ni 43-101 updated prefeasibility technical report bovill

NI 43-101 Updated Prefeasibility Technical Report Bovill Kaolin Project Latah County, Idaho Effective Date: April 20, 2014 Report Date: June 26, 2014 Report Prepared for

880 – 580 Hornby Street Vancouver, B.C., V6C 3B6 Canada Report Prepared by

SRK Consulting (U.S.), Inc. 7175 West Jefferson Avenue, Suite 3000 Lakewood, CO 80235 SRK Project Number: 165800.080

Reviewed by: Bret Swanson, BEng Mining, MAusIMM, MMSAQP

Signed by QP(s): Bart Stryhas, PhD, CPG Harry Gatley, PE David S. Hallman, PE Manuel Rauhut, PE Valerie Obie, BS Mining, RM-SME

SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page i

Table of Contents 1 Summary ....................................................................................................................... 1

1.1 Property Description and Ownership .............................................................................................. 1

1.2 Geology and Mineralization ............................................................................................................ 1

1.3 Status of Exploration, Development and Operations ....................................................................... 1

1.4 Mineral Processing and Metallurgical Testing ................................................................................. 1

1.5 Mineral Resource Estimate ............................................................................................................ 3

1.6 Mineral Reserve Estimate .............................................................................................................. 5

1.7 Mining Methods.............................................................................................................................. 5

1.8 Project Infrastructure ...................................................................................................................... 6

1.9 Environmental Studies and Permitting ............................................................................................ 6

1.10 Capital and Operating Costs .......................................................................................................... 7 1.11 Economic Analysis ......................................................................................................................... 7

1.11.1 Sensitivity Analysis ............................................................................................................. 8

1.12 Conclusions and Recommendations .............................................................................................. 8

2 Introduction ................................................................................................................ 10

2.1 Terms of Reference and Purpose of the Report ............................................................................ 10

2.2 Qualifications of Consultants (SRK) .............................................................................................. 10 2.3 Details of Inspection ..................................................................................................................... 11

2.4 Sources of Information ................................................................................................................. 12

2.5 Effective Date .............................................................................................................................. 12

2.6 Units of Measure .......................................................................................................................... 12

3 Reliance on Other Experts ........................................................................................ 13

4 Property Description and Location .......................................................................... 14

4.1 Property Location ......................................................................................................................... 14

4.2 Mineral Titles ............................................................................................................................... 14

4.3 Royalties, Agreements and Encumbrances .................................................................................. 16

4.4 Environmental Liabilities and Permitting ....................................................................................... 17

4.4.1 Environmental Liabilities ................................................................................................... 17

4.4.2 Required Permits and Status ............................................................................................ 17

4.5 Other Significant Factors and Risks .............................................................................................. 18

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ........ 19

5.1 Topography, Elevation and Vegetation ......................................................................................... 19

5.2 Accessibility and Transportation to the Property ........................................................................... 19

5.2.1 Proximity to Population Center and Transport ................................................................... 19

5.3 Climate and Length of Operating Season ..................................................................................... 20

VO/MLM Bovill_Kaolin_NI43-101_PFS_Update_Report_165800.080_015_MLM.docx June 26, 2014

SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page ii

5.4 Sufficiency of Surface Rights ........................................................................................................ 20

5.5 Infrastructure Availability and Sources .......................................................................................... 20

6 History ......................................................................................................................... 21

6.1 Prior Ownership, Exploration, Resource Estimates and Production............................................... 21

7 Geological Setting and Mineralization ..................................................................... 23

7.1 Regional Geology ........................................................................................................................ 23

7.2 Local Geology .............................................................................................................................. 24

7.3 Significant Mineralized Zones ....................................................................................................... 27

7.3.1 Feldspars.......................................................................................................................... 27

7.3.2 Quartz .............................................................................................................................. 28 7.3.3 Clay Minerals .................................................................................................................... 28

8 Deposit Type .............................................................................................................. 30

8.1 Mineral Deposit ............................................................................................................................ 30

9 Exploration ................................................................................................................. 31

9.1 Procedures and Parameters of Surveys and Investigations .......................................................... 31

9.2 Sampling Methods and Sample Quality ........................................................................................ 31

9.3 Significant Results and Interpretation ........................................................................................... 31

10 Drilling ......................................................................................................................... 32

10.1 Type and Extent ........................................................................................................................... 32

10.2 Interpretation and Results ............................................................................................................ 34

11 Sample Preparation, Analysis and Security ............................................................ 35

11.1 Sampling and Preparation Methods .............................................................................................. 35

11.2 Laboratories ................................................................................................................................. 36

11.3 Analysis ....................................................................................................................................... 36

11.4 Security Measures ....................................................................................................................... 38 11.5 QA/QC Procedures and Results ................................................................................................... 38

11.6 Opinion on Adequacy ................................................................................................................... 44

12 Data Verification ......................................................................................................... 45

12.1 Procedures .................................................................................................................................. 45

12.2 Limitations ................................................................................................................................... 46

12.3 Opinion on Data Adequacy........................................................................................................... 46

13 Mineral Processing and Metallurgical Testing ........................................................ 47

13.1 Testing and Procedures ............................................................................................................... 47

13.1.1 Initial Testing of Primary Clay ............................................................................................ 47

13.1.2 Mechanical Testing of Hydrocyclones ............................................................................... 52

13.1.3 Pilot Plant Testing of Primary Clay .................................................................................... 52


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page iii

13.1.4 Pilot Plant Testing of WBL K-Spar Tailings ........................................................................ 53

13.1.5 Bench-scale Testing of K-Spar Flotation Tailings for Quartz .............................................. 55

13.1.6 Pilot Plant Testing of K-Spar Flotation Tailings for Quartz .................................................. 56

13.1.7 Repeat Bench-Scale Testing of K-Spar Flotation Tailings for Quartz ................................. 56

13.2 Relevant Results .......................................................................................................................... 57

13.2.1 Results and Conclusions Based on Primary Clay Tests ..................................................... 57

13.2.2 Results and Conclusions Based on Mechanical Testing of Hydrocyclones ........................ 60

13.2.3 Results and Conclusions Based on Pilot Plant Testing of Primary Clay ............................. 61 13.2.4 Results and Conclusions for Testing of WBL K-spar Tailings ............................................. 70

13.2.5 Results and Conclusions for Bench-scale Testing of K-spar Flotation for Quartz ............... 72

13.2.6 Results and Conclusions for Pilot Plant Testing of K-Spar Flotation Tailings for Quartz ..... 73

13.2.7 Results and Conclusions for Repeat Bench-scale Testing of K-Spar Flotation Tailings for Quartz .............................................................................................................................. 73

13.3 Recovery Estimate Assumptions .................................................................................................. 74

13.4 Sample Representativeness ......................................................................................................... 74

13.5 Significant Factors ....................................................................................................................... 74

14 Mineral Resource Estimate ....................................................................................... 76

14.1 Geology of the Resource Estimation ............................................................................................ 76

14.2 Drillhole Database ........................................................................................................................ 76

14.3 Capping and Compositing ............................................................................................................ 76

14.4 Variogram Analysis ...................................................................................................................... 77

14.5 Density ........................................................................................................................................ 77

14.6 Block Model and Topography ....................................................................................................... 78

14.7 Resource Modeling ...................................................................................................................... 78

14.8 Model Validation .......................................................................................................................... 85

14.9 Resource Classification ................................................................................................................ 89 14.10 Mineral Resource Statement ........................................................................................................ 89

14.11 Relevant Factors .......................................................................................................................... 91

15 Mineral Reserve Estimate .......................................................................................... 92

15.1 Conversion Assumptions, Parameters and Methods ..................................................................... 92

15.2 Relevant Factors .......................................................................................................................... 92

16 Mine Design ................................................................................................................ 93

16.1 Geotechnical Conditions .............................................................................................................. 93

16.2 Pit Optimization ............................................................................................................................ 93

16.2.1 Block Models .................................................................................................................... 94

16.2.2 Parameters ....................................................................................................................... 94

16.2.3 Results ............................................................................................................................. 97


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page iv

16.3 Pit Design .................................................................................................................................. 102

16.3.1 Parameters ..................................................................................................................... 102

16.3.2 Results ........................................................................................................................... 102

16.3.3 Pit Design versus Optimized Shell................................................................................... 107

16.3.4 Waste Dumps ................................................................................................................. 111

16.3.5 Pit and Waste Dump Layout............................................................................................ 114

16.4 Mine Schedule ........................................................................................................................... 115

16.5 Mine Operations......................................................................................................................... 119 16.5.1 Mining Operations ........................................................................................................... 119

16.5.2 Mining Operations Parameters ........................................................................................ 119

16.5.3 Drilling and Blasting ........................................................................................................ 120

16.5.4 Loading (SRK Estimation) ............................................................................................... 120

16.5.5 Haulage (SRK Estimation) .............................................................................................. 121

16.5.6 Mine Contractor Labor (SRK Estimation) ......................................................................... 121

16.5.7 Mining Contractor Infrastructure (SRK Estimation) .......................................................... 122

16.5.8 I-Minerals (Owner) Mining Functions ............................................................................... 122

17 Recovery Methods ................................................................................................... 123

17.1 Operation Results ...................................................................................................................... 123

17.1.1 Yield/Recovery Results ................................................................................................... 125

17.1.2 Material Balance Criteria ................................................................................................. 125

17.2 Processing Methods ................................................................................................................... 125

17.2.1 General Considerations .................................................................................................. 128

17.2.2 Unit Operations ............................................................................................................... 128

17.2.3 Additional Tests & Recommendations ............................................................................. 130

17.3 Flowsheet .................................................................................................................................. 130

17.4 Plant Design and Equipment Characteristics .............................................................................. 130 17.4.1 Ore Preparation .............................................................................................................. 130

17.4.2 Clay/Feldspathic Sands Classification ............................................................................. 130

17.4.3 Degritting ........................................................................................................................ 130

17.4.4 Halloysite/Kaolin Clay Separation ................................................................................... 131

17.4.5 Halloysite Clay Product ................................................................................................... 131

17.4.6 Kaolin Clay/Metakaolin Products ..................................................................................... 131

17.4.7 Feldspathic Sand Grinding, Sizing, and Attrition Scrubbing ............................................. 132

17.4.8 Iron and K-spar Flotation................................................................................................. 132

17.4.9 K-spar Product Processing ............................................................................................. 132

17.4.10 Quartz Regrinding, Sizing, and Attrition Scrubbing ...................................................... 132

17.4.11 Quartz Flotation .......................................................................................................... 132


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page v

17.4.12 Quartz Products Processing ....................................................................................... 133

17.5 Consumable Requirements ........................................................................................................ 133

18 Project Infrastructure ............................................................................................... 134

18.1 Site Access Road ....................................................................................................................... 134

18.2 On-Site Services ........................................................................................................................ 134

18.3 Energy Supply ........................................................................................................................... 134

18.3.1 Electric Power ................................................................................................................. 134

18.3.2 Natural Gas .................................................................................................................... 134

18.4 Water ......................................................................................................................................... 135

18.5 Tailings Storage Facility (TSF) ................................................................................................... 137

18.6 Product Transportation ............................................................................................................... 137

19 Market Studies and Contracts ................................................................................ 138

19.1 Product Markets ......................................................................................................................... 138

19.1.1 Halloysite ........................................................................................................................ 138

19.1.2 Kaolin ............................................................................................................................. 138

19.1.3 Metakaolin ...................................................................................................................... 139

19.1.4 Feldspar ......................................................................................................................... 139

19.1.5 Quartz ............................................................................................................................ 139

19.2 Commodity Price Projections ..................................................................................................... 140

19.3 Contracts and Status .................................................................................................................. 140

19.3.1 Terms ............................................................................................................................. 140

20 Environmental Studies, Permitting and Social or Community Impact ................ 141

20.1 Environmental Setting ................................................................................................................ 141

20.2 Environmental Studies and Permitting ........................................................................................ 141

20.2.1 Required Permits and Status .......................................................................................... 144

20.2.2 Post-Performance or Reclamations Bonds ...................................................................... 146

20.3 Social and Community ............................................................................................................... 146

20.4 Mine Closure .............................................................................................................................. 146

21 Capital and Operating Costs ................................................................................... 147

21.1 Capital Cost Estimates ............................................................................................................... 147

21.1.1 Mining............................................................................................................................. 147

21.1.2 Process & Infrastructure.................................................................................................. 148

21.1.3 Tailings Storage Facility (TSF) ........................................................................................ 149

21.1.4 Owner’s Costs ................................................................................................................ 151

21.2 Operating Cost Estimates .......................................................................................................... 151 21.2.1 Mining............................................................................................................................. 152

21.2.2 Processing ...................................................................................................................... 153


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page vi

21.2.3 G&A ............................................................................................................................... 154

22 Economic Analysis .................................................................................................. 155

22.1 Principal Assumptions ................................................................................................................ 155 22.2 Cashflow Forecasts and Annual Production Forecasts ............................................................... 155

22.3 Taxes, Royalties and Other Interests .......................................................................................... 157

22.3.1 Federal Income Tax ........................................................................................................ 157

22.3.2 State Taxes .................................................................................................................... 158

22.3.3 Royalty ........................................................................................................................... 158

22.4 Sensitivity Analysis..................................................................................................................... 158

23 Adjacent Properties ................................................................................................. 159

24 Other Relevant Data and Information ..................................................................... 159

25 Interpretation and Conclusions .............................................................................. 159

25.1 Results ...................................................................................................................................... 159

25.2 Significant Risks and Uncertainties ............................................................................................. 160

25.2.1 Exploration ..................................................................................................................... 160

25.2.2 Mineral Resource Estimate ............................................................................................. 160

25.2.3 Mineral Reserve Estimate ............................................................................................... 160

25.2.4 Metallurgy and Processing .............................................................................................. 160

25.2.5 Tailings Disposal............................................................................................................. 160

25.2.6 Environmental ................................................................................................................. 161

25.2.7 Projected Economic Outcomes ....................................................................................... 161

26 Recommendations ................................................................................................... 162

26.1 Exploration ................................................................................................................................. 162

26.2 Mining Study .............................................................................................................................. 162

26.3 Environmental Permitting and Supporting Studies ...................................................................... 162

26.4 Tailings Characterization and Design ......................................................................................... 163

26.5 Metallurgy and Processing ......................................................................................................... 163

26.6 Recommended Work Programs and Costs ................................................................................. 163

26.6.1 Costs .............................................................................................................................. 164

27 References ................................................................................................................ 165

28 Glossary .................................................................................................................... 168

28.1 Mineral Resources ..................................................................................................................... 168 28.2 Mineral Reserves ....................................................................................................................... 168

28.3 Definition of Terms ..................................................................................................................... 169

28.4 Abbreviations ............................................................................................................................. 170


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page vii

List of Tables Table 1.5.1: Measured and Indicated Mineral Resource Statement (as of April 20, 2014)* ............................. 4

Table 1.6.1: Mineral Reserve Statement, (as of June 14, 2014) ..................................................................... 5

Table 1.11.1: LoM Capital Cost Summary (US$000’s) ................................................................................... 7

Table 1.11.2: LoM Operating Cost Summary ................................................................................................. 7 Table 1.12.1: Financial Model Results ........................................................................................................... 8

Table 1.12.1.1: Project NPV Sensitivities (US$ millions) ................................................................................ 8

Table 4.2.1: Mineral Leases ........................................................................................................................ 15

Table 7.3.2.1: Average Quartz Composition Calculated from Electron Microprobe Analyses ........................ 28

Table 11.3.1: Clay Code Assignment .......................................................................................................... 38

Table 13.2.1.1: Phase I Characterization Data ............................................................................................. 58

Table 13.2.3.1: Primary Clay Pilot Plant Trials Compilation Recovery Data (1) .............................................. 68

Table 13.2.4.1: Grade vs. Hydrofluoric Acid Dosage .................................................................................... 71

Table 13.2.4.2: Grade vs. Collector Amine Dosage ..................................................................................... 71

Table 13.2.4.3: Feldspar Middlings Evaluation ............................................................................................. 71

Table 13.2.5.1: Quartz Products Evaluation ................................................................................................. 72

Table 13.2.7.1: Quartz Products Evaluation ................................................................................................. 74 Table 14.3.1: Capping and Compositing Results ......................................................................................... 77

Table 14.5.1 Block Model Material Densities ............................................................................................... 77

Table 14.6.1: Block Model Limits WBL and Middle Ridge Areas .................................................................. 78

Table 14.6.2: Block Model Limits Kelly’ Hump Areas ................................................................................... 78

Table 14.7.1 Percentage of Model Blocks in Clay Shell ............................................................................... 80

Table 14.7.2 Resource Estimation Parameters ............................................................................................ 80

Table 14.8.1: Grade Estimation Performance Parameters ........................................................................... 86

Table 14.8.2: Statistical Model Validation .................................................................................................... 87

Table 14.8.3: Nearest Neighbor Model Validation ........................................................................................ 88

Table 14.10.1: Indicated Mineral Resource Statement, (as of 20 April 2014) ................................................ 90

Table 15.1: Mineral Reserve Statement, (as of June 14, 2014) .................................................................... 92 Table 16.1.1: Slope Angles from Geotechnical Report by STRATA (2012)................................................... 93

Table 16.2.2.1: Whittle™ Kelly’s Hump Block Model Import Parameters ...................................................... 95

Table 16.2.2.2: Whittle™ Middle Ridge Block Model Import Parameters ...................................................... 95

Table 16.2.2.3: Combined Recoveries for Kaolinite and Sand Minerals........................................................ 96

Table 16.2.2.4: Combined Selling Prices for Kaolinite and Sand Minerals .................................................... 96

Table 16.2.2.5: Whittle™ Economic Parameters .......................................................................................... 97

Table 16.2.3.1: Whittle™ Economic Pits Middle Ridge ................................................................................. 98

Table 16.2.3.2: Whittle™ Economic Pits Kelly’s Hump ................................................................................. 99

Table 16.3.1.1: Pit Design Parameters ...................................................................................................... 102


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page viii Table 16.3.2.1: Kelly’s Hump Phases Tonnages ........................................................................................ 107

Table 16.3.2.2: Kelly South Phases Tonnages .......................................................................................... 107

Table 16.3.2.3: Middle Ridge Phases Tonnages ........................................................................................ 107

Table 16.3.2.4: WBL Phases Tonnages .................................................................................................... 107

Table 16.3.3.1: Kelly’s Hump and Kelly South Pit Design versus Optimized Shells .................................... 107

Table 16.3.3.2: Middle Ridge and WBL Pit Design versus Optimized Shells .............................................. 108

Table 16.3.4.1: Waste Dump Design Parameters ...................................................................................... 111

Table 16.4.1: Mine Schedule ..................................................................................................................... 116 Table 16.5.2.1: SRK Estimate of Mining Equipment ................................................................................... 120

Table 16.5.4.1: SRK Estimate of Material Loading Statistics by Loading Unit Type .................................... 120

Table 18.5.1: Boyle TSF Design Criteria .................................................................................................... 137

Table 19.2.1: Commodity Prices ................................................................................................................ 140

Table 20.2.1.1: Required Environmental Studies by Permit ....................................................................... 145

Table 21.1.1: LoM Capital Cost Summary (US$000’s) ............................................................................... 147

Table 21.1.1.1: Mine Owner Cost Summary (US$000’s) ............................................................................ 148

Table 21.1.1.2: Support Equipment (US$000’s) ......................................................................................... 148

Table 21.1.2.1: LoM Process & Infrastructure Capital (US$000’s) .............................................................. 148

Table 21.1.2.2: Process Capital Cost (US$000’s) ...................................................................................... 149

Table 21.1.3.1: TSF LoM Summary Capital Cost (US$000’s) ..................................................................... 149 Table 21.1.3.2: TSF Capital Cost (US$000’s) ............................................................................................ 150

Table 21.1.4.1: Owner Costs (US$000’s)................................................................................................... 151

Table 21.2.1: LoM Operating Cost Summary ............................................................................................. 151

Table 21.2.1.1: Mine Operating Cost Summary ......................................................................................... 152

Table 21.2.1.2: Mine Contractor Cost Estimate .......................................................................................... 152

Table 21.2.1.3: Mine G&A (Owner) ............................................................................................................ 152

Table 21.2.1.4: Mine Support Equipment (Owner) ..................................................................................... 153

Table 21.2.2.1: LoM Process Operating Cost ............................................................................................ 153

Table 21.2.2.2: Annual Process Cost Detail ............................................................................................... 154

Table 21.2.3.1: G&A Operating Cost Summary ......................................................................................... 154

Table 22.1.1: Model Parameters ............................................................................................................... 155

Table 22.1.2: Financial Assumptions ......................................................................................................... 155 Table 22.2.1: Financial Model Results ....................................................................................................... 156

Table 22.2.2: Annual Production and Cashflow US$(000) .......................................................................... 157

Table 22.4.1: Project NPV Sensitivities (US$ million) ................................................................................. 158

Table 26.6.1.1: Summary of Costs for Recommended Work (US$) ............................................................ 164

Table 28.3.1: Definition of Terms ............................................................................................................... 169

Table 28.4.1: Abbreviations ....................................................................................................................... 170


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page ix

List of Figures Figure 4.1.1: Location Map of the Bovill Kaolin Project ................................................................................ 14

Figure 4.2.1: Land Management/Ownership Map ........................................................................................ 16

Figure 7.1.1: Regional Geology ................................................................................................................... 24 Figure 10.1.1: Drillhole Locations and Resource Areas ................................................................................ 33

Figure 11.5.1: UOI vs. GMT Analyses for 2011,-325 Mesh Duplicates ......................................................... 39

Figure 11.5.2: UOI vs. CCL Analyses for 2011,-325 Mesh Duplicates .......................................................... 40

Figure 11.5.3: UOI vs. GMT Analyses for 2013, -325 Mesh Duplicates ........................................................ 41

Figure 11.5.4: UOI vs. Non Certified Reference Material for the -325 Mesh, 2013 Analyses ......................... 42

Figure 11.5.5: UOI vs. GMT Analyses for 2013, -325 Mesh Duplicates, with Bias Removed ........................ 43

Figure 11.5.6: UOI vs. Non Certified Reference Material for the -325 Mesh, 2013 Analyses with Bias Removed ........................................................................................................................................ 44

Figure 13.1.1.1: Helmer-Bovill Clay Processing Plant Phase I Testing Block Flow Diagram ......................... 49

Figure 13.1.1.2: Helmer-Bovill Clay Processing Plant Phase II Testing Block Flow Diagram ........................ 51

Figure 13.2.3.1: SEM of 3" Hydrocyclone Overflow Showing Clays .............................................................. 62 Figure 13.2.3.2: SEM of 3: Hydrocyclone Underflow Showing Feldspathic Sand Grit ................................... 63

Figure 13.2.3.3: SEM of 0.5" Hydrocyclone 2nd Pass Overflow (Predominantly Halloysite) .......................... 64

Figure 13.2.3.4: SEM of 0.5" Hydrocyclone 2nd Pass Underflow (Predominantly Kaolinite) .......................... 65

Figure 14.7.1: Drillhole Locations (Black Dots), Clay Shell >=10% (Blue), Clay Shell >1-<10% (Teal) .......... 79

Figure 14.7.2 WBL, East West Cross Section 1,904,800N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste ............................................... 81

Figure 14.7.3 Middle Ridge, East West Cross Section 1,906490N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste ...................................... 82

Figure 14.7.4 Kelly’s Hump North, East West Cross Section 1,907,200N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste ..................... 83

Figure 14.7.5 Kelly’s Hump South, East West Cross Section 1,904,200N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste ..................... 84

Figure 16.2.3.1: Kelly’s Hump Results ....................................................................................................... 100

Figure 16.2.3.2: Middle Ridge Results ....................................................................................................... 101

Figure 16.3.2.1: Kelly’s Hump Final Design ............................................................................................... 103

Figure 16.3.2.2: Kelly South Final Design .................................................................................................. 104

Figure 16.3.2.3: Middle Ridge Final Design ............................................................................................... 105

Figure 16.3.2.4: WBL Final Design ............................................................................................................ 106

Figure 16.3.3.1: Kelly’s Hump and Kelly South Pit Designs and Optimized Shells ...................................... 109

Figure 16.3.3.2: Middle Ridge and WBL Pit Designs and Optimized Shells ................................................ 110

Figure 16.3.3.3: Cross Section (Designed pit versus Whittle™ shell) ......................................................... 111

Figure 16.3.4.1: Kelly’s Hump Waste Dump (Plan View) ............................................................................ 112


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page x Figure 16.3.4.2: Middle Ridge Waste Dump (Plan View) ............................................................................ 113

Figure 16.3.5.1: Pit and Waste Dump Layout............................................................................................. 114

Figure 16.3.5.2: Pit and Waste Dump Layout with Constraints ................................................................... 115

Figure 16.4.1: Total Material Movement .................................................................................................... 117

Figure 16.4.2: Clay Material Feed.............................................................................................................. 117

Figure 16.4.3: Sand Material Feed ............................................................................................................ 118

Figure 16.4.4: Resource Classification Feed.............................................................................................. 118

Figure 17.1: Simplified Material Balance Block Diagram ............................................................................ 124 Figure 17.2.1: WBL Primary Clay Process Simplified Flow Diagram .......................................................... 126

Figure 17.2.2: Feldspar and Quartz Process Simplified Flow Diagram ....................................................... 127

Figure 18.4.1: Simplified Water Balance Block Diagram ............................................................................ 136

Figure 22.4.1: Project NPV Sensitivities (US$ million) ................................................................................ 158

Appendices Appendix A: Certificates of Qualified Persons


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 1

1 Summary This report on the Bovill Kaolin-Halloysite-Quartz-Potassium Feldspar Project (Bovill Kaolin Project) was prepared as a National Instrument 43-101 (NI 43-101) Prefeasibility Technical Report for I-Minerals Inc. (I-Minerals) by SRK Consulting (U.S.), Inc. (SRK), Tetra Tech Inc. (Tetra Tech), HDR Engineering, Inc. (HDR) and Roberts & Schaefer Company, a KBR Company (R&S/KBR), collectively the Consultants. A brief summary of the results of prefeasibility study are presented below.

1.1 Property Description and Ownership The Project is a development stage open pit mining operation which will produce quartz sand, K-feldspar sand, kaolinite clay, metakaolin clay, and halloysite clay. The Project area has been mined historically for similar products. It is located at geographical coordinates 46° 53’ 14.7” N. latitude and 116° 28’ 11.7” W longitude (UTM NAD 27 Zone 11 N, coordinates 5,192,807 N and 440,392 W,) in Latah County, Idaho, USA. The property currently totals 5,140.64 acres in area. The mineral leases are not contiguous, but are concentrated within three surveyed townships near the town of Bovill, Idaho. I-Minerals has rights to develop the Project through mineral leases issued by the State of Idaho.

1.2 Geology and Mineralization The Project geology is underlain by the Thatuna Batholith, a granitic intrusive of Cretaceous age, composed mainly of Na-feldspar, K-feldspar and quartz. The mineral deposit is hosted within a weathered saprolite horizon which directly overlies the bedrock from which it was derived. During the natural processes of weathering, the original plagioclase feldspars have preferentially broken down to produce the clays kaolinite and halloysite. The K-feldspars have resisted weathering to a degree and much of the original component remains as free grains. Similarly, the quartz component of the host rock remains as free grains in the weathered material. The mineral resource products include kaolinite, halloysite, K-feldspar and quartz.

1.3 Status of Exploration, Development and Operations The exploration programs supporting the mineral resource estimate consists primarily of diamond core drilling, sampling, chemical analysis and material characterization studies. The drillhole database supporting the resource estimation of this report consists of 322 diamond core drillholes totaling 35,909 ft. The drillholes average 112 ft in length, all are oriented vertical and spaced on 100 to 200 ft centers. The original assay sample lengths generally range from 5 to 10 ft with an average of 5.8 ft. Analytical testing consists primarily of material characterization studies used to support the resource estimation of this report. This work involved two general areas of study including particle size analysis and clay characterization.

1.4 Mineral Processing and Metallurgical Testing The proposed processing operations for the Bovill Kaolin primary clays will produce Halloysite clay (7,500 t/y), Kaolinite clay (37,600 t/y), , K-feldspar (47,400 t/y), and quartz products (118,000 t/y). Conventional metallurgical techniques will be required to produce these products and will include:



clay/feldspathic sands classification, clay degritting, halloysite/kaolinite classification, wet grinding, flotation, filtration, drying, dry magnetic separation, dry fine grinding, and Processing techniques, flow schematics, and a material balance for the Bovill plant processing circuit was developed to assess product recoveries and material losses. Extensive laboratory and pilot plant scale testing was required to obtain this information, and was primarily performed by utilizing two well-known and respected industrial minerals laboratories that are considered to be industry leaders in their respective areas of expertise. Ginn Mineral Technology, Inc. (GMT) is located in the heart of the Georgia kaolin belt at Sandersville, GA and is a recognized expert in clay processing. North Carolina State University Minerals Research Laboratory (MRL) is located in Asheville, NC near extensive historical feldspar and quartz operations, and is a recognized expert in feldspar and quartz processing.

Clay

Initial tests to determine the responsiveness of the clay to commercial processing technology were completed by GMT with the following key results:

• The clays have good brightness values with extensive processing methods; • SEM images from the crude and processed materials for each sample illustrated the range

in particle size and particle morphology of the minerals and showed the presence of the halloysite with the kaolinite; and

• The products produced from the crude samples concentrated the clay (halloysite and kaolinite) to almost 100% and removed the majority of other mineral contaminants.

Based upon the metallurgical test work, a bulk crude ore sample was collected for pilot plant testing. The primary focus of the pilot plant evaluations was to separate the clay from the feldspathic sand and then concentrate the halloysite from the kaolin clay. Two tests were performed and reported in July 2008 and July 2010. Three additional smaller scale pilot plant testing followed up the July 2010 testing. Key results from these tests included:

• The 3 inch cyclone successfully split the grit fraction from the clay; • SEM’s illustrated excellent separation and concentration of the halloysite in the 2nd pass

overflow fraction using the 0.5 inch diameter cyclone; • A metakaolin product was successfully produced and provided for potential customer

development; • Total clay recovery should be in the range of 23%, halloysite 8%, and kaolinite 15%; and • Feldspathic sand fraction recovery should be in the range of 70%.

Additional metallurgical test work was performed in 2011 and 2012 on both a bulk crude ore sample as well as composite core ore material to confirm qualitative and quantitative data from previous work, produce additional material for product development, and investigate alternative separation processing techniques. Key results from these test included:

• The 0.5 inch cyclones could be successfully replaced with solid bowl decanter centrifuge technology which would simplify processing requirements for the halloysite/kaolinite separation;

• Significant additional halloysite product was produced for product development; and



• Recovery data for the composite core material mirrored that of previously run bulk samples,

both qualitatively and quantitatively, to confirm desired results.

Feldspathic Sand (K-feldspar and Quartz)

The MRL at NCSU performed a series of pilot plant tests on K-spar tailings. The objective was to prove the technical feasibility and developmental integration of processing techniques to produce commercial grade K-spar and quartz products. The quartz material was subjected to further bench-scale testing to produce varying degrees of quartz product quality. The MRL successfully achieved this objective and the following key results were obtained:

• The optimized feldspar circuit consistently produced quality K-spar product grades ranging from 18.0% to 18.5% Al2O3 and 13.0% to 13.5% K2O;

• K-spar yields from the feldspathic sand feed ranged from 17% to 20%; • The optimized quartz circuit, using varying degrees of flotation steps, resulted in the ability to

produce three grades of high quality quartz products. The three 50 x 200 mesh product grades ranged from 355 to 280 ppm impurities; and

• Quartz yields from the feldