitrogen oxides in the vapor phases are further oxidised and inducted into the slurry where

they react with concentrate providing an electron sink in the extraction process.

NO(g) + ½O2(g) → NO2(g)

NO2(g) → NO2(a)

4NO2(a) + S2-(a) → SO42-(a) + 4NO(g)

The cyclical process of oxidation and reduction continues repetitively with the nitrogen oxides

being the catalyst. Most of the nitrate is lost in the leachates where it is ultimately bled from

the circuit to the copper iron removal circuit. A tight control on water addition to the leach

circuit is essentially to minimise the bleed and hence catalyst loss to the cobalt circuit.

Majority of the NOx are recovered in the NOx recovery process, where the hot gas is contacted

with the cooled dilute HNO3. More than 50% of the NOx and steam are stripped from the vent

gas and recovered as HNO3. A portion of the recovered HNO3 is bled back to the autoclave

minimising the addition of fresh HNO3.

The temperature in the autoclave is regulated by a patented flash-cool system where:

About 75% of the autoclave discharge is drawn from the lead autoclave compartment

and flashed to atmospheric conditions at 102ºC. The discharge slurry is subsequently

cooled to about and directed to the subsequent acid neutralisation section

The remaining flow (about 25%) is discharged from the last compartment of the

autoclave and flashed to atmospheric conditions at 102ºC. The discharge slurry is

subsequently cooled to around 60ºC and forwarded to the Leach Filtration section.

Leach Neutralisation/Thickening.

168

The purpose of this circuit is to:

Neutralise the free acid in the discharge slurry prior to reintroduction to the autoclave,

allowing for a more manageable acid control in the autoclave.

Produce a cooled and clear pregnant leach solution (PLS) for feed to the Cu-SX circuit.

The discharge from the front end compartment is forwarded to the Leach Thickening section,

where free acid is neutralised with MgO in a series of acid neutralisation tanks. Raw water and

Cu crystallizer condensates are also added in the first tank to dilute the cobalt concentration

and minimise any localised precipitation of cobalt from the solution. The resulting slurry is

then forwarded to a conventional thickener where flocculant is added to aid the settling. The

slurry is thickened to approximately 50% w/w solids underflow and is recycled back to the

autoclave for further recovery of copper and cobalt. The thickener overflow is sent to a heat

exchanger to lower the temperature from 60ºC to 45ºC prior to further clarification in the Cu

PLS clarifier. The clear clarifier PLS overflow is then forwarded to Cu-SX while the clarifier

underflow is recycled back to the lead acid neutralisation tank for recirculation.

17.2.4 Cu-SX and Crystallization

The objective of this circuit is to refine the Cu-PLS and remove the impurities before

recovering the copper as copper sulphate pentahydrate product in the crystallizer.

The cooled and clear Cu PLS is fed to the Cu-SX section which is composed of:

Three extraction stages.

Two strip stages.

The PLS is pumped to the first extraction stage where the majority of the copper is extracted

using 20% organic added at an O:A ratio of 1:1. The loaded organic exits the first extraction

stage and stripped with 180 g/L H2SO4 in the subsequent stripping stages. The strip solution is

a combination of mother liquor bleed from the copper crystallizer and 30% sulphuric acid

make-up. It is introduced in the last stripping stage.

The stripped organic is sent back to final extraction stage, where Cu-raffinate is also drawn.

The recirculation of the Cu raffinate is split into several streams. The majority of the flow is

sent to Acidulation stage to repulp the reground concentrate, a bleed portion is sent to the Cu-

Fe Removal Circuit.

The pregnant electrolyte from the first strip stage is forwarded to the evaporative crystallizer,

where copper pentahydrate sulphate crystals are produced for sale. The mother liquor bleed

from the crystallizer is sent back to the Cu-SX circuit as strip solution. The evaporated steam

is collected as condensates and returned to the Leach Neutralisation/Thickening circuit.

Crud formation at the organic/aqueous interfaces in the settler can inhibit effective phase

separation and potentially contribute to organic loss. Crud is removed from the settler and

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pumped by a portable air operated diaphragm pump, as required, to the crud recovery tank, to

allow crud to be accumulated and be treated on a batch basis.

17.2.5 Leach Filtration

The objective of this circuit is to provide a solid-liquid separation of the end discharge of the

autoclave, where the solid residue is sent for gold recovery.

The end discharge of the autoclave is cooled in a coil-fitted tank from 60ºC to 30ºC using

cooling water. The cooled slurry is mixed with flocculant to aid in the filtration and

subsequently fed to a belt filter with three stages of displacement washing.

The washed cake of about 83% w/w solids is forwarded to a repulp tank for sulphur flotation

while the primary filtrate is sent back to the autoclave for recirculation. The second and third

wash filtrates feed clarifier, the overflow from which is recycled back as wash water while the

underflow is recycled to the front end belt filter feed tank.

17.2.6 Sulphur Flotation

This circuit recovers, by flotation, the sulphur pellets that tend to form in the autoclave (and

that can entrain unleached or partially leach copper and cobalt sulphides), producing a sulphur

concentrate and a tails which contains most of the undissolved gold for further recovery as a

byproduct.

The leach filtration cake is repulped to about 40% w/w solids with raw water. MgO is also

added to neutralise any residual acid. The resulting slurry is subsequently forwarded to a 3-

bank flotation cell, where PAX and frother are added to float the sulphur and unleached /

partially leached cobaltite and chalcopyrite. The recovered concentrate is returned to the

concentrate regrind circuit prior to acidulation while the tailings are sent to a series of three

oxidation tanks, where blower air is injected to oxidise the surface of the material in preparation

for cyanide leaching. The oxidised slurry is then forwarded to a disk filter for dewatering.

Majority of the filtrate is sent to poor water recovery for recirculation in various areas while a

bleed is sent back to the front-end repulping tank. The filter cake is discharged to a repulping

tank, where fresh water is added to repulp the cake to about 50% w/w solids ahead of

cyanidation.

17.2.7 Gold Recovery

This circuit recovers the gold values from the sulphur flotation tails via cyanidation and

carbon-in pulp (CIP) process.

The repulped tails from the sulphur flotation circuit feeds a series of four agitated tanks for

gold leaching. In the lead tank, lime is added to neutralise any residual acid and to ensure that

the tank discharge pH is around 10 to 10.5, which the requirement for the cyanidation. Sodium

cyanide (NaCN) is added to the distribution box at a rate dictated by the cyanide analyser which

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determines free and WAD cyanide levels in the leach. Hydrogen cyanide detectors are also

provided to ensure operator safety by alarming at levels lower than could cause harm.

Gold is solubilised according to the following chemical reaction:

4Au(s) + 8NaCN + 2H2O + O2(g) → 4NaAu(CN)2 + 4NaOH

The gold leach slurry gravitates from the final leach tank to a series of 4 adsorption (CIP) tanks.

Fresh activated carbon is loaded periodically to the last adsorption tank and carbon is

transferred forward to the next upstream tank daily by recessed impeller pumps. During its

passage through the adsorption circuit carbon becomes progressively loaded with gold, silver

and copper cyanide complexes which are adsorbed onto the carbon.

Loaded carbon is pumped from the first adsorption stage to the loaded carbon screen. Spray

bars are provided to wash the loaded carbon before it is discharged for packaging and

transportation for further treatment to recover gold and silver. Screen undersize is recycled

back to the first adsorption tank.

17.2.8 Secondary Belt Filter/Cyanide Destruction

The cyanide destruction circuit is required for the operation of the process plant to ensure that

effluent discharged from the process plant meets environmental requirements.

The screen undersize from the CIP circuit is pumped to an in-line mixer, where flocculant is

added to aid the dewatering in the subsequent filtration stage. The pre-flocculated slurry is then

filtered and washed. The wash filtrates are collected and sent to a clarifier. The clarifier clear

O/F is returned to the belt filter as wash water while the U/F is recycled to theSulphur Flotation

circuit or sent directly to the tails collection tank together with the filter cake.

The primary filtrate reports to the cyanide destruction tank where air and sodium

metabisulphite (SMBS) is added in sufficient quantity to destroy all of the cyanide complexes.

The discharge from the cyanide destruction circuit is then forwarded to the tails collection tank

for disposal.

17.2.9 Cu-Fe Removal

This circuit removes the majority of the impurities (mostly copper and iron) in the cobalt rich

PLS prior to downstream refining.

This circuit treats the heated Cu-SX PLS from the autoclave along with other recycle streams

that contain recoverable cobalt values (Co precipitation cyclone U/F, Co Scrub VSEP filtrate)

with MgO to around pH 4.5 in a series of 4 agitated tanks. At this pH more than 95% of the

copper and ferric iron is removed as stable hydroxide precipitate.

The resulting slurry is pumped to the Cu/Fe removal thickener for solid-liquid separation.

About 70% of the thickener U/F is recycled back to the lead Cu/Fe Removal Tank for seeding.

The balance is sent to the autoclave for further recovery of copper and co-precipitated cobalt

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values. The clear overflow, which is virtually free of copper and iron impurities, is pumped to

the downstream cobalt refining circuit.

17.2.10 Cobalt Precipitation

This circuit selectively precipitates cobalt from the PLS by neutralisation and re-dissolving

with spent acid from the cobalt crystallizer for further refining in the Co-SX circuit.

The impurity free and cobalt rich PLS is received in the series of five agitated cobalt precipitate

tanks along with other recycle streams such as Co-Raff RO bleed, Co-Scrub RO concentrate

and poor water bleed. MgO is added to the lead tanks to increase the pH of the solution to 8.5.

At this pH around 90% of the Co is precipitated from the solution (as per below reaction) along

with other minute impurities present in the solution such as iron, copper, nickel and manganese.

CoSO4(a) + Mg(OH)2(a) → Co(OH)2(s) + MgSO4 (a)

The precipitation of cobalt from the solution also yields MgSO4 in the barren liquor and a bleed

of this stream is treated to minimise the build-up of MgSO4 in the system.

The discharge slurry from the last cobalt precipitate tank is pumped to the cobalt precipitation

cyclone to separate the coarse particles, which are mostly composed of SiO2, hydrated

magnesium sulphate salts and unreacted MgO, from the fine cobalt hydroxide precipitate. The

cyclone U/f is recycled back to the Cu/Fe Removal Circuit, while a bleed portion is sent to the

MgSO4 crystallizer as part of the Mg control in the overall system.

The cyclone O/F is cooled then filtered. Entrained solution in the filter cake is washed with

process water for displacement. The collected filtrate, which is mostly saturated with MgSO4,

is sent to the Trace Metals Precipitation circuit prior to MgSO4 recovery in the crystallization

step.

The filter cake is discharged to the cobalt resolution tank where, RO water, Co-SX Raff RO

permeate, Co crystallizer condensate and blow-outs are added to repulp the precipitate. The

acid contained in some of these streams redissolves the cobalt hydroxide into the solution as

sulphate. Make-up acid is added to supplement the requirement for re-dissolution of the

precipitate.

17.2.11 Cobalt SX

The objective of this circuit is for further impurity removal and concentration of cobalt PLS

ahead of crystallization.

The discharge slurry from the Co Re-solution circuit is pumped to the Co-SX circuit which is

composed of:

Single extraction stage.

Four scrubbing stages.

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Single stripping stage.

The Co PLS feed is received in the cobalt extraction mixer settler where it is mixed with

organic (20 vol% Cyanex 272) at an O:A ratio of 1:1. Co and Zn are preferentially loaded onto

the organic over Ni, Co and Mg, which remain in the raffinate solution. More than 70% of the

cobalt in solution is extracted to the organic. During the extraction, H2SO4 is generated in the

aqueous phase according to the following reaction:

CoSO4(a) + 2oRcoH(o) → oRco2Co(o) + H2SO4 (a)

The product acid helps in the complete re-dissolution of the remaining Co(OH)2 in the feed

PLS. pH control is regulated by the addition of MgO.

The loaded organic is held in the loaded organic surge tank to reduce the carry-over of

entrained aqueous solution and is pumped from there to the last scrubbing stage. The scrubbed

organic is pumped to the cobalt strip tank where it is contacted with dilute sulphuric acid and

Co-crystallizer purge to strip Co from the organic. The loaded strip solution then gravitates

from the strip mixer-settler to the loaded strip solution tank. Entrained organic in the loaded

strip solution is removed ahead of crystallization by a series of multimedia filters and a

coalescer.

The raffinate from the cobalt extraction mixer-settler gravitates to the raffinate after settler and

pumped to carbon columns in order to recover the soluble gold (from pressure oxidation acid

leaching) via adsorption.

The cobalt raffinate is fed to the Co raffinate RO VSEP Package where the majority of the

trace heavy metals, such as Ni, Co, Fe, Cu and Mg, and some acid are removed in the reverse

osmosis process. The concentrate is discharged periodically to the metals loaded solution

return tank and pumped to nickel cementation column for nickel removal. A bleed of the

concentrate is recycled to Cu/Fe Precipitation circuit in order to recover the cobalt and build-

up the nickel concentration within the system. The Co-raffinate RO water is sent to the Co-

precipitate repulp tank.

17.2.12 Crud Treatment

Entrained organic from aqueous streams emanating from the SX circuits is recovered by the

crud treatment facility, which comprises an agitated crud treatment tank where clay is added,

and two stages of centrifuging. . Centrifuged organic is returned to the circuit. Solid matter is

deposited into containers for removal. Recovered aqueous is pumped back intermittently to the

Zn scrub pump tank.

17.2.13 Cobalt Sulphate Crystallization

The objective of this circuit is to recover the cobalt from the solution as cobalt sulphate

heptahydrate crystals via multi-effect evaporative crystallization.

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Following organic removal, the pure cobalt sulphate feed solution is pumped to the crystallizer

along with the seed. The flashed steam from the autoclave is injected in the heater in order to

transfer the heat to the cobalt feed solution. As water evaporates the solution is concentrated

and becomes supersaturated, promoting the formation of nuclei for crystallization. The vapour

that is generated as a result of the evaporation is condensed in a water-cooled condenser, with

the clean condensate recycled as clean water to various areas of the plant.

The supersaturated cobalt solution is forwarded to the cobalt crystallizer where it contacts more

cobalt crystals through agitation. The seeding promotes additional site for nucleation. The

growth of the crystals is dependent on the extent of super saturation and recirculation control.

When the target particle size of is reached, the crystals are withdrawn from the crystallizer

thickened and further dewatered in the cobalt crystals centrifuge, where the crystal discharge

is split into portions:

Seed which is recycled to the feed tank.

Product stream which is sent to the last stage of drying in the tray crystal dryer.

The final moisture of the crystal product is controlled by drying with hot air. The dried cobalt

sulphate heptahydrate crystals is then discharged to a conveyor for bagging.

The mother liquor from the crystallizer is periodically bled to minimise the build-up of

impurities in the crystallizer system.

17.2.14 Trace Metal Precipitation

The Trace Metal Precipitation circuit removes trace impurities in the MgSO4 bleed streams

prior to recovery of MgSO4 via crystallization.

The cobalt cyclone underflow bleed and Co-ppt filtrate are fed to a series of four trace metal

precipitation tanks, where MgO is added to precipitate the trace metals such as Cu, Co, Fe, Zn

and Ni as metal hydroxides. The discharge slurry is then pumped to the trace metals lamella

clarifier for solid-liquid separation. The clarifier underflow is recirculated back to the autoclave

for recovery of target metals and the clarifier O/F is sent to a polishing filter for further removal

of fine solid precipitates. Diatomaceous earth is used as a pre-coat to aid the filtration of the

fine solids.

The filter cake from the polishing filter is periodically removed by backwashing and

recirculated back to the autoclave.

17.2.15 Magnesium Sulphate Crystallization

Recovery of solid magnesium sulphate (Mg2SO4) is achieved by heating of the feed brine to

the boiling point, evaporating water and as a result, increasing the Mg2SO4 concentration in

the remaining brine above its saturation point. The super saturation of the dissolved solids

causes crystals of Mg2SO4 to precipitate out of solution and to grow in size by attaching

themselves to existing previously precipitated crystals in the brine. Removing a portion of the

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brine and solids from the evaporator vessel allows for the solids to be separated from the brine,

with the brine being returned to the evaporator vessel along with fresh feed brine. The water

vapor driven off from the brine is condensed and cooled, and the resulting relatively pure water

can be reused for other purposes.

At equilibrium conditions, the volume of fresh feed introduced is equal to the water vapor

boiled off from the boiling surface, plus the volume of solids and brine sent to the dryer. The

solids tonnage removed from the system at equilibrium is equal to the tonnage of Mg2SO4

dissolved in the feed brine that is introduced into the evaporator system. An intermittent purge

is removed from the centrate brine to control the concentration of contaminant ions below a

critical level and is made up with a corresponding intermittent increase of feed brine as

required.

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18.0 PROJECT INFRASTRUCTURE

18.1 MINE AND MILL SITE

The mine site is located in an area generally characterized by deep, relatively narrow valleys

with steep side slopes and rugged mountains. The mine portal is located in one of these valleys.

Adjacent to this valley is a flat-topped mountain which is referred to as the Big Flat. The

processing facilities and the tailings and waste rock storage facilities (TWSF) are located on

the Big Flat. The Big Flat is a gently sloping area at an elevation of approximately 8,000 ft,

which is approximately 1,000 ft. higher than the mine portal. Material is moved from the mine

portal to the Big Flat using an aerial tramway system.

Facilities located on the Big Flat include:

The tram head structure.

Coarse ore and waste storage.

Crushing and screening facilities located in a crushing building.

Processing facilities located in the concentrator/process building.

Backfill preparation plant located in the process building.

Wastewater treatment plant located in the process building.

Tailings and waste rock storage facility.

Mine office building and change facilities.

Fuel and lube storage and dispensing facilities.

Mine equipment maintenance facility (truck shop), planned for construction in Year 2.

Topsoil stockpile.

Electrical substation.

18.1.1 Site Layout

The overall site layout is shown in Figure 18.1.

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Figure 18.1

Site Facility Map

18.1.2 Work Completed to Date

The concentrator and mine site facilities and infrastructure were previously under construction

and as such, preliminary work has been completed, including:

Completion of the access road from highway 93 to the mine site.

Security/Gate House, which has been purchased and is installed along the access road

to the mine site.

Site preparation including stripping and grading.

Earthworks for the first cell of the Tailings Waste Storage Facility (TWSF) was nearly

completed during the 2011 construction phase. The installed portion of the liner has

been damaged since construction ended and will need to be replaced or repaired.

Some building footings have been installed for the crusher building and the

concentrator building.

The administration building has been purchased and installed at site. No utilities have

been installed to the building.

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The incoming power supply line was completed during the last phase of construction.

Tie-ins to the supply line and the site distribution system will be installed during the

next phase of construction.

The road to the portal location and portal bench has been completed. A Hilfiker

retaining wall will be constructed during final construction prior to mine development.

A small warehouse and yard south of Salmon, Idaho, has been purchased. The Salmon

depot is currently used for storage of the purchased equipment. In future, this site will

be used as a mustering point for construction and operations employees who will be

bussed to site. It will also serve as temporary storage of concentrate prior to shipment

to the hydrometallurgical facility (CPF) and incoming shipments bound for the mine

site.

18.1.3 Mine Site Access Roads

Vehicle access to the ICP mine site is via a series of well maintained, public access roads which

are open year round. Access to the road is approximately 6 miles south of Salmon, Idaho, on

Highway 93 which also services the Blackbird mine (currently not in operation). The total

distance from the Salmon depot to the ICP is approximately 48 miles.

Employees are expected to live in the Salmon area and will be transported to the project site

by buses or vans from the Salmon depot. Access to the site from Salmon is shown in Figure

18.2. This route will also be used for transportation of concentrate, equipment, reagents, and

other freight.

Figure 18.2

General Mine Site Area Map

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18.1.4 Buildings

The process facility and ancillary buildings include the following:

Crusher Building – has been purchased and is stored at the Salmon Depot.

Concentrator Building – has been purchased and is stored at the Salmon Depot.

Control Room (enclosure within the Concentrator Building).

Sample Prep Room (enclosure within the Concentrator Building).

Administration Building - has been purchased and is installed at the mine site.

Dry/Change House.

Security gate/house – has been purchased and is installed along the access road to the

mine site.

Warehousing of spare parts, reagents and consumables will be in the crushing and screening

building and in the concentrator building.

No separate maintenance shop is provided. Routine maintenance of the surface equipment is

performed in the crushing and screening building. Breakdown maintenance is performed

offsite. The mining contractor is responsible for maintenance of its equipment and the

maintenance functions are carried out inside the mine or on the portal bench.

18.1.5 Electrical Power Supply and Distribution

The project mine site is supplied by a 69-kV power line provided by the Idaho Power

Company. A power supply line to the adjacent Blackbird Mine Site, which currently feeds only

the Blackbird water treatment plant, already exists.

18.1.5.1 Mine Site Incoming Power Supply Line

The 69-kV incoming power line originates in Salmon and services the Blackbird Mine Site.

The power line to supply the ICP will be from a new tap on the existing line to the new

substation located near the concentrator.

18.1.5.2 Site Power Distribution

Transformers located at the concentrator substation reduce the voltage to 4.16 kV for further

power distribution. Power for the tram drive system and the concentrator ball mill are operated

at 4.16 kV. All other loads operate at 480 V, with the exception of lighting, instrumentation

and other small loads.

An overhead power line at 4.16 kV runs parallel to the tram from the concentrator substation

to the mine portal area. Transformers located at the portal reduce the voltage to 480 V for

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distribution within the tram loading facility. Power distribution within the mine will be at 4.16

kV.

Power for the water management pond pumping system and the surface mounted mine vent

fans are provided by 4.16 kV overhead power lines that run from the concentrator to the point

of consumption with transformer used to reduce the voltage to 480 V for supply to the

equipment.

18.1.6 Surface Facilities Fire Protection

Neither local building codes nor the FCC insurance carrier require a permanent fire protection

system. Fire protection for the surface facilities is provided using a combination of hand-held

(20-pound) fire extinguishers and wheel-mounted (120-pound) fire extinguishers.

All vehicles and mobile equipment are equipped with fire extinguishers.

Forest Service regulations require that fire suppression equipment be available during the fire

season. A water truck equipped with pump, hoses and nozzles is sufficient to meet this

requirement. Shovels and other tools for suppressing small fires will also be stocked and made

readily available for employees.

18.1.7 Mine and Concentrator Communications

Administrative functions for the mine and concentrator are performed primarily from the

existing FCC office in Salmon Idaho. This includes senior management, human resources,

accounting, payroll, accounts payable and procurement. The office is currently connected to

all required data and voice communications networks.

Surface communication at the mine site is primarily by cellular phones. Communication to the

Salmon administrative site is via cell service, land line or network data system.

Communications within the underground mine use radios and a leaky feeder system. This

system allows radio communication from all locations within the mine.

.

18.1.8 Water Supply, Treatment and Discharge

The primary demand for water is for processing at approximately 960 gallons per ton processed

(768,000 gallons per day) at the nominal production of 800 tons per day. Except for water lost

to the concentrate and the tailings, the effluent from the milling operation reports to a water

management pond. This water mixes with mine water and other waters reporting to the pond

and is recycled back to the mill.

The primary source of water for the operation comes from dewatering the Ram deposit. The

average flow during the life of the mine is estimated to be approximately 45 gpm from the Ram

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Mine. Additional water for the operation comes from the collection of runoff from the TWSF

and storm water. These flows report to the water management pond.

18.1.8.1 Water Management Pond

The water management pond is sized to contain all process and mine drainage waters, and all

drainage waters from the TWSF. Additional capacity is required to contain runoff from the

500-year storm event. The pond has a total required capacity of 10 million gallons. The pipeline

from the pond to the mill is double-contained and complete with leak detection at all low points

and at pipe-to-pipe connections.

The liner design for the water management pond consists of a double synthetic liner system

with leak detection and leak collection.

18.1.8.2 Mine Dewatering

Discharge from the mine dewatering system is delivered to a holding sump at the portal. The

sump is sized to contain the entire backflow from draining the pipeline from the Ram portal to

the mill on the Big Flat.

Pumping from the portal to the mill is accomplished via a winterized steel pipe with secondary

containment. During an emergency shutdown or production curtailment, the mine pumps

continue to operate in order to maintain the entire water balance and control system.

18.1.8.3 Water Treatment

The objective of the water treatment facility is to produce the highest quality discharge stream

that is reasonably achievable. A secondary objective is to operate the system with as close to a

zero liquid-waste discharge condition as possible. Water management is based on operating a

water treatment plant and releasing water in accordance with a National Pollutant Discharge

Elimination System (NPDES) permit in conjunction with temporary storage in the water

management pond.

The water treatment plant will incorporate the following treatment methodologies:

Metals removal by oxidation, precipitation and filtration/settling (co-precipitation).

Metals polishing removal by cation ion exchange (IX).

Nitrogen removal biological denitrification (biodenitrification) via moving bead

bioreactor (MBBR).

Co-precipitation.

The design of the water treatment system is based on the water flow diagram outlined in Figure

18.3.

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Figure 18.3

Water Treatment Block Flow Diagram

Except during periods of very high inflow, the water treatment plant treats incoming water on

an as-received basis. During periods of high inflow, water will accumulate in the water

management pond for treatment during lower inflow periods. Mine water quality is predicted

by the Dynamic Systems Model (DSM) to contain elevated concentrations of nitrate, sulphate,

and metals (aluminum, cobalt, copper, iron, manganese, zinc).

Final effluent from the treatment plant will be discharged into Big Deer Creek immediately

downstream of WQ-24a in accordance with the NPDES permit which is to be issued by EPA.

FCC has stated in its NPDES permit application to EPA that the treated water will meet all in-

stream standards and will not require a mixing zone.

18.1.9 Tailings and Waste Rock Storage (TWSF)

A single surface disposal facility is used to store both the tailings from the concentrator and

the waste rock material. This facility serves to minimize the area of disturbance by sharing

containment and drainage collection facilities while providing storage for these materials.

The TWSF is located east of and down slope from the mill on the Big Flat. This location was

chosen as the best site for the facility in the project area because of its relatively flat

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topography, avoidance of jurisdictional wetlands, soil characteristics, and distance from active

drainages and streams.

Specific design elements of the TWSF are:

Storage of 800,000 tons of waste rock and 960,000 tons of tailing, based on production

estimates for the feasibility study, with separation of tailing and waste rock to the extent

practicable.

2.5H:1V inter-bench side slopes.

Geomembrane liner system with drainage collection.

Diversion of runoff around operating areas of the facility.

Collection of and conveyance of runoff and seepage from tailing and waste rock to

synthetically lined water management ponds.

Collection and conveyance of shallow groundwater flow to wetland mitigation ponds.

Toe berm to provide geotechnical stability and storm water control.

Occupies an area of about 36 acres.

A general layout of the TWSF and water management ponds is shown Figure 18.4.

Figure 18.4

TWSF Plan View

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The liner system for the TWSF consists of a 40-mil PVC synthetic liner placed over a

geosynthetic clay liner (GCL). An 80-mil HDPE rubsheet will be placed over the primary PVC

liner on the perimeter of the TWSF to protect the primary liner from UV and physical damage.

A drainage collection system will be constructed over the PVC liner to collect water infiltrating

through the tailing and waste rock. The drainage collection system consists of a series of

perforated pipes connected to a header pipe that conveys flow to the water management pond.

The drainage collection system will be constructed within a protective sand layer, which also

acts to protect the PVC liner from damage during tailing and waste rock placement.

18.1.9.1 Topsoil Stockpile

Growth media salvaged during the construction of the project site roads and other facilities

will be stockpiled in one area. The area is located adjacent to the TWSF. The total amount of

topsoil salvaged is estimated to be 279,000 yd3. Approximately 7.2 acres are required for the

growth media stockpile area. Precipitation runoff is diverted around the area by perimeter

ditches. As topsoil materials are placed, this area will be seeded to stabilize the stockpile.

18.1.10 Explosives Storage and Transport

Explosive will be delivered by the powder manufacturer/distributor to the designated surface

explosive storage facility at the ICP.

Explosive requirements for the underground operation will then be transported by designated

underground explosive vehicles certified to transport explosive to the underground storage

facilities.

From the underground storage facilities, explosive will then be distributed to the working areas

with the similar designated vehicles.

18.2 CPF INFRASTRUCTURE

The CPF will accept the flotation concentrate from the mine for final processing. All products

will be shipped from the CPF site via rail or truck. Solid process residue will be transported

via rail to an offsite disposal facility.

18.2.1 CPF Site Access

Access to the site is via Pioneer road which is appropriately rated for the anticipated truck

traffic to and from site. The site has access to highways 15 through Blackfoot Idaho for delivery

of concentrate and export of products to any location in North America. The site is adjacent

to a Union Pacific (UP) rail line with access to Blackfoot rail yards which provides a

connection point to the primary rail lines in Idaho.

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18.2.2 Process Plant Layout

The CPF location is a greenfield site located approximately 3.5 miles north west of Blackfoot

Idaho on Pioneer Road in a light industrial area. The 15 acre site is currently used as

agricultural land and is relatively flat as shown in Figure 18.5.

Figure 18.5

CPF Site Location

Facilities on site include:

Refinery building (including the primary process building and solvent extraction

building).

Crystallizer pad.

Administration building.

Rail spur lines.

Truck scale.

18.2.3 Buildings

The overall refinery building dimensions are 450 ft. x 169 ft. x 50 ft. with equipment located

both at group level and in elevated platforms. With the exception of the crystallizer, all process

equipment is located within this building. The crystallizers are separated from the refinery

building on a 128 ft. x 56 ft. concrete pad.

18.2.3.1 Primary Process Building

The primary process building reuses a pre-engineered building which was previously

purchased by FCC. The building will house the following areas:

CPF

Site

Blackfoot, Idaho

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Cobalt concentrate feed preparation.

Acid neutralization.

Reagents handling.

Trace metals precipitation and recovery.

Autoclave circuit.

Cobalt precipitation and re-dissolution.

Acid neutralization.

Electrical and control rooms.

18.2.3.2 Solvent Extraction Building

The solvent extraction building is a building extension to the process building which houses:

Shops and warehouse.

Primary filtration, gold recovery and residue handing.

Cobalt solvent extraction.

Copper solvent extraction.

A new pre-engineered building is used for the extension and is separated from the main process

building by a concrete block wall. All building materials in this area are resistant to an acidic

environment including epoxy coated steel, acid resistant concrete and stainless steel

mechanical and electrical equipment. A concrete block wall separates the solvent extraction

building form the primary building due to the potential for explosion in the solvent extraction

areas. Explosion proof motors and heaters will be used in the solvent extraction area.

18.2.4 Crystallizer Pads

The crystallizer pad area contains all equipment required for production of:

Copper sulphate

Cobalt sulphate.

Magnesium sulphate.

All crystal products are bagged at the crystallizer pad for loading onto the adjacent rail spur

via forklift. No buildings are installed at the pad but service, inspection and equipment

platforms are required to house the dryers, pumps, thickeners, filters and bagging equipment.

The platforms are constructed with painted structural steel with steel grating on the platform

levels.

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18.2.5 Administration Complex

The administration complex includes two 24' x 60' modular office trailers and a 10' x 20' wash

car. The complex includes space for administration functions as well as laboratory space and

an employee change room.

18.2.6 Rail Spur Line and Loading Area

A Union Pacific (UP) railway is located adjacent south of the CPF site. Railcars picked up

from site will be transported to the rail yard in Blackfoot, Idaho, which can provide product

transport to any location in North America.

18.2.7 Hydrometallurgical Facility Fire Protection

Fire protection consists of a buried fire main and fire hydrants located around the perimeter of

the facilities. Hand-held fire extinguishers are located throughout the facilities as additional

fire protection.

The solvent extraction area has an automatic-spray fire suppression system for each

mixer/settler and has an automatic foam system that will suppress fire in upper and lower decks

in surrounding rooms. A manual override system will start the system manually.

18.2.8 Power Supply and Distribution

Power supply to site is provided by the Idaho Power company from the 15 kV overland power

lines located 100 m north of the CPF site. Idaho power has confirmed their system will

accommodate the estimated additional loading from the CPF plant and no upgrades will be

required.

The electrical and control room is housed in a two story steel and concrete block. The building

houses the stepdown transformers, switch gear, MCCs and capacitor banks on the ground floor.

A mezzanine level contains the control room and DCS system, UPS, and battery room.

Buried cables bring the power to inside the electrical room. The electrical room has a 4.160

kV switch gear and MCC for distribution to the plant air compressor system. A 4.160 kV-480

V substation including a and 2500 kVa 60 hz 3 phase dry type transformer is used for general

plant distribution to the refinery.

Four MCCs are installed, with one dedicated to the 4.160 kV air compressor system and the

remaining three for the 480 V system. A capacitor bank is included for the 4.160 kV system.

Cable trays are used to distribute cabling throughout the plant along two main branches. Cable

trays in the SX building are fiberglass to provide corrosion resistance. All other cable trays are

galvanized steel.

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18.2.8.1 Emergency Power

A 480 V - 500 kW emergency diesel generator provides power to the following areas:

Fire protection system in the solvent extraction building.

Ventilation system in the cyanide system.

Agitators in the autoclave.

Air compressor system.

The NOx cooling loop system.

18.2.9 Process Control System

The overall CPF operating philosophy is that the whole plant shall be controlled from a Facility

Control Room (FCR). It is possible to monitor and control all the processes equipment of the

CPF from this FCR. The FCR will be continuously manned by operators during operations.

Remote control panels are installed at a number of key locations within the CPF to allow

operators the ability to monitor and control specific equipment.

A UPS system is used to ensure complete supervision of the process and allow for safe plant

shutdown in the event of a power outage.

18.2.10 Communications

The CPF is located near Blackfoot Idaho and tie-ins to the existing data and voice

communications networks are available within 100' of the site boundary. This includes the

availability of mobile and land lines as well as network connectivity. Communication at the

CPF administration building are primarily via land line or cell phone communications. Plant

operators and yard personnel use two way radio communication within the plant.

18.2.11 Water Supply

A 12” water line from Blackfoot is located on Pioneer road which will service the domestic

and process water needs of the plant. In total, the CPF requires approximately 26 T/h (103

USgpm) of water make-up to sustain the operation during steady state (see Table 18.1).

Table 18.1

Total CPF Make up Water

Total Water Requirement to CPF Flowrate (T/d)

Water Make-up 383

Unaccounted Water Requirement 236

Total 619

Flowrate, T/h (USgpm) 26 (103)

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The water balance calculated by the METSIM model is well accounted with the exception of

6% (~24 T/d), difference between the water in and water out flowrates. This difference is

assumed to be due to venting and other evaporation losses, and chemically bound water in the

solid products and residue streams, which are difficult to model and quantify accurately.

18.2.12 Waste Disposal

Water effluent from the facility will be disposed of in the Blackfoot municipal sanitary sewer

lines located under Pioneer Road in front of the CPF. The effluent from the plant is estimated

to be approximately 16 gpm. Liquid effluent from the CPF complies with the City of Blackfoot

Sewer Department criteria.

Solid waste residue from the facility will be collected for transport to an offsite disposal

facility. Approximately 160 lb per day of solid residue (dry basis) will be generated at the

facility.

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19.0 MARKET STUDIES AND CONTRACTS

19.1 INTRODUCTION

The feasibility study is based on the recovery of battery grade cobalt sulphate heptahydrate,

together with copper sulphate, magnesium sulphate and gold, and a minor volume of copper

concentrate as saleable by-products.

CRU Consulting (CRU) was retained by eCobalt to provide market analysis for cobalt sulphate

and by-products. CRU’s report, “Market Study for the Idaho Cobalt Project (ICP)”, dated

September, 2017, includes the following:

An assessment of the battery market and the technologies in use and under development

to support electric vehicles and other rechargeable battery applications.

Analysis of the market for cobalt, with particular emphasis on the use of cobalt sulphate

in the battery market.

Analysis of the current and future supply of cobalt sulphate and accessibility of that

market to the ICP.

An assessment of the market for the associated by-products of the ICP (i.e., copper

sulphate, magnesium sulphate, gold and copper concentrate).

CRU was founded in 1969 and provides market analysis and price assessments for a range of

metals, minerals and fertilizers, including minor and specialty metals and products. Based in

London, United Kingdom, it is a well-regarded, independent consulting firm.

Readily-available general information on cobalt supply and demand is published by

organizations such as the United States Geological Survey (USGS). However, detailed

information on cobalt chemicals and cobalt sulphate in particular, as well as copper sulphate

and magnesium sulphate, is available only through specialist consultancies such as CRU.

CRU has provided reasoned analysis of the markets for products from the ICP to support its

projection of unit prices on an ex-works basis. The following descriptions are based on that

report.

19.2 COBALT

The majority of cobalt is recovered as a by-product of nickel and copper. Historically, the most

important cobalt end-use sector was in the superalloys used to make parts for gas turbine

engines. By around 2005, growth in the rechargeable battery sector resulted in cobalt

consumption in rechargeable batteries matching that in superalloys. Over the past decade,

demand for cobalt in non-metallurgical applications, driven by demand for batteries, has

outpaced that in metallurgical applications. CRU estimates that cobalt demand was

approximately 96,000 t in 2016, of which approximately 44% was consumed in rechargeable

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batteries. Demand growth for rechargeable batteries has been led by the electric vehicle (EV)

sector, including plug-in hybrid electric vehicles (PHEV) and hybrid electric vehicles (HEV).

Over the period to 2026, CRU’s analysis indicates that global refined cobalt demand will

increase to approximately 165,000 t/y from 96,000 t in 2016. Between 2016 and 2021, the

compound average growth rate (CAGR) is expected to be 6% based on continued growing

demand for lithium-ion batteries; between 2021 and 2026, the rate of growth is expected to

moderate to around 4% as the EV sector matures and the metallurgical sector continues to

show robust growth. Within this overall demand projection, demand in non-metallurgical

applications will continue to outpace total demand growth throughout the forecast period.

The lithium-ion battery sector will overshadow growth in most other applications and CRU

projects non-metallurgical applications to account for 67% of global demand in 2026. Given

the length of time required for development, CRU does not anticipate that other battery

technologies for the EV sector will negatively affect cobalt consumption within the next 10

years.

Mined production of cobalt is dominated by the Democratic Republic of Congo (DRC) which

accounted for approximately 54% of world output in 2016 (USGS, Mineral Commodity

Summary, 2017). CRU expects the country’s share of output to increase to 67% by 2021,

despite political instability, required infrastructure development and risks to energy supply.

China is the largest importer of cobalt concentrates and intermediate products and is the largest

producer of cobalt chemicals.

CRU’s analysis of cobalt demand considers both metallurgical and non-metallurgical end-use

sectors; analysis of supply takes account of mine and refinery output, and secondary sources,

including recycling of catalysts and batteries; and stockpiled material.

19.2.1 Cobalt Sulphate

Battery-grade cobalt sulphate is produced by the following processes:

Dissolution of refined cobalt metal.

Dissolution of cobalt powder.

Refining cobalt concentrates.

Refining cobalt hydrometallurgical intermediates.

Recovering and refining recycled material.

China is by far the largest producer of cobalt sulphate and the majority is produced by refining

concentrates and hydrometallurgical intermediates. Outside China, cobalt sulphate mainly

takes place through dissolution of refined metal and powder.

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Three types of lithium-ion batteries, lithium cobalt oxide (LCO), lithium-nickel-cobalt-

aluminium-oxide (NCA) and lithium-nickel-manganese-cobalt (NMC) use cobalt in the

cathode. While LCO batteries have a higher cobalt content than NCA and NMC batteries, it is

these two which have dominated the growth in cobalt demand over the past three years and

both rely specifically on cobalt sulphate as a key component of the cathodes.

These contain between 7% and 20% cobalt in their cathode material and are used in a range of

applications. As well as EVs, they are also used in e-bikes and power tools. A minor amount

of sulphate is also consumed in copper solvent extraction electrowinning, the plating industry

and as a supplement in animal feeds.

Cobalt sulphate demand is rising strongly and is likely to outperform demand for other cobalt

chemicals and, in fact, demand in metallurgical applications in the future.

CRU estimates global cobalt sulphate consumption at 14,544 t contained metal in 2016, a

25.3% y/y increase. It is driven by strong growth in the electric vehicles (EV) sector; in

particular, 23.7% y/y increase in EV, plug-in hybrid electric vehicles (PHEV) and hybrid

electric vehicles (HEV) production.

Energy storage has become an important part of the electricity generation, transmission and

distribution chain, due to the surge in renewable energy generation systems globally. CRU

includes consumption in this end-use sector in its demand forecast. However, it believes that

there are number of competing technologies that remain more cost effective, efficient or safer

than the use of cobalt-bearing batteries for grid storage and, therefore, includes only minimal

consumption for this end-use.

CRU recognizes that there are a number of alternative technologies in existence or in

development that could challenge future cobalt demand, particularly if long-term supply

continues to be unstable for manufacturers. Low availability could make manufacturers turn

away from the use of cobalt-rich cathode chemistries, particularly in EVs. This will not

necessarily destroy cobalt demand, but could result in more battery material makers

endeavouring to reduce the amount of cobalt per unit in their NCA and NMC batteries.

However, CRU does not anticipate that these new technologies to significantly affect cobalt

demand in the next 10 years, although the intensity of use of cobalt in batteries may decline

slightly in the 2020s.

Over the next 10 years, CRU foresees that tightness in both the metallurgical and non-

metallurgical cobalt sectors will result in prices above current levels. CRU projects that the

shortfall will be more pronounced in the non-metallurgical sector where supply is expected to

increase at CAGR 7.0% compared with demand increasing at CAGR 7.9%. CRU notes further

that the global supply of cobalt chemicals is increasingly subject to bottlenecks in mine supply,

resulting in upward pressure on prices for cobalt chemicals.

As a result of its analysis, CRU concludes that the ICP has the opportunity to become a reliable

source of cobalt sulphate to markets within the United States and internationally. As noted

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above, the United States is a net importer of both copper sulphate and magnesium sulphate and

the ICP should be in a position to replace a proportion of imports in regions where it has a

freight advantage.

19.3 COPPER SULPHATE

Copper sulphate is usually sold in technical and animal feed grades, and as anhydrous copper

sulphate. It is widely used as a fungicide on fruit crops such as melons, grapes and other berries.

As an algaecide, it is used to control algal growth in lakes and reservoirs. As a plant nutrient,

copper sulphate is used to address copper deficiency in cereals. In animal feeds, it promotes

weight gain and feed efficiencies in poultry and pigs.

Copper sulphate is also used in a wide variety of industrial applications including adhesives,

timber preservation, colours and dyeing and in electrolytic processing.

Domestic United States production of copper sulphate has averaged 22,000-23,000 t/y since

2010. CRU’s analysis of trade data shows that the United States is a net importer of

approximately 30,000 t/y and it indicates that material from ICP will have the opportunity to

displace imported material in market regions where there is a freight advantage.

19.4 MAGNESIUM SULPHATE

Magnesium sulphate occurs naturally as the mineral kieserite (MgSO4.H2O) and in a number

of other naturally-occurring double salts of magnesium with potassium, sodium and calcium.

It is also produced synthetically by reacting magnesium oxide, magnesium hydroxide or

magnesium carbonate with sulphuric acid.

It is used as a fertilizer to correct magnesium deficiency in soils and for certain crops, such as

potatoes, roses, tomatoes, lemon trees, carrots and peppers which require additional

magnesium to support crop quality.

Synthetic magnesium sulphate is used in food processing, pharmaceuticals and a wide range

of industrial applications including detergents, leather, metal plating and pulp and paper.

Production in the United States averages around 45,000-48,000 t/y CRU’s analysis of trade

data shows that the United States is a net importer of approximately between approximately

12,000 and 21,000 t/y and it indicates that material from ICP will have the opportunity to

displace imported material in market regions where there is a freight advantage.

19.5 GOLD

Gold is a readily marketable metal and, in contrast to the sulphate products, pricing is

transparent. Supply is made up of mined gold; secondary, recycled gold recovered from

previously fabricated products; gold released from the official sector (gold bullion reserves

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held by central banks, government and supranational bodies), sales or leasing arrangements;

and gold released by producer hedging operations.

Gold demand is broadly divided into fabrication demand and investment demand. Fabrication

demand includes demand for manufacture of items such as jewellery, coins and medallions,

most of which is partly driven by the investment considerations, and gold used in electronics

and dental applications, and in minor industrial uses. Investment demand includes ingot and

bullion purchased by central banks, governments and other institutions.

19.6 COPPER CONCENTRATE

Copper concentrate produced by the ICP is expected to grade 32% Cu, which is above the

standard 30% Cu grade, and also will have favourable copper:iron:sulphur ratios. CRU

considers that since the projected production volumes are relatively small, the material will be

acceptable to smelters and/or traders in the United States for blending to take advantage of the

high copper content and to reduce contaminants such as arsenic.

19.7 PROJECTED REVENUE AND MARKET POSITION

Based on the CRU analysis, the following LOM average prices have been used in the financial

evaluation of the ICP project:

Cobalt sulphate $26.65/lb contained Co (average premium of $1.47/lb)

Copper sulphate $4.00/lb contained Cu (premium of $1.4/lb)

Magnesium sulphate $250/t

Gold $1,200/oz

Copper $5,732.01/t (base case copper price)

Micon has reviewed the CRU report and supports its rationale for projections of unit revenues

for cobalt, copper and magnesium sulphates, gold and copper in concentrate.

The projected breakdown of revenues is shown in Figure 19.1.

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Figure 19.1

ICP Breakdown of Projected Revenue

eCobalt, 27 September, 2017 press release.

19.8 CONTRACTS

FCC has not entered into any material contracts relating to development of the ICP.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT

The Idaho Cobalt Project (ICP) can be divided into the mine/mill complex and the refinery,

known as the Cobalt Processing Facility (CPF). The mine/mill complex lies 38 km west of the

town of Salmon, Lemhi County, Idaho and the CPF will be located 5 km northwest of the town

of Blackfoot, Bingham County, Idaho.

20.1 ENVIRONMENTAL BASELINE STUDIES AND IMPACT ASSESSMENTS

20.1.1 Mine and Mill

The mine and mill are located on National Forest lands managed by the Salmon-Challis

National Forest. As such, it is subject to the National Environmental Policy Act (NEPA). This

requires a thorough series of environmental baseline studies and an Environmental Impact

Statement that provides the Company, state and federal government agencies a complete

property description, identification of all environmental impacts both positive and negative and

the development of mitigation methods to reduce or eliminate negative impacts utilizing best

practices.

The lead government agency was the US Department of Agriculture, Forest Service, Salmon-

Cobalt Ranger District, Salmon-Challis National Forest (SCNF) with the US Environmental

Protection Agency (US EPA), the Idaho Department of Environmental Quality (IDEQ) and

Tribal governments being major cooperating agencies.

The Baseline Studies included:

Air Quality.

Subsurface Geology.

Surficial Geology (Soils and overburden).

Hydrology.

o Surface Water Quality and Quantity.

o Groundwater Quality and Quantity.

Wetlands and Other U.S. Waters.

Aquatic Resources (Fisheries, Threatened, Endangered and Candidate Species).

Vegetation Resources (Invasive, Threatened, Endangered and Candidate Species).

Wildlife Resources (Threatened, Endangered, Candidate Species and FS Management

Indicator Species).

Land Use Management (Recreational, Visual, Wilderness Resources).

Cultural and Historic Resources.

Noise and Light.

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Transportation.

Socio-Economic Resources.

Blackbird Mine – CERCLA

The studies examined threatened and endangered species of wildlife (Canada Lynx, Grey

Wolf, Bald Eagle and Yellow-billed Cuckoo), sensitive species (Wolverine, Fisher, Northern

Goshawk, Three-toed Woodpecker, Spotted Frog), management indicator species (Greater

Sage Grouse, Pileated Woodpecker, Spotted Frog), migratory birds, big game (Elk, Deer,

Moose, Black Bear, Big Horn Sheep, Mountain Goats) and other species.

The Draft Environmental Impact Statement (DEIS, February, 2007) identified several

significant and non-significant issues, based on the above Baseline Studies. The DEIS only

analysed significant issues which included:

Blackbird Mine CERCLA remediation and restoration.

Groundwater quality of the Panther Creek watershed.

Surface water quality of the Panther Creek watershed.

Water use, management, treatment and disposal.

Sediment delivery (storm water management).

Roads and access.

Transport of product, chemicals and fuel.

Social-economics.

Vegetation and reclamation.

Wetlands and other waters of the US.

Fish populations and their habitat.

Air quality, visual resources, wilderness experience.

Wildlife populations and their habitat.

Cultural resources and Tribal Trust responsibilities.

Planning (land use).

Issues deemed insignificant included:

Claim validity.

Water rights.

Public access and recreation.

Soil productivity.

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The Final Environmental Impact Statement (FEIS, June, 2008) discussed the project,

alternatives to the project, environment effects (direct, indirect and cumulative) and

consultation with aboriginal groups, communities and other stakeholders. No issues were

identified that could not be mitigated using best practices.

Alternative IV was identified by SCNF as the preferred alternative to the original Plan of

Operation that would meet the objectives of the Company as well as reduce resource impacts

to surface water, groundwater, wetlands and native vegetation.

Alternative IV modifications include reducing the size of the tailings disposal site to match

existing ore reserves and avoid direct impacts to isolated wetlands, modification of the

groundwater capture system to ensure adequate post closure groundwater capture,

modification of the proposed water treatment system to reduce the volume of water treatment

waste products while meeting NPDES permit requirements for the discharge to Big Deer

Creek, addition of amendments to mine waste backfill to improve long-term geochemical

stability and re-routing of the water discharge pipeline to avoid impacts to a cultural site.

An extensive environmental monitoring plan has been developed by the Company (Formation,

2015). The plan covers the following:

Water Quality Monitoring.

Biological Monitoring.

Wetlands Monitoring.

Storm Water Monitoring.

Weather Monitoring.

Air Quality Monitoring.

Geochemical Monitoring.

Figure 20.1 identifies the groundwater sampling locations, and the surface water and spring

monitoring network is illustrated in Figure 20.2.

Figure 20.3 shows a plan view of the tailings and waste rock storage facility.

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Figure 20.1

Groundwater Sampling Locations

199

Figure 20.2

Mine and Mill - Surface Water and Spring Monitoring Network

200

Figure 20.3

Mine and Mill - Tailings and Waste Storage Facility

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20.1.2 CPF

The processing facility is on private land and located within agricultural and industrial areas.

Therefore, it is not subject to the NEPA process or extensive environmental baseline studies.

However, the CPF will be required to meet certain other county, state and federal permitting

requirements, as described in Section 20.5.2 (below).

20.2 SOCIAL COMMUNITY RELATIONS

20.2.1 Mine and Mill

Both the Company and SCNF have conducted numerous consultation meetings with the local

community, aboriginal groups, stakeholders and special interested groups and non-

governmental organizations. Impacts, whether real or perceived, have been recorded and

addressed during meetings and in the EIS and the Record of Decision (RoD) processes.

Aboriginal groups (American Indian Tribes) that have been consulted throughout the process

are the Shoshone-Bannock and Nez Perce Tribes. Consultation included meetings with Tribal

councils and their staff, periodic project update letters and on-site tours of some of their

representatives. They have had full access to all project documents on the project. SCNF did

not identify any traditional cultural properties in the project area.

20.2.2 CPF

The CPF site has only been selected relatively recently and, therefore, consultation with the

surrounding community, the County of Bingham, the City of Blackfoot and other stakeholders

has been less extensive than at the mill/mill site. One potential stakeholder might be the

Danskin Ditch Company, that is part of the United Canal Company and is the local authority

for this area. Although the site itself is presently zoned as industrial, neighbouring agricultural

land is irrigated utilizing a system of irrigation canals to supply water to the area.

20.3 PLAN OF OPERATIONS

The Forest Service regulations at 36 CFR, Part 228 Subpart A required that the mine/mill be

operated in accordance with an approved Plan of Operations (PoO).

The original PoO, dated June 2006, included mining the RAM and Sunshine deposits. It was

updated November, 2008 and submitted to the SCNF. It was approved in January, 2009.

The current PoO, dated December 2009, is based on the recommendations from the FEIS

Alternative IV and the revised Record of Decision (RoD), dated January, 2009 under which

only the RAM deposit will be mined. A modified PoO, that includes the operations at the

Sunshine deposit, will therefore need to be submitted before a new RoD can be approved. The

total disturbance of the RAM mine and mill complex was estimated to be approximately 53 ha

(132 acres).

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An Inter-Agency Task Force (IATF) was established to oversee the mine/mill complex as it

progresses. The IATF comprised SCNF, IDEQ, US EPA, National Marine Fisheries Service

(NMFS), US Fish and Wildlife Service (US FWS), Nez Perce Tribe and Shoshone-Bannock

Tribe. The Task Force has not met since May, 2011 due to the project being put on care and

maintenance. The IATF will reconvene when construction of the mine/mill resumes.

20.3.1 Tailings and Waste Rock Storage Facility

Dry stacking will be employed for tailings not scheduled to go underground. The tailings and

waste rock storage facility (TWSF) will combine the waste rock and dry stack tailings into one

facility (see Figure 20.3, above). Waste rock from the RAM deposit, not being used

underground will be deposited in part of the TWSF underground workings. Tailings, not being

used as underground backfill will come from the mill. The TWSF will serve to minimize

surface disturbance by sharing containment of the waste rock and tailings and also the drainage

collection systems.

The majority of the waste rock will be quartzite which is not expected to present an acid

generation problem since it has a low pyrite percentage and hence low sulphide content.

However, some waste rock will contain sulphides that have a low buffering capacity as well as

soluble metals. The design of the TWSF will include liners and a drainage system that will

capture any acid rock drainage and soluble metals (ARD/ML) for treatment before effluent is

released into the environment. The tailings, on the other hand are almost entirely void of acid

generating capacity due to the mill processing system. It is felt that the tailings will assist in

encapsulating waste rock to reduce and possibly eliminate ARD/ML generation.

The final design of the TWSF will include:

A closure cap that includes a minimum of 1.2 m (4 ft.) of soil cover material to protect

the liner from potential damage from trees growing on the reclaimed surface.

A plan for placement of tailings into the TWSF during winter designed to maintain the

design density and moisture content of the dry stack tailings.

Co-disposal of tailings and waste rock in the TWSF to reduce the oxidation rate of the

higher permeability waste rock component and reduce log-term risk to the environment

of metals release.

A QA/QC Plan will specify that construction monitoring will proceed under the supervision of

a qualified professional engineer. The TWSF will be constructed east of and downslope from

the mill on the Big Flat area. This location was chosen as the best site for the facility in the

project area because of its relatively flat topography, avoidance of jurisdictional wetlands, soil

characteristics, and distance from active drainages and streams.

Specific design elements of the TWSF are:

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A designated facility footprint approximately 35 percent greater than the current RAM

deposit production estimates.

Separation of tailings and waste rock to the extent practicable.

Waste rock will be placed in designated portions of the TWSF in 2 m (5 ft) compacted

lifts.

Waste rock will be covered with at least a 1.2 m (3 ft) thick tailings cap annually.

Dewatered tailings will be placed and compacted in 0.8 m (2 ft) lifts to 90 percent of

the Modified Proctor maximum density.

A composite liner system with drainage collection.

Staged construction and reclamation.

A collection of runoff from waste rock and tailings with conveyance to the water

management pond.

A snow removal storage area with conveyance to the water management pond.

A diversion of runoff around the operating areas of the facility.

Geotechnical stability of the TWSF.

The configuration of the TWSF will separate the tailings from waste rock, except for a co-

disposal zone where the waste rock will be encapsulated by tailings. The TWSF will have a

slope of 4 horizontal to 1 vertical (4H:1V) side slopes constructed in three 15-m (50 ft.) raises

with two 30-m (100-ft) wide benches. The 4H:1V side slopes and benches will enhance

erosional and structural stability. A toe berm will be constructed at the base of the tailings

facility to provide containment for seepage and runoff water from the tailings stack and to

enhance geotechnical stability. The facility will occupy an area of approximately 22.3 ha (55

acres) and will measure approximately 488 m by 518 m (1,600 ft. by 1,700 ft.). The stack will

reach a maximum depth of about 27 m (90 ft.).

20.3.2 Water Management

Water management is based on operating a water treatment plant and releasing water in

accordance with an NPDES permit (Table 20.1). The water treatment plant will have the ability

to treat up to 568 L/min (150 g/min) of water for discharge through the NPDES Outfall 001.

Except during periods of very high inflow, the water treatment plant will treat incoming water

on an as-received basis, with very little water being stored in the water management pond.

During periods of high inflow, water will accumulate in the water management pond for

treatment during lower inflow periods.

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Table 20.1

Water Treatment Concentrations and Limits

Constituent Influent Concentration

(mg/L)

NPDES Limits

(mg/L)

Removal Target

(%)

Alkalinity (CaCO3) 0.0

Aluminum (Al) 0.122

Ammonia (NH3) 3.0 2.8 7

Arsenic (As) 0.093 0.01 89

Cadmium (Cd) 0.0

Calcium (Ca) 44.0

Chloride (Cl) 1.0

Cobalt (Co) 0.287 0.0704 75

Copper (Cu) 0.032 0.0024 93

Fluoride (F) 0.2

Iron (Fe) 1.0

Lead (Pb) 0.0

Magnesium (Mg) 67.0

Manganese (Mn) 5.6

Mercury (Hg) 0.0

Nickel (Ni) 0.003

Nitrate – (NO3-) 25.0 10.0 60

Potassium (K) 184.0

Silica (SiO2) 5.0

Sodium (Na) 131.0

Sulphate (SO4) 556.0

Thallium (Tl) 0.0

Zinc (Zn) 0.044 0.01845 58 Source Bruner et al, 2016

20.3.2.1 Water Balance

The water balance is unchanged from the Plan of Operations, as presented in the 2008

feasibility study, and is based on the flow diagram shown in Figure 20.4. A dynamic system

model (DSM) has been developed for the project that considers the relationships between the

project components and predicts the impact on them throughout the life of the mine (Telesto,

2005). It allows for the determination of storage requirements, based on the water treatment

plant capacity, to maintain a balanced system.

The DSM includes specific water balance calculations for each year of the project’s life. Each

year of the project’s life is unique, with variations in precipitation, ground water inflow

variances into the mine, and degree of build-out of the TWSF. As such, there is no typical year.

20.3.2.2 Water Treatment

The project water has been determined to contain elevated concentrations of nitrate, sulphate,

aluminum cobalt, copper, iron manganese and zinc. There has been a review of several

advanced technologies (Apex, Veolia and Linkan, Bruner et al, 2016), see Table 20.2.

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Figure 20.4

Mine/Mill Water Balance

Table 20.2

Water Treatment Systems Comparison

Characteristics Apex Veolia Linkan

Operator requirements 1 1 1

Operator qualifications Medium Medium Low

Maintenance Medium High Low

Outside support required Low Medium Low

Filtration type MMF UF Bag Filter

Filtration complexity Low High Low

Nitrogen treatment Packaged MBBR IX

Treatment sensitivity Medium Medium Low

IX waste volume (L) 78,737 85,433 111,848

System footprint (m2) 2,428 1,214 1,133

Capital Cost (2016) $4.743 M $6.467 M $3.014 M

O&M Cost (2016)/ yr $323,300 $239,754 $375,000

Present Value (15 yrs/6%) $7.883 M $9.148 M $6.656 M Notes: MMF = multi-media filtration, UF = ultra-filtration, IX = Ion exchange, O&M = Operation and

Maintenance, MBBR = moving bed biofilm reactor. Source Bruner et al, 2016

206

The water treatment system will incorporate the following stages:

pH adjustment and oxidation followed by clarification and bag filtration to primarily

remove iron, manganese and arsenic and reduce the turbidity of the effluent. Solids

would be sent to the TWSF.

Oil and grease removal.

Ion exchange (IX) with a synthetic resin that absorbs cobalt, copper and zinc.

Nitrogen removal by IX with a synthetic resin that absorbs nitrates rather than

biological processes.

After these steps, monitoring the results, and with approval from NPDES, the water will be

released to the environment.

20.3.3 Reclamation – Closure

The estimated closure and reclamation costs for operations under the approved PoO for the

mine/mill were determined using a present net value analysis and were based on the

preliminary designs available, including those from the PoO. The preliminary bond required

will be reviewed annually by SCNF, to ensure it is adequate to cover all reclamation costs.

Components included in calculating the financial assurance include:

Interim operations and maintenance.

Hazardous materials removal and disposal.

Operational water treatment.

Demolition and disposal.

Site re-grading, capping and other earthwork.

Revegetation.

Groundwater capture.

Post-closure operations and maintenance.

Post-closure water treatment.

Indirect and overhead costs.

The estimated financial assurance requirement for the ICP in the 2009 RoD was estimated to

be US$43.9 million dollars, plus or minus 20 percent. A detailed closure cost analysis was

completed in 2006, and updated in 2009. These costs have been escalated to 2017 dollars,

resulting in a total reclamation/closure cost of US$40.2 M. This cost includes, as required by

government, the post closure water treatment costs over the 100-year treatment period.

207

Bonding for surface disturbance is understood to have been put in place to cover the initial

construction reclamation costs of US$6.4 M. Provisions for bonding in respect of post-closure

water treatment, insurance on the balance of closure costs, and final closure costs have been

made in the project cash flow projection as part of this study, and amount to approximately

US$5.6 M.

As the CPF is on private land, it does not require a reclamation bond being posted as a

prerequisite of operation from a government agency. However, when the facility ceases

operation, the company will need to remove all hazardous material and waste products from

the site. This cost should be less than approximately US$20,000. The material will have to be

disposed of in an approved government facility.

If the property is not sold to another industrial user, the company may want to dispose of the

buildings and infrastructure and re-grade the site to reduce potential future safety liability

issues.

20.3.4 Closure Considerations

The closure of the mine and mill will meet the requirements of the USFS and reduce and

eliminate future reclamation and liabilities on the site. A phased approach to the reclamation

is recommended to even out the company’s operating expenses. It is hoped that the post-closure

operations, maintenance and water treatment at the mine and mill will only be required for 5

to 10 years before the USFS releases the company from any future obligations, however there

is provision for treatment to continue for 100 years as mandated by the government.

20.4 CPF OPERATIONS

The CPF will not require a Plan of Operations since it is in an industrial area of Blackfoot,

Bingham County and outside of USFS and other federal and state lands.

Figure 20.5 shows a site plan of the proposed facility. Copper and cobalt concentrate will be

shipped by truck from the Mine-Mill complex to the CPF site. The operation of the plant is

described in detail in another section of the report. A rail spur from the Union Pacific Railway

will be incorporated in the plant site to accommodate supplies being brought to the plant. The

spur will also be used to ship product and solid waste off-site.

208

Figure 20.5

Cobalt Processing Facility, Blackfoot, Idaho

20.5 PERMITS

20.5.1 Mine/Mill

The mine and mill portion of the ICP is located on National Forest lands managed by the

Salmon-Challis National Forest. As such it is subject to the National Environmental Policy Act

(NEPA). The Forest Service regulations at 36CFR part 228 require that the mine/mill be

operated in accordance with an approved Plan of Operations. FCC received the approval for

the Plan of Operations on December 10, 2009 following the issuance of an Environmental

Impact Statement and Record of Decision from the Forest Service. Alterations to the original

Plan of Operations will require discussions with government agencies and a revised Record of

Decision. In addition to the approved Plan of Operations the mine/mill requires a number of

permits and authorizations to operate. These permits and authorizations are listed in Table 20.3.

Table 20.3

ICP Permits

Permit Agency Comment

Plan of Operations USFS December 10, 2009

Environmental Impact Statement USFS June, 2008

Record of Decision USFS April,2009

Road Use USFS December 18,2009, annually renewed

Individual Discharge Permit EPA – NPDES Issued April 1, 2009, administratively extended

209


Storm Water EPA – NPDES July, 8, 2012, requires on-going inspections.

Air Quality IDEQ Issued April 23, 2009

Stream Channel Alteration USACE, IDWR Issued March 28, 2012, Clarification July 15, 2012

Water Rights IDWR 75-13977, completed

Injection Well IDWR Not required until backfill is used underground

Dam Safety IDWR, USACE June 24, 2011

Public Water System IDHW Pending final design

404 – Water Discharge Pipeline USACE May 15, 2012

Septic System IDHW Issued November 4, 2009

Mine Identification Number MSHA 10-02221, issued June 23, 2011

Large Scale Development Lemhi County Issued May 20, 2009

Road Maintenance Agreement Lemhi County Issued May 11, 2009

Building Permit Lemhi County In progress

Note: USFS – United States Forest Service, Salmon Idaho

EPA – NDPES – Environmental Protection Agency - National Pollutant Discharge Elimination System

IDEQ – Idaho Department of Environmental Quality

USACE – United States Army Corps of Engineers

IDWR – Idaho Department of Water Resources

MSHA – Mine Safety and Health Administration

20.5.2 CPF

The Cobalt Processing Facility site has been selected, and is located near the town of Blackfoot,

Idaho. The permits and approvals required for this facility are minimal, with the Air Quality

permit being the most important and time consuming to acquire. The CPF will require the

following permits listed in Table 20.4.

Table 20.4

CPF Permits


Storm Water EPA. NPDES Pending

Air Quality IDEQ Pending (6-8 months)

Building Bingham County Pending (3-4 weeks)

Sanitary Water approval Bingham County Pending (3-4 weeks)

The Sanitary Water system will be hooked into the Blackfoot water treatment collection

system, thus an independent septic system, permit and approval will not be necessary.

The southern boundary of the CPF property fronts on a Union Pacific (UP) railway line. The

rail line is classified as an “Industry Parks, Leads and Other Customer Complexes”. A rail spur

will be incorporated into the plant to bring supplies in and product and possibly waste/residue

out of the property. Negotiations with UP will need to be finalized, as well as a Track

Agreement between the Company and UP. Approvals from the UP Regional Vice President,

the UP-Marketing Business Team, the UP-Network Access team and the UP-Business

Rationalization team. A lead time from infrastructure design to implementation is a minimum

18 to 24 months.

210

21.0 CAPITAL AND OPERATING COSTS

21.1 CAPITAL COST ESTIMATE

The LOM capital cost estimate is summarised in Table 21.1. The estimate is given in US dollars

($), with a base date of third quarter, 2017. Owing to rounding of the estimates, some totals

may not agree.

Table 21.1

LOM Capital Estimate

Area Initial Capital

$'000

Sustaining Capital

$'000

LOM Total Capital

$'000

Mining 22,463 70,661 93,124

Processing + Infrastructure 26,355 5,000 31,355

Indirect costs 8,764 0 8,764

Contingency 5,165 3,207 8,373

Sub-total Mine/Mill/Concentrator 62,748 78,869 141,616

Direct – CPF 88,861 5,000 93,861

Indirect – CPF 20,495 0 20,495

Contingency 14,644 0 14,644

Sub-total Cobalt Production Facility 124,000 5,000 129,000

Rehabilitation and Mine Closure 588 16,942 17,530

Total 187,336 100,810 288,146

The capital cost estimate for this project presented herein is considered to be at a feasibility

study level with an accuracy of + 15%/-15% and carrying a contingencies totaling

approximately 12% on initial capital and 9% on LOM capital expenditures.

21.1.1 Mining Capital Cost

It is assumed that a contractor will carry out all underground mining activities in the pre-

production and ramp-up phases, and therefore the purchase of a mobile equipment fleet may

be deferred until the third year of operations, when it will be treated as sustaining capital.

Ongoing development of the underground mine is also treated as a sustaining capital expense.

Table 21.2 provides a breakdown of the initial and sustaining mining capital expenditure for

the project, and Figure 21.1 shows the LOM annual mining capital expenditures.

Table 21.2

LOM Mining Capital Estimate

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Portal Bench and Retaining Wall 1705 655 0 655

U/G Power Supply and Distribution 1710 1,186 0 1,186

Mine Ancillary Facilities 1720 565 0 565

Explosive Storage 1724 60 0 60

211

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Ventilation Shaft 1731 0 0 0

Ventilation Fans 1732 240 0 240

Mine Dewatering 1750 754 0 74

Underground Communications 1760 105 0 105

Mining Equipment 1770 1,382 12,259 13,641

U/G Development 1780 17,516 58,402 75,918

Total Mining Capital Expenditure 22,463 70,661 93,124

Figure 21.1

LOM Annual Mining Capital Expenditure

21.1.2 Mill/Concentrator and Infrastructure - Direct Capital Cost

A breakdown of the capital costs estimate for the processing plant and associated infrastructure

is given in Table 21.3. Since maintenance costs are included in operating expenses, sustaining

capital is required only for the retro-fitting of a copper scalping circuit, as described elsewhere

in this report.

Table 21.3

Mill/Concentrator Capital Estimate

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Site Development and Common Systems 1100 13,510 0 13,510

Tram System 1200 3,472 0 3,472

Concentrating 1300 5,516 5,000 10,516

Tailings Disposal 1400 1,904 0 1,904

Tailings Waste Rock Storage Facility 1600 1,123 0 1,123

Mine Area Facilities (see Mining, above) 1700 0 0 0

Ancillary Facilities and Reagents 1800 441 0 441

Utilities 1900 389 0 389

Mill/Concentrator Capital Exp. 26,355 5,000 31,355

0

5,000

10,000

15,000

20,000

25,000

30,000

Yr-2 Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12

$'0

00

Mining Capital Expenditure

212

21.1.3 Indirect Capital Costs

The estimated indirect capital costs applied to the process and infrastructure are shown in Table

21.4. These costs include the estimated Owner’s costs for site management, recruitment and

training.

Table 21.4

Mine + Mill Indirect Capital Estimate

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Temporary Construction Facilities 2210 243 0 243

Construction Support 2220 1,429 0 1,429

Constr. Equipment, Tools and Supplies 2240 0 0 0

Pre-commissioning 2250 100 0 100

Freight incl. Duties 2260 544 0 544

Spares 2280 654 0 654

First Fills 2290 13 0 13

EPCM – Mine + Mill 3000 4,157 0 4,157

Mobile Equipment 4120 1,490 0 1,490

Communication Equipment/Systems 4130 54 0 54

Admin Furniture, Office Equipment 4140 50 0 50

Safety, First Aid and Security 4140 30 0 30

Indirect Capital – Mine + Mill 8,764 0 8,764

21.1.4 Contingency – Mine and Mill

In addition to the costs identified above for the mine and mill site, contingencies of $5.2 million

and $3.2 million were provided in respect of initial and sustaining capital expenditures

respectively.

21.1.5 Cobalt Production Facility – Direct Capital Cost

A breakdown of the capital costs estimate for the processing plant and associated infrastructure

is given in Table 21.5. Since maintenance costs are included in operating expenses, sustaining

capital is required only for the retro-fitting of a copper scalping circuit, as described elsewhere

in this report.

Table 21.5

CPF- Direct Capital Estimate

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Hydrometallurgical area 0 1,755 0 1,755

Refinery Building 200 17,175 0 17,175

Cobalt Concentrate Feed Preparation 210 2,736 0 2,736

Copper Scalping and Conc. Handling 220 1,312 5,000 6,312

Filtration/Gold Rec./Tailings Handling 230 8,956 0 8,956

213

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Acid Neutralization 240 3,034 0 3,034

Copper Solvent Extraction 250 3,389 0 3,389

Copper Sulphate Production 260 4,777 0 4,777

Cobalt Plant 600 4,721 0 4,721

Autoclave 610 6,483 0 6,483

Copper / Iron Removal 620 2,723 0 2,723

Cobalt Precipitation 630 3,378 0 3,378

Cobalt Re-dissolution 640 133 0 133

Cobalt Solvent Extraction 650 6,406 0 6,406

Cobalt Sulphate Production 660 5,749 0 5,749

Effluent Treatment 670 7,354 0 7,354

Reagents 680 4,536 0 4,536

Utilities 690 4,243 0 4,243

Total 88,861 5,000 93,861

21.1.6 Cobalt Production Facility – Indirect Capital Cost

The estimated indirect capital costs applied to the process and infrastructure are shown in Table

21.6. These costs include the estimated Owner’s costs for site management, recruitment and

training.

Table 21.6

CPF Indirect Capital Estimate

Area WBS

Code

Initial Capital

$'000

Sustaining Capital

$'000

Total Capital

$'000

Temporary Construction Facilities 2110 287 0 287

Construction Support 2120 1,637 0 1,637

Constr. Equipment, Tools and Supplies 2140 0 0 0

Pre-commissioning 2150 269 0 269

Freight incl. Duties 2160 2,612 0 2,612

Spares 2180 1,677 0 1,677

First Fills 2190 38 0 38

EPCM - CPF 3000 13,329 0 13,329

Mobile Equipment 4120 465 0 465

Admin Furniture, Office Equipment 4130 150 0 150

Safety, First Aid and Security 4140 30 0 30

Total 20,495 0 20,495

21.1.7 Contingency – CPF

In addition to the costs identified above for the cobalt processing facility, a contingency of

$14.64 million (13.4%) was provided in respect of initial capital expenditures. No contingency

was applied to the sustaining capital estimate.

214

21.1.8 Closure Costs

The costs of surface rehabilitation and post closure water treatment (PCWT) have previously

been estimated and those estimates have been discussed with the permitting authorities. For

the purpose of this study those estimates have been updated by applying appropriate

allowances for escalation to 2017 terms. It is anticipated that during construction a deposit of

30% of these estimated costs will be required (where not already in place) and that the balance

of this liability will be covered by purchasing insurance at a premium of 2% of the liability,

for an annual cost equating to approximately $0.2 million over the LOM period.

21.1.8.1 Surface Disturbance

A cost of $6.4 million has been estimated for rehabilitation of surface disturbances on mine

closure, and a deposit of 30% of that amount ($1.9 million) has already been paid and hence is

treated as a sunk cost. It is anticipated that in real terms the value of that deposit will grow at

2.8%/y for 12 years, leaving a balance of $3.7 million as the net cost of surface disturbance

rehabilitation to be expended on mine closure.

21.1.8.2 Post Closure Water Treatment

An annual cost of $0.56 million has been estimated for PCWT in 2017 terms, to be incurred

over the 100-year period following mine closure. Applying an annual discount rate of 2.8% to

this amount values of this liability at $18.8 million at the time of mine closure. Accordingly, a

deposit of 30% of that amount ($5.6 million) has been provided for in the cash flow forecast.

In real terms, it is anticipated that the value of that deposit will grow at 2.8%/y for 12 years,

leaving a balance of $10.9 million as the net cost of surface disturbance rehabilitation to be

expended on mine closure.

21.2 OPERATING COST ESTIMATE

The estimated life-of-mine total project operating costs are summarized in Table 21.7.

Table 21.7

Summary of LOM Operating Costs

Area LOM total Operating

Costs ($’000)

Unit cost

$/tonne milled

$/lb Contained

Co in sulphate

Mining 196,692 53.71 6.19


Transport (residue disposal) 5,199 1.42 0.16

Hydromet Plant (CPF) 149,121 40.72 4.69

G&A 37,309 10.19 1.17

Sub-total Direct Operating Costs 440,815 120.38 13.88


Total Cash Operating Costs

before by-product credits

442,932 120.96 13.94

Less By-product credits (282,510) (77.15) (8.89)

Cash Operating Costs (net) 160,422 43.81 5.05

215

21.2.1 Mining Operating Cost

Table 21.8 shows a breakdown of the mine operating cost estimate.

Table 21.8

LOM Mine Operating Cost Estimate

Area Stoping

$/t mined

LOM Total

($’000)

Unit cost

$/tonne milled

Unit cost

$/lb Cobalt

Development – horizontal 70,928 19.37 2.23

Development – vertical 3,864 1.06 0.12

Development – contingency 3,805 1.04 0.12

Sub-total Development 78,597 21.46 2.47

Cut & Fill Stoping 45.91 49,591 13.54 1.56

Long-Hole Stoping 30.62 78,739 21.50 2.48

Equipment Parts Costs 7,678 2.10 0.24

Fuel Costs 6,089 1.66 0.19

Power Cost 2,506 0.68 0.08

Mine Staff Costs 17,111 4.67 0.54

Mine Labour Costs 36,337 9.92 1.14

Total Mining Costs 276,648 75.55 8.71

Less Capitalized Development (75,918) (20.73) (2.39)

Less Pre-Production Mining Cont. (4,038) (1.10) (0.13)

Net Mining Costs 196,692 53.71 6.19

21.2.2 Mill/Concentrator Operating Cost

Operating costs for processing (Table 21.9) have been calculated on the basis of labour

requirements and consumption of power, reagents, grinding media and other operating

consumables and spares.

Table 21.9

LOM Mill/Concentrator Operating Cost Estimate

Area LOM Costs

($’000)

Unit cost

$/tonne milled

$/lb Contained

Co in sulphate

Power 8,205 2.24 0.26

Consumables 885 0.24 0.03

Maintenance 8,791 2.40 0.28

Operating Supplies 4,211 1.15 0.13

Tailings Disposal 293 0.08 0.01

Water Treatment 4,460 1.22 0.14

Supervision 2,376 0.65 0.07

Labour 23,274 6.36 0.73

Total 52,494 14.34 1.65

216

21.2.3 Cobalt Production Facility Operating Cost

Table 21.10 gives a breakdown of the CPF operating costs. These have been estimated on the

basis of labour requirements and the consumption of electrical power, natural gas, reagents,

water, and other operating consumables and spares.

Table 21.10

LOM CPF Operating Cost Estimate

Area LOM Cost

($’000)

CPF unit cost

$/ton conc.

Ore unit cost

$/ton milled

Unit cost

$/lb Co

Sulphuric Acid, 98% 7,894 48.97 2.16 0.25

MgO 25,978 161.16 7.09 0.82

Gen Reagents/Consumables 6,841 42.44 1.87 0.22

Activated Carbon 196 1.22 0.05 0.01

SX Reagents 4,081 25.32 1.11 0.13

Water & Water Treatment 866 5.37 0.24 0.03

Assay / Laboratory Consumables 1,509 9.36 0.41 0.05

Grid Power 8,350 51.80 2.28 0.26

Natural Gas 1,250 7.76 0.34 0.04

Labour 45,506 282.30 12.43 1.43

General Expenses 18,901 117.25 5.16 0.59

Maintenance Materials 9,423 58.45 2.57 0.30

Contract Services 4,769 29.59 1.30 0.15

Contingency 13,556 84.10 3.70 0.43

Total 149,121 925.10 40.72 4.69

21.2.4 Residue Disposal

In addition to the on-site cash operating costs listed above, a separate allowance is made in the

project cash flow model for transport of residue from the CPF to a registered toxic waste

disposal facility at a cost of $5.2 million over the LOM period, equating to $25/ton residue, or

$1.42/ton milled.

21.2.5 General and Administrative Operating Costs

General and Administrative (G&A) operating costs for the project are given in Table 21.11.

Table 21.11

LOM G&A Operating Cost Estimate

Area LOM Costs

($’000)

Unit cost

$/tonne milled

$/lb Contained

Co in sulphate

Supervision 3,717 1.02 0.12

Labour 3,717 1.02 0.12

Concentrate Transport 10,972 3.00 0.35

Transportation Expenses 6,418 1.75 0.20

Office expenses 1,975 0.54 0.06

Insurance 5,398 1.47 0.17

217

Area LOM Costs

($’000)

Unit cost

$/tonne milled

$/lb Contained

Co in sulphate

Corporate Services 1,382 0.38 0.04

Environmental 1,250 0.34 0.04

Right of Way and Land acquisitions 1,850 0.51 0.06

Legal Permits and Fees 461 0.13 0.01

Recruiting and Relocation 168 0.05 0.01

Total 37,309 10.19 1.17

21.2.6 Selling Costs and Royalty

Selling costs have been estimated for the elution and smelting of gold doré from loaded carbon

shipped from the CPF plant, as well as for the transport and treatment of copper sulphide

concentrates that will be produced intermittently in later years of the LOM production period.

The sulphates of cobalt, copper and magnesium are assumed sold from the CPF on FOB basis,

and so do not attract any additional selling costs.

There is no royalty payable on production from the ICP project.

218

22.0 ECONOMIC ANALYSIS

22.1 BASIS OF EVALUATION

Micon has prepared its assessment of the project on the basis of a discounted cash flow model,

from which Net Present Value (NPV), Internal Rate of Return (IRR), payback and other

measures of project viability can be determined. Assessments of NPV are generally accepted

within the mining industry as representing the economic value of a project after allowing for

the cost of capital invested.

The objective of the study was to determine the potential viability of the proposed development

of an underground mine and on-site mill and concentrator near Salmon, Idaho, and the

establishment of an off-site hydrometallurgical processing plant to further refine the product,

to be located on an industrial site in Blackfoot, Idaho. In order to do this, the cash flow arising

from the base case has been forecast, enabling a computation of the NPV to be made. The

sensitivity of this NPV to changes in the base case assumptions is then examined.

22.2 MACRO-ECONOMIC ASSUMPTIONS

22.2.1 Exchange Rate and Inflation

Price assumptions for each product and by-product are given in United States dollar ($) terms

and, unless otherwise stated, all financial results are also expressed in U.S. dollars. All material

capital and operating cost estimates and other inputs to the cash flow model for the project

have been prepared using constant, third quarter 2017 money terms, i.e., without provision for

escalation or inflation. Since these costs are estimated in U.S. dollars, no exchange rate

assumptions are relevant.

22.2.2 Weighted Average Cost of Capital

In order to find the NPV of the cash flows forecast for the project, an appropriate discount

factor must be applied which represents the weighted average cost of capital (WACC) imposed

on the project by the capital markets. The cash flow projections used for the valuation have

been prepared on an all-equity basis. This being the case, WACC is equal to the market cost

of equity, and can be determined using the Capital Asset Pricing Model (CAPM):

where E(Ri) is the expected return, or the cost of equity. Rf is the risk-free rate (usually taken

to be the real rate on long-term government bonds), E(Rm)-Rf is the market premium for equity

(commonly estimated to be around 5%), and beta (β) is the volatility of the returns for the

relevant sector of the market compared to the market as a whole.

Micon has applied a real discount rate of 7.5% as its base case.

219

22.2.3 Expected Metal Prices

The base case cash flow projection assumes a variable price of cobalt metal, with cobalt

sulphate heptahydrate (with a minimum grade of 20.5% Co) trading at a premium of around

2% on a 100% cobalt basis. The basis for these price assumptions are discussed in Section 19

of this report. Figure 22.1 shows the annual prices and premium applied for cobalt in sulphate.

Figure 22.1

Cobalt Price and Cobalt Sulphate Premium (Recent and Forecast)

Copper sulphate sales are forecast at a constant price of $2.60/lb Cu, with a premium of 54%

for the sulphate resulting in gross revenue of $4.00/lb Cu. Copper concentrate sales are forecast

with payability of 98%, treatment charges of $185/t including transport, and $0.10/lb Cu

refining. Gold revenue and credits are based on a price of $1,200/oz Au, and magnesium

sulphate sales are forecast on a price averaging $250/t MgSO4.

22.2.4 Taxation Regime

Idaho state and U.S. federal income taxes payable on the project have been provided for in the

cash flow forecast after deductions for relevant depreciation allowances.

The net taxes payable on the forecast project cash flow has been estimated by an independent

third party with specialist expertise in this area, and Micon has relied on this analysis in its

economic evaluation of the project.

22.2.5 Royalty

No royalty has been provided for in the cash flow model.

0.0%

1.0%

2.0%

3.0%

4.0%

5.0%

6.0%

-

5.00

10.00

15.00

20.00

25.00

30.00

35.00

Pre

miu

m (

%)

$/l

b

Co (99.3%) Premium for Sulphate (%)

220

22.2.6 Selling Expenses

Concentrate transport between the mine and the hydrometallurgical plant is included within

cash operating costs. Both the primary cobalt sulphate product and by-products are assumed

to be sold on FOB basis at the refinery.

22.3 TECHNICAL ASSUMPTIONS

The technical parameters, production forecasts and estimates described elsewhere in this report

are reflected in the base case cash flow model. These inputs to the model are summarised

below. The measures used in the study are metric throughout.

22.3.1 Mine Production Schedule

Figure 22.2 shows the annual tonnage of mill-feed material mined from underground, as well

as the mill head grades for cobalt, copper and gold content.

Figure 22.2

Annual Mining Schedule

As shown in the figure, the grade of the mill feed demonstrates the focus on higher cobalt

grades in the early part of the production period. Material with a relatively high copper/cobalt

ratio of 2.0 or more is extracted later in the mine life. Treatment of this material necessitates

the commissioning of a copper scalping circuit, the construction of which is provided for in

the sustaining capital estimate.

Annual production of cobalt and by-products over the LOM period is shown in Figure 22.3.

0.000

0.200

0.400

0.600

0.800

1.000

1.200

0.0

50.0

100.0

150.0

200.0

250.0

300.0

350.0

1 2 3 4 5 6 7 8 9 10 11 12 13

Gra

de

(p

pm

Au

, % C

o, C

u)

Mill

ed

(0

00

t)

Mill feed (ROM) Cobalt Grade Copper Grade Gold Grade

221

Figure 22.3

Annual Processing Schedule

Annual sales value of cobalt and by-products over the LOM period is shown in Figure 22.4.

Figure 22.4

Annual Sales Revenues by Product

Over the LOM period, cobalt sulphate sales account for 75% of total revenue. Copper sulphate

contributes a further 15%, magnesium sulphate 5%, gold 4%, and copper concentrates 1%.

0

5,000

10,000

15,000

20,000

25,000

30,000

0

2

4

6

8

10

12


MgS

O4

.7H

2O

(t)

Co

SO4

.7H

2O

, Cu

SO4

.5H

2O

, C

u-C

on

c (t

),G

old

(o

z)

Cobalt Sulphate Copper Sulphate Gold in doré Copper Conc. Magnesium Sulphate

0

20,000

40,000

60,000

80,000

100,000

120,000

140,000

160,000


Rev

enu

e ($

'00

0)

Cobalt Sulphate Copper Sulphate Magnesium Sulphate Gold in doré Copper Conc.

222

22.3.2 Operating Costs

Cash operating costs over the LOM period average $120.38/t milled, a breakdown of these is

presented in Table 22.1.

Table 22.1

Operating Cost Estimate

Area $/tonne

milled

Mining 53.71

Mill/Concentrator 14.34

Transport (residue disposal) 1.42

CPF 40.72

G&A 10.19

Sub-total Direct Operating Costs 120.38

Selling Costs 0.58

Total Cash Operating Costs before by-product credits 120.96

By-product credits (77.15)

Cash Operating Costs (net) 43.81

Figure 22.5 shows these expenditures over the LOM period.

Figure 22.5

LOM Cash Operating Costs

22.3.3 Capital Costs

Pre-production capital expenditures are estimated to total $186.75 million. This sum includes

$22.46 million for mining, $26.36 million in the milling/concentrator plant, $88.86 million in

0

5,000

10,000

15,000

20,000

25,000

30,000

35,000

40,000

45,000

50,000


Rev

enu

e ($

'00

0)

Selling Costs (Cu-conc, gold) Mining Mill/Concentrator Transport CPF G&A

223

the CPF, $29.26 million indirect costs and owner’s costs, and contingencies totalling $19.81

million.

Sustaining capital is estimated at $83.87 million over the LOM period, mainly for underground

development but including $10 million for retro-fitting a copper sulphide scalping circuit. A

further $17.53 million is required to cover mine closure and associated bonding costs.

Working capital has been estimated to include 30 days allowance for product inventory on site,

in transit, and accounts receivable on concentrates delivered. Stores provision is for 30 days of

consumables and spares inventory, less 60 days accounts payable. On this basis, an average of

$4.92 million of working capital is required during the mine/mill operating period.

22.4 BASE CASE CASH FLOW

The LOM base case project cash flow is presented in Table 22.2. Annual cash flows are

presented in Table 22.3 (over) and summarized in Figure 22.6 (following page).

Table 22.2


Item LOM total

($ 000) $/t milled $/lb Cobalt

Cobalt Sales 846,837 231.26 26.66


Mining 196,692 53.71 6.19


Transport 5,199 1.42 0.16

CPF 149,121 40.72 4.69

G&A 37,309 10.19 1.17




EBITDA 686,415 187.45 21.61

Capital Costs 288,146 78.69 9.07


Tax 66,814 18.25 2.10


224

Figure 22.6

Life-of-Mine Cash Flows

The project demonstrates an undiscounted pay back of 3.3 years, or approximately 4.0 years

when discounted at 7.5%, leaving a tail of over 8 years of production.

22.5 DISCOUNTED CASH FLOW EVALUATION

The base case evaluates to an IRR of 25.1% before taxes and 21.3% after tax. At a discount

rate of 7.5%, the net present value (NPV7.5) of the cash flow is $177 million before tax and

$136 million after tax.

225

Table 22.3

Base Case Life of Mine Annual Cash Flow

CASH FLOW PROJECTION Item Units Period LOM TOTAL Yr-2 Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12 Yr13 Yr14 Yr15

Copper Conc. Gross Value 5,851,152 0 0 0 0 0 0 0 0 0 0 2,039,169 3,408,287 403,696 0 0 0 0

Cobalt Sulphate sales 846,836,514 0 0 79,214,801 112,510,709 112,483,342 91,275,045 59,050,865 61,748,417 47,875,265 52,302,006 43,479,152 39,451,614 43,240,995 46,527,869 57,676,435 0 0

Copper Sulphate sales 171,443,319 0 0 11,995,984 19,717,931 21,236,391 8,098,312 14,056,980 14,655,342 9,690,790 13,931,110 12,258,519 11,153,688 12,146,740 10,044,793 12,456,739 0 0

Gold doré sales 47,088,764 0 0 4,935,350 6,129,511 4,958,518 3,292,667 2,886,725 3,631,479 2,387,962 2,732,931 2,740,186 2,925,747 3,731,593 3,635,870 3,100,226 0 0

Magnesium Sulphate 58,126,765 0 0 4,485,832 6,951,609 7,122,121 3,664,280 4,684,029 4,859,520 3,440,276 4,388,533 3,812,292 3,466,587 3,780,601 3,335,333 4,135,751 0 0

REVENUE Gross Revenue $ 1,129,346,514 0 0 100,631,966 145,309,760 145,800,371 106,330,304 80,678,598 84,894,758 63,394,293 73,354,581 64,329,319 60,405,922 63,303,625 63,543,865 77,369,152 0 0

OPERATING COSTS Selling Costs (Cu-conc, gold) 2,116,548 0 0 137,212 170,411 137,856 91,542 80,256 100,962 66,390 75,980 360,599 551,586 156,479 101,084 86,192 0 0

Royalties 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Mining 196,691,537 0 0 16,601,837 23,989,719 16,672,544 16,417,597 14,906,955 16,335,905 15,892,592 14,662,878 16,076,726 15,726,033 16,649,648 12,759,103 0 0 0

Mill/Concentrator 52,494,076 0 0 3,527,235 4,135,925 4,135,925 4,135,925 4,141,789 4,135,925 4,135,925 4,135,925 4,141,789 4,135,925 4,135,925 4,135,925 3,459,940 0 0

Bulk Conc.Transport 5,199,168 0 0 411,285 628,182 635,170 349,800 416,999 432,039 311,429 384,500 332,738 302,510 330,053 296,643 367,821 0 0

Hydromet Plant 149,121,267 0 0 12,686,305 14,932,633 14,188,944 13,147,732 11,551,564 11,666,885 10,848,798 10,552,245 9,921,699 9,633,656 9,892,557 10,129,800 9,968,449 0 0

G&A 37,309,244 0 0 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 3,109,104 0 0 0

Total Cash Operating Costs $ 442,931,840 0 0 36,472,977 46,965,975 38,879,542 37,251,700 34,206,667 35,780,819 34,364,237 32,920,631 33,942,655 33,458,813 34,273,765 30,531,658 13,882,402 0 0

Operating Margin (EBITDA) 686,414,674 0 0 64,158,989 98,343,785 106,920,829 69,078,605 46,471,931 49,113,938 29,030,056 40,433,950 30,386,664 26,947,109 29,029,860 33,012,207 63,486,750 0 0

CAPITAL COSTS Initial Capital 186,747,526 35,472,207 151,124,188 151,131 0 0 0 0 0 0 0 0 0 0 0 0 0 0

Sustaining Capital 83,868,504 0 0 28,886,719 23,102,958 9,130,733 1,783,078 948,410 4,170,624 1,103,035 10,981,153 817,132 1,148,367 1,239,794 556,500 0 0 0

Reclamation & Closure 17,529,960 294,000 294,000 5,824,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 181,140 14,587,140 -5,643,000

Working Capital Mvmt 0 0 -148,830 5,258,496 -545,305 3,677,173 -1,604,409 -1,686,652 223,197 -1,653,561 912,927 -800,436 -296,807 168,840 364,148 3,067,647 -6,936,426 0

Capital Invested $ 288,145,990 35,766,207 151,269,359 40,120,486 22,738,793 12,989,046 359,810 -557,102 4,574,961 -369,386 12,075,220 197,836 1,032,700 1,589,773 1,101,788 3,248,787 7,650,714 -5,643,000

CASH FLOW Net Cash Flow before tax $ 398,268,684 -35,766,207 -151,269,359 24,038,503 75,604,993 93,931,784 68,718,795 47,029,034 44,538,977 29,399,442 28,358,730 30,188,828 25,914,409 27,440,087 31,910,420 60,237,963 -7,650,714 5,643,000

Taxation Payable $ 66,814,474 0 0 783,747 9,961,692 11,695,597 9,504,507 5,603,469 6,318,879 2,611,889 5,897,555 4,419,584 4,062,031 4,530,354 6,096,016 -4,709,218 38,372 0

Net Cash Flow after tax $ 331,454,210 -35,766,207 -151,269,359 23,254,756 65,643,300 82,236,187 59,214,288 41,425,565 38,220,098 26,787,554 22,461,175 25,769,244 21,852,378 22,909,733 25,814,404 64,947,181 -7,689,087 5,643,000

IRR Payback

CUMULATIVE C/F Cum. Cash Flow before tax 25.1% 2.9 yrs -35,766,207 -187,035,565 -162,997,062 -87,392,070 6,539,714 75,258,509 122,287,543 166,826,520 196,225,962 224,584,692 254,773,520 280,687,929 308,128,016 340,038,435 400,276,398 392,625,684 398,268,684

0.0 0.0 1.0 1.0 0.9 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Cum. Cash Flow after tax 21.3% 3.3 yrs -35,766,207 -187,035,565 -163,780,809 -98,137,509 -15,901,322 43,312,966 84,738,531 122,958,629 149,746,182 172,207,357 197,976,601 219,828,979 242,738,712 268,553,116 333,500,297 325,811,210 331,454,210

0.0 0.0 1.0 1.0 1.0 0.3 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

NPV Discount

DISCOUNTED Net Cash Flow before tax 176,864,854 7.5% -34,495,993 -135,718,256 20,062,556 58,697,668 67,838,226 46,166,716 29,390,775 25,892,665 15,898,903 14,266,136 14,127,242 11,280,906 11,111,679 12,020,381 21,108,022 -2,493,853 1,711,080

Net Cash Flow after tax 135,760,458 7.5% -34,495,993 -135,718,256 19,408,440 50,963,680 59,391,579 39,781,390 25,888,890 22,219,195 14,486,421 11,299,314 12,059,042 9,512,647 9,277,142 9,724,064 22,758,182 -2,506,361 1,711,080

Payback

CUMUL. DISCOUNTED Cum DCF before tax 3.5 yrs -34,495,993 -170,214,249 -150,151,693 -91,454,024 -23,615,799 22,550,917 51,941,692 77,834,357 93,733,260 107,999,396 122,126,638 133,407,544 144,519,222 156,539,604 177,647,626 175,153,773 176,864,854

0.0 0.0 1.0 1.0 1.0 0.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Cum DCF after tax 4.0 yrs -34,495,993 -170,214,249 -150,805,809 -99,842,128 -40,450,549 -669,159 25,219,731 47,438,926 61,925,347 73,224,660 85,283,702 94,796,349 104,073,491 113,797,556 136,555,738 134,049,377 135,760,458

0.0 0.0 1.0 1.0 1.0 1.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

226

22.6 SENSITIVITY STUDY

The sensitivity of project returns to changes in all revenue factors (including grades, recoveries,

prices and exchange rate assumptions) and also to capital and operating costs was tested over

a range of 30% above and below base case values. See Figure 22.7, showing net present values

on an after-tax basis.

The chart suggests that the project is most sensitive to revenue drivers, moderately sensitive to

operating costs and least sensitive to changes in capital cost. Within a range of 30% above and

below base case values, operating and capital costs both maintain a positive NPV outcome.

Figure 22.7

NPV Sensitivity Diagram

22.7 CONCLUSION



assumptions.







(50,000)

0

50,000

100,000

150,000

200,000

250,000

300,000

70 75 80 85 90 95 100 105 110 115 120 125 130

$'0

00

Percentage of Base Case

Prices Operating Costs Capital Expenditure

227

23.0 ADJACENT PROPERTIES

The historical Blackbird mine is immediately adjacent to the ICP property. The Blackbird mine

is no longer in production, has undergone remediation and continues with water treatment for

the mine and tailings runoff waters.

Micon has not verified whether the mineralization on the defunct Blackbird mine is indicative

of the mineralization on the ICP.

228

24.0 OTHER RELEVANT DATA AND INFORMATION

All relevant data and information is presented in other sections of this report.

229

25.0 INTERPRETATION AND CONCLUSIONS

25.1 GEOLOGY AND MINERAL RESOURCES

The Ram deposit consists of a Hanging-wall Zone with 3 primary and 4 minor horizons, a

Main Zone comprising 3 horizons, and a Footwall Zone with 3 horizons. These sub-parallel

horizons generally strike N15oW and dip 50o-60o to the northeast. The overall mineral resource

is summarized Table 25.1 Table 25.1

Summary of the Ram Deposit Mineral Resources at 0.2% Co Cut-off

Category Co%

Cut-off

Resource

(Tons)

Co

(%)

Co

(lbs)

Au

(oz/t)

Au

(ounces)

Cu

(%)

Cu

(lbs)

M + I 0.2 3,436,000 0.59 40,577,700 0.016 54,200 0.73 50,435,500

Inferred 0.2 1,543,000 0.51 15,593,800 0.012 18,700 0.68 21,032,200

M + I = Measured & Indicated

The Main Zone (horizons 3021, 3022 and 3023) contribute about 87% of the Measured and

Indicated resource. However, the exact extents of the Hangingwall and Footwall Zones

horizons of the deposit remain to be fully investigated and could have material effect in terms

of increasing the resource and life of mine.

The mineralization of the Ram deposit remains open at depth (down-dip) and along strike. The

geological corridor/structure controlling the mineralization is persistent for the entire strike

length of FCC’s ICP area and beyond. The already known Sunshine deposit is within easy

reach (i.e., only one mile south) from the infrastructure at the Ram. Hence, the outlook in terms

of increasing the resource is favourable.

It is also worth noting that previous drill-testing by earlier operators in the greater region

identified additional areas of mineralization near the ICP deposits (see Figure 7.2). These

mineralized zones represent promising targets for future drilling.

25.2 MINING AND MINERAL RESERVES

Table 25.2 summarizes the mineral reserve estimate for the Idaho Cobalt Project.

Table 25.2

Mineral Reserve for the ICP at 0.25% Co Cut-off










230















25.3 ECONOMIC EVALUATION

The LOM base case project cash flow is presented in Table 25.3.

Table 25.3


Item LOM total

($000) $/t milled $/lb Cobalt

Cobalt Sales 846,837 231.26 26.66


Mining 196,692 53.71 6.19


Transport 5,199 1.42 0.16

CPF 149,121 40.72 4.69

G&A 37,309 10.19 1.17




EBITDA 686,415 187.45 21.61

Capital Costs 288,146 78.69 9.07


Tax 66,814 18.25 2.10


The base case cash flow evaluates to an IRR of 25.1% before taxes and 21.3% after tax. At a

discount rate of 7.5%, the net present value (NPV7.5) of the cash flow is $177 million before

tax and $136 million after tax.



assumptions.

231







232

26.0 RECOMMENDATIONS

26.1 GEOLOGY/MINERAL RESOURCES

In Micon’s view, the critical issues pertaining to the successful development of the ICP are

precision in predicting the grade and geometry of the various components of the deposit and

availability of additional resources to sustain the operations. To address these issues, Micon

makes the following recommendations:

While the block size of 6 ft. by 2 ft. by 5 ft. is an appropriate size for the narrow deposit

widths encountered at the Ram deposit and the envisaged SMU, the ability to estimate

grades and geometry with precision to this resolution requires a much closer drill

spacing. Accordingly, infill development drilling (for longhole stopes) and

development drifting (for cut and fill stopes) is recommended prior to commercial

underground mining production and before final stope design. The suggested infill drill

hole spacing is 30 to 35 ft.

Concurrently with infill development drilling, a drilling program to upgrade the

Inferred mineral resources should be initiated to increase the life of mine.

Additional exploration in the form of systematic step-out drilling should be conducted

following the main trend of mineralization in the north-westerly and south easterly

direction along strike and down dip.

A review and mineral resource update of the Sunshine and East Sunshine deposits is

recommended together with economic studies on trucking ore from these deposits to

the mill/concentrator facilities at the Ram deposit.

26.2 MINING

The following summarizes the recommendations observed during the preparation of the current

feasibility, even though “all drill-hole information, geotechnical data, and hydrological data

have been developed to a feasibility level” (PEA, 2015) in previous studies:

Backfill testing – Results from the 2017 pastefill material testing indicates that the

strength of the pastefill is not dependent of the type of cement or binder types but rather

on the water:cement ratio (P&C, 2017). It is recommended that additional material

testing to be carried out with increased binder addition to the current testing matrix

(50% cement: 50% slag) to potentially reduce the cement costs.

Backfill plant – Currently the backfill plant has only one silo for the cement storage. A

trade-off study to identify the technical and cost benefit for an additional binder storage

system will be advantageous to the project. Micon agrees with P&C that a re-evaluation

“of the paste delivery pipeline system design completed in 2008 to ensure that the

expected pumping pressures are appropriate given the change in tailings properties

from 2008 to 2017” (P&C, 2017) and to ensure these are compatible to the purchased

backfill equipment.

233

Geotechnical – Minefill indicated that the available geotechnical data is limited, and

additional data collection is warranted, including laboratory uniaxial compressive

strength testing. To date, there have been no geotechnical drillholes completed at the

Ram and no oriented core measurements collected from this deposit. Adit mapping

from the neighbouring Black Bird mine was carried out at adjacent adits near the Ram

portal, however the mapping of those adits encountered none of the principal units

expected in the Ram deposit (Minefill, 2006). Additional classification of the rock mass

in relation to it spatial location will assist in stope dimension and overall mine design.

The current mine design is very similar to the mine design in which the mine ventilation

study was performed. An updated mine ventilation study is recommended in the next

phase of engineering study before construction when a finalized mine design is

available.

Optimization – Additional optimization of the mine design, plan and especially the

production schedule can potentially improve the economics of the project.

26.3 PROCESSING – FUTURE TESTWORK

Copper Flotation

Additional tests are recommended to verify the copper scalping and cleaning flotation

performance using fresh samples that represent the relatively high Cu:Co ratio

mineralization planned to be mined and processed in the later years of the mine life.

Cobalt Solvent Extraction

Pilot plant cobalt solvent extraction testwork needs to be completed in order to provide

design details for the process. The objective of this additional testwork will be to

confirm extraction kinetics, determine optimum percent solids MgO vs. cobalt

recovery, confirm Co/Mg selectivity, determine strip liquor impurities and confirm the

overall circuit mass balance. The cobalt and zinc stripping conditions also need to be

confirmed.

Copper Solvent Extraction

The design of the copper solvent extraction circuit is based on the 2005 mini pilot plant

test program, the object of which was to produce cathode copper not copper sulphate

crystals. There may be a benefit of reviewing this circuit as the differences in the

optimal PLS specifications for these two applications (electrowinning vs

crystallization) could result in a simpler system and lower capital costs.

Crystallization

Although adequate bench scale testwork has been completed to provide a design for

the cobalt crystallizer circuit, additional detailed work needs to be completed to

establish the actual maximum recovery rate per pass and the critical impurity

concentration prior to the finalized design and procurement of the system. It is

234

recommended that extended continuous operations be performed using a high purity

feed electrolyte to produce additional cobalt sulphate crystals and investigate the

impact of impurity buildup of the product over a more prolonged period of operation.

A process to treat the bleed stream and recycle cobalt will also need to be developed.



Testwork needs to be completed using representative solution samples to provide

detailed design details of the magnesium sulphate crystallizer circuit.

Based on the recent copper crystallization testwork at SGS-L, it is recommended to

perform additional neutralization tests on both the feed solution and the copper raffinate

with the objective to (i) minimize cobalt and copper losses in the primary precipitate

stage and (ii) reduce the copper concentration in the feed to cobalt recovery, without

losing cobalt to the copper precipitate. This work should also include an evaluation of

a two stage precipitation process at two target pH levels for both processes.

Gold Recovery Circuit

Additional testwork is required to optimize the elemental sulphur flotation and the

cyanide leaching circuit circuits. Testwork also needs to be completed in order to model

the CIL circuit and gold/silver carbon loading as well as the cyanide destruction circuit.

CPF Pilot Plant

Much of the CPF processing circuits have been designed using batch tests or continuous

pilot tests using synthetic solutions. It is therefore recommended that the complete CPF

process be tested using a continuous pilot plant using composite samples of flotation

concentrate.

During the pilot plant testwork program it is suggested that solid/liquid separation and

washing of precipitates should be evaluated using pressure filtration and/or

centrifuging to develop an industrially robust methodology for removing the

precipitates produced within the process flowsheet.

Process Modelling and Simulation

As part of the feasibility study process engineering completed by SLI, a MetSim model

was developed for the CPF. This model needs to be developed to a higher level of detail

using the results from the additional testwork recommended above. The more robust

model will be available to stress test the final detailed design of the CPF.

HAZOP Studies

During the detailed design phase it is important to complete a hazard and operability

study (HAZOP) in order to identify and evaluate potential risks to personnel or

equipment so that the design can mitigate these risks.

235

27.0 DATE AND SIGNATURE PAGE

The effective date of this technical report is September 27th, 2017.

“Barnard Foo” {signed and sealed}

Barnard Foo, P.Eng., MBA

Date of signature: November 10th, 2017.

“Richard Gowans” {signed and sealed}

Richard Gowans, B.Sc., P.Eng.


“Christopher Jacobs” {signed and sealed}

Christopher Jacobs, CEng MIMMM


“David Makepeace” {signed and sealed}

David Makepeace, M.Eng., P.Eng.


“Charley Murahwi” {signed and sealed}

Charley Murahwi, M.Sc., P.Geo., FAusIMM


“Jane Spooner” {signed and sealed}



236

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SEDAR

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Samuel Engineering Inc., 2015: Preliminary Economic Assessment NI 43-101 Technical

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Anderson, Corby G. 2000a: Metallurgical Testing of RAM Ore, Consultant's report for

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Anderson, Corby G. 2000b: Task V Technical Assistance for Determination of Smelter

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237

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Project, Lemhi County, Idaho, unpublished report by Mine Development Associates for

Formation Capital Corporation.

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238

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Lemhi County, Idaho, USA, unpublished report by Mine Development Associates for

Formation Capital Corp., 140 p.

Slack, J.F., 2006: High REE and Y concentrations in Co-Cu-Au ores of the Blackbird district,

Idaho, Econ.Geol. 101,275-280.

Tysdal, R. G., 2000, Revision of Middle Proterozoic Yellowjacket Formation, Central Idaho,

USGS Professional Paper 1601-A, p. 1A-13A.

von Schwerin, M., February, 2015, Cobalt Market Review report unpublished report to

Formation Capital by Skybeco Inc.

Mining

Bieniwaski, Z. T., 1993 “Classification of rock masses for engineering: The RMR systems and

future trends”. Comprehensive Rock Engineering, (ed. Hudson), Oxford: Pergamon, p 553-

573.

Mitchell R.J., Olsen R.S., Smith J.D., 1982. Model studies on cemented tailings used in mine

backfill. Can. Geotech. J., Vol.19, No.1, pp. 14-28.

Minefill Services, Inc., 2006 “Underground geotechnical design parameters Ram/Sunshine

deposit – Idaho Cobalt Project”. January 10, 2006.

Minefill Services, Inc., 2006b “Idaho Cobalt Project – Updated ‘Preliminary’ Ground Support

Recommendations”, MineFill Services, Inc., January 30, 2006.

Nickson, S. D., 1992 “Cable support guidelines for underground hard rock mine operations.

M.A.Sc Thesis, Dept. Mining and Mineral Processing, University of British Columbia”.

Mine Development Associates (MDA), 2006, “Reserve Estimate and Mine Plan Idaho Cobalt

Project Feasibility Study, Lemhi County, Idaho, USA”, Mine Development Associates,

December 05, 2006.

Ouchi, A., Pakalnis, R., Brady, T., 2009 “Weak Rock Mass Span Design – Best Practices”,

ROCKENG09: Proceedings of the 3rd CANUS Rock Mechanics Symposium, Toronto, May

2009.

Paterson & Cooke, 2012 “Piston Pump Requirements, Reference: ICD-4021 R02 Rev B”, 15

May 2012.

239

Paterson & Cooke, 2011a “Idaho Cobalt Paste Backfill System: High Level Control

Philosophy, Reference: IDC-4021 R03 Rev A”, 17 May 2011.

Paterson & Cooke, 2011b “Idaho Cobalt Paste Backfill Design Criteria Document. Reference:

IDC-4021 R01 Rev B”, April 2011.

Paterson & Cooke, 2017 “eCobalt Idaho Cobalt Paste Testing: Test Work Report. Reference:

32-0228-00-TW-REP-0001 Rev C”, September 19, 2017.

Potvin, Y., 1998 “Empirical open stope design in Canada”, PhD Thesis, Dept. Mining and

Mineral Processing, University of British Columbia”.

FCC, Personal communications from eCobalt Solutions Inc. and Formation Capital Corp.,

2016-2017.

Metallurgy and Process Design

Anderson C.G., Nordwick S.M., Society for Mining, Metallurgy and Exploration, Inc “Novel

Precious Metal Processes”, March, 1996.

Anderson C.G., CAMP, “The Mineral Processing and Industrial Nitrogen Species Catalyzed

Pressure Leach Plant Treatment of Formation Capital Corporation’s Colbaltite and

Chalcopyrite Concentrates”, not dated.

Center for Advanced Mineral & Metallurgical Processing (CAMP), “MLA Characterization

of Autoclave Residue”, November 22, 2011.

Cytec Solvay Group, “Summary of Lab Work Conducted for Idaho Cobalt Project through

2015”, not dated.

FLSmidth Minerals, “Report on Testing for Minefill Services Inc., Idaho Cobalt Project,

Sedimentation, Rheology and Filtration Tests on Cobalt Tailings”, December, 2007.

GE Water & Process Technologies, “Final Report, Cobalt Sulfate Testing for Formation

Metals Inc.”, December 7, 2015.

Hazen Research, Inc., “Laboratory Program for the Flotation of Copper and Cobalt Minerals

for the Idaho Cobalt Project”, August 28, 2015.

Hazen Research, Inc., “Laboratory Program to Generate Process Data for the Idaho Cobalt

Production Facility”, September 25, 2015.

Hydromet (Pty) Ltd., “Overview of Test Work Conducted for Formation Capital Corporation”,

May, 2007.

240

Hydromet (Pty) Ltd., “Report on Batch Leach & Metathesis Tests Conducted for Formation

Capital Corporation”, May, 2007.

Mintek, “Idaho Mini Plant”, 26 July, 2005.

Mintek, “Laboratory Investigation into Cobalt Precipitation for the Idaho Project”, May 24,

2007.

Mintek, “Idaho Cobalt Pilot Campaign 2007”, 10 December, 2008.

Pocock Industrial, Inc, “Flocculant Screening, Gravity Sedimentation, Pulp Rheology,

Vacuum and Pressure Filtration, Studies”, conducted for Formation Chemical Corporation,

Idaho Cobalt Project, April 2005.

SGS Lakefield Research Limited, “A Mineralogical Description of Six Crushed Samples from

Formation Capital Corporation”, May 05, 2005.

SGS Lakefield Research Limited, “An Investigation into a Cobalt Copper Concentrate”,

prepared for Formation Capital Corporation U.S., April 05, 2005.

SGS Lakefield Research Limited, “An Investigation into the Recovery of Copper and Cobalt

by Laboratory Flotation”, prepared for Formation Capital Corporation U.S., May 05, 2005.

SGS Lakefield Research Limited, “An Investigation into the Recovery of Cobalt and Copper

from a Co-Cu-As Flotation Concentrate submitted by Formation Chemicals”, June 3, 2005.

SGS Lakefield Research Limited, “The Recovery of Copper and Cobalt by Laboratory

Flotation”, prepared for Formation Capital Corporation U.S., May 5, 2005.

SGS Minerals Services, “An Investigation into Continuous Cobalt Crystallization Testing -

Idaho Cobalt”, prepared for eCobalt Solutions Inc., October 26, 2017.

SGS Minerals Services, “An Investigation into Copper Crystallization Testing - Idaho Cobalt”,

prepared for eCobalt Solutions Inc., October 17, 2017.

SGS Minerals Services, “Bench Scale Hydrometallurgical Testing for Feasibility Study- Idaho

Cobalt”, prepared for eCobalt Solutions Inc., April 26, 2017.

SGS Minerals Services, “Bench Scale Flotation Testing for Feasibility Study- Idaho Cobalt”,

prepared for eCobalt Solutions Inc., May 12, 2017.

Telesto Solutions Inc., “Tailings and Waste Rock Storage Facility Design Report for the Idaho

Cobalt Project”, February, 2011.

241

Environmental

Bruner, D. and Thurman, G., 2016. Water Treatment System Options Assessment for eCobalt

Solutions, Inc. (formerly ICP) – Lemhi County, Idaho, 35 pp.

EPA, 2009. Record of Decision, Idaho Cobalt Project, 26 pp.

Formation Capital Corporation U.S., 2015. 2014 Idaho Cobalt Project Monitoring Summary,

281 pp.

SCNF, 2007. Draft Environmental Impact Statement, Idaho Cobalt Project, Salmon-Cobalt

Ranger District, Salmon-Challis National Forest, Lemhi County, Idaho; United States

Department of Agriculture, Forest Service, 370 pp.

SCNF, 2008. Final Environmental Impact Statement, Idaho Cobalt Project, Salmon-Cobalt

Ranger District, Salmon-Challis National Forest, Lemhi County, Idaho; United States

Department of Agriculture, Forest Service, 365 pp, 7 chapters plus appendices, 464 pp.

SCNF, 2009. Record of Decision, Idaho Cobalt Project, Salmon-Cobalt Ranger District,

Salmon-Challis National Forest, Lemhi County, Idaho; United States Department of

Agriculture, Forest Service, 60 pp.

Telesto, 2005. Environmental Response to Mining at the Idaho Cobalt Project.

Marketing

CRU Consulting, Market study for the Idaho Cobalt Project (ICP), A final report for eCobalt

Solutions Inc., dated September 2017.

CRU Consulting, 2017, Market study for the Idaho Cobalt Project (ICP), A final report for

eCobalt Solutions Inc., Updated Executive Summary, dated October, 2017.

USGS, 2016, Mineral Commodity Summaries, Cobalt, January, 2017.

242

29.0 CERTIFICATES

243

CERTIFICATE OF AUTHOR

Barnard Foo

As co-author of this report entitled “Feasibility Study for the Idaho Cobalt Project, Idaho, U.S.A”, with an

effective date of September 27th, 2017 (the “Technical Report”), I, Barnard Foo, do hereby certify that:

1. I am employed as a senior mining engineer by, and carried out this assignment for:

Micon International Limited,

Suite 900 - 390 Bay Street, Toronto, ON, M5H 2Y2

Tel. (416) 362-5135 Email: [email protected]

2. I hold the following academic qualifications:

Laurentian University, B.Eng., Mining Engineering 1998

University of British Columbia, M. Eng., Rock Mechanics 2007

University of Northern British Columbia, Executive MBA 2010

3. I am a registered Professional Engineer with the Professional Engineers of Ontario (Membership #

100052925);

4. I have worked as a mining engineer in the minerals industry for 19 years;

5. I am familiar with NI 43-101 and, by reason of education, experience and professional registration; I

fulfill the requirements of a Qualified Person as defined in NI 43-101. My work experience includes 4

years as an mining engineer in cassiterite, base and precious metal deposits, 5 years in underground and

open pit geotechnical engineering, and 10 years consulting in mine design and mining project evaluation

for the mineral industry;

6. I visited the Idaho Cobalt Project on July 13-14, 2016;

7. I am responsible for Sections 15, 16, 21.1.1, 21.2.1 and the portions of Sections 1, 25 and 26 summarized

therefrom, of the Technical Report.

8. I am independent of eCobalt Solutions Inc. and related entities, as defined in Section 1.5 of NI 43-101;

9. I have had no previous involvement with the Property except for the purposes of an independent due

diligence carried out on behalf of unrelated financial institutions in 2010-2012.

10. I have read NI 43-101 and the portions of this report for which I am responsible for which have been

prepared in compliance with the instrument;

11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this

Technical Report for which I am responsible contain all scientific and technical information that is

required to be disclosed to make this report not misleading.

Dated this November 10th, 2017

“Barnard Foo” {signed and sealed}

Barnard Foo, M.Eng., P.Eng., MBA.

Senior Mining Engineer

244


Richard Gowans


effective date of September 27th, 2017, (the “Technical Report”), I, Richard Gowans do hereby certify that:

1. I am employed by, and carried out this assignment for, Micon International Limited, Suite 900, 390 Bay

Street, Toronto, Ontario M5H 2Y2.

tel. (416) 362-5135, e-mail [email protected].


B.Sc. (Hons) Minerals Engineering, The University of Birmingham, U.K. 1980.

3. I am a registered Professional Engineer of Ontario (membership number 90529389); as well, I am a

member in good standing of the Canadian Institute of Mining, Metallurgy and Petroleum.

4. I am familiar with NI 43-101 and by reason of education, experience and professional registration, fulfill

the requirements of a Qualified Person as defined in NI 43-101. My work experience includes over 30

years of the management of technical studies, management of numerous metallurgical testwork programs,

design of metallurgical processing plants and due diligence reviews of a number of cobalt and copper

projects.

5. I have read NI 43-101 and this Technical Report has been prepared in compliance with the instrument.

6. I visited the mine and mill site near Salmon, Idaho on February 7, 2012.



8. I am independent of eCobalt Solutions Inc. and related entities as defined in Section 1.5 of NI 43-101.

9. I am responsible for Sections 13, 17 and 18 and the portions of Sections 1, 25 and 26 summarized

therefrom, of this Technical Report.

10. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report

contains all scientific and technical information that is required to be disclosed to make this technical

report not misleading.


“Richard Gowans” {signed and sealed as of the report date}

Richard Gowans, P.Eng.

mailto:[email protected]

245


Christopher Jacobs


effective date of September 27th, 2017, (the “Technical Report”), I, Christopher Jacobs, do hereby certify that:

1. I am employed by, and carried out this assignment for, Micon International Limited, Suite 900 - 390 Bay

Street, Toronto, Ontario M5H 2Y2. tel. (416) 362-5135, email:[email protected].


B.Sc. (Hons) Geochemistry, University of Reading, 1980;

M.B.A., Gordon Institute of Business Science, University of Pretoria, 2004.

3. I am a Chartered Engineer registered with the Engineering Council of the U.K.

(registration number 369178).

4. Also, I am a professional member in good standing of: The Institute of Materials, Minerals and Mining;

and The Canadian Institute of Mining, Metallurgy and Petroleum (Member).

5. I have worked in the minerals industry for more than 35 years; my work experience includes 10 years as

an exploration and mining geologist on gold, platinum, copper/nickel and chromite deposits; 10 years as

a technical/operations manager in both open-pit and underground mines; 3 years as strategic (mine)

planning manager and the remainder as an independent consultant when I have worked on a variety of

deposits including cobalt, copper and gold.

6. I do, by reason of education, experience and professional registration, fulfill the requirements of a

Qualified Person as defined in NI 43-101.

7. I visited the Idaho Cobalt Project mine and mill site near Salmon, Idaho on July 13-14, 2016.

8. I am responsible for Section 21 (other than 21.1.1 and 21.2.1), Section 22, 24 and the portions of Sections

1, 25 and 26 summarized therefrom, of this Technical Report.

9. I am independent of eCobalt Solutions Inc. and related entities, as defined in Section 1.5 of NI 43-101.



11. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared in

compliance with the instrument.

12. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this

Technical Report for which I am responsible contain all scientific and technical information that is

required to be disclosed to make this report not misleading.


“Christopher Jacobs” {signed and sealed}

Christopher Jacobs, CEng, MIMMM


246


David K. Makepeace


effective date of September 27th, 2017, (the “Technical Report”), I, David Makepeace, do hereby certify that:

1. I am employed by and carried out this assignment for:

Micon International Limited, Suite 205 - 700 West Pender Street, Vancouver, British Columbia, V6C

1G8, Canada. Telephone : (604) 647-6463, Fax : (604) 647-6455.


Bachelor of Applied Science - Geological Engineering, Queen’s University at Kingston, Ontario,

1976,

Master of Engineering - Environmental Engineering, University of Alberta, 1994.

3. I am a registered member of the:

Association of Professional Engineers and Geoscientists of British Columbia, licence 14912.

Association of Professional Engineers, Geologists and Geophysicists of Alberta, licence 29367.

4. I have worked as a geological engineer for a total of 34 years since my graduation from university.

5. I have read the definition of Qualified Person set out in National Instrument (NI) 43-101 and certify that

by reason of my education, affiliation with professional associations as defined in NI 43-101 and past

relevant work experience, I fulfill the requirements to be a Qualified Person for this report for the purposes

of NI 43-101. My relevant experience includes several years as an environmental engineer on several

projects in Canada, USA, Mexico, DRC and Zambia.

6. I am the author of section 20 and any summaries therefrom in sections 1, 25 and 26 of the Technical

Report.

7. I visited the property on July 12 to July 14, 2016 and inspected the mine and mill site (ICP) west of

Salmon Idaho.

8. I have had no prior involvement with the property which is the subject of this Technical Report.

9. As of the date of this certificate, I am not aware of any material fact or material change with respect to

the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to

disclose which makes the Technical Report misleading.

10. I am independent of the issuer applying all the tests in section 1.5 of NI 43-101 other than providing

consulting services.

11. I have read NI 43-101, Companion Policy 43-101CP and Form 43-101F1, and the Technical Report has

been prepared in compliance with that instrument, companion policy and form.


(signed) “David K. Makepeace” (Sealed)

_____________________________________ __________________________

David K. Makepeace, M.Eng., P.Eng. Professional Engineering Stamp

Senior Geologist - Environmental Engineer

Micon International Limited

247


Charley Murahwi


effective date of September 27th, 2017, (the “Technical Report”), I, Charley Murahwi, do hereby certify that:

1. I am employed as a Senior Geologist by, and carried out this assignment for, Micon International

Limited, Suite 900, 390 Bay Street, Toronto, Ontario M5H 2Y2, telephone 416 362 5135, fax 416 362

5763, e-mail [email protected].


B.Sc. (Geology) University of Rhodesia, Zimbabwe, 1979;

Diplome d΄Ingénieur Expert en Techniques Minières, Nancy, France, 1987;

M.Sc. (Economic Geology), Rhodes University, South Africa, 1996.

3. I am a registered Professional Geoscientist in Ontario (membership # 1618) and in PEGNL

(membership # 05662), a registered Professional Natural Scientist with the South African Council for

Natural Scientific Professions (membership # 400133/09) and am a Fellow of the Australasian Institute

of Mining & Metallurgy (FAusIMM) (membership number 300395).

4. I have worked as a mining and exploration geologist in the minerals industry for over 30 years.


Qualified Person as defined in NI 43-101. My work experience includes 18 years on gold, silver,

copper, cobalt, tin and tantalite projects (on and off mine), and 14 years on Cr-Ni-Cu-PGE deposits in

layered intrusions/komatiitic environments.

6. I visited the Idaho Cobalt Project on 9 December 2010 and from 13 to 14 July 2016.



8. As of the date of this certificate to the best of my knowledge, information and belief, the Technical

Report contains all scientific and technical information that is required to be disclosed to make this

report not misleading;

9. I am independent of the parties involved in the Idaho Cobalt Project as described in Section 1.5 of NI

43-101.

10. I have read NI 43-101 and the portions of this Technical Report for which I am responsible have been

prepared in compliance with this Instrument.

11. I am responsible for Sections 1.1 to 1.6, 1.15.1, 2 to 12, 14, 23, 25.1, 26.1 and 28.1 of the Technical

Report.


“Charley Murahwi” {signed and sealed}

Charley Murahwi, M.Sc., P. Geo. FAusIMM


248



As a co-author of this report entitled “NI 43-101 F1 Technical Report Feasibility Study for the Idaho Cobalt

Project Idaho, USA”, dated 10 November, 2017, I, Jane Spooner, P.Geo., do hereby certify that:

1. I am employed by, and carried out this assignment for

Micon International Limited

Suite 900, 390 Bay Street

Toronto, Ontario

M5H 2Y2

tel. (416) 362-5135 fax (416) 362-5763

e-mail: [email protected]


B.Sc. (Hons) Geology, University of Manchester, U.K. 1972

M.Sc. Environmental Resources, University of Salford, U.K. 1973

3. I am a member of the Association of Professional Geoscientists of Ontario (membership number

0990); as well, I am a member in good standing of the Canadian Institute of Mining, Metallurgy and

Petroleum.

4. I have worked as a specialist in mineral market analysis for over 30 years.


Qualified Person as defined in NI 43-101. My work experience includes the analysis of markets for

base and precious metals, industrial and specialty minerals, coal and uranium.

6. I have not visited the project site.

7. I am responsible for Section 19.0 and Section 1.11 of this report entitled “NI 43-101 F1 Technical

Report Feasibility Study for the Idaho Cobalt Project Idaho, USA”, dated 10 November, 2017.

8. I am independent of eCobalt Solutions Inc. and related entities, as described in Section 1.5 of NI 43-

101.

9. I have had no prior involvement with the mineral property in question except for the purposes of an

independent due diligence carried out on behalf of unrelated financial institutions in 2010-2012.

10. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared

in compliance with the instrument.

11. As of the date of this certificate, to the best of my knowledge, information and belief, the sections of

this Technical Report for which I am responsible contain all scientific and technical information that

is required to be disclosed to make this report not misleading.

Signing Date:

“Jane Spooner” {signed and sealed}



formation capital corporation, u.s. ecobalt solutions … · suite 900 - 390 bay street, toronto...

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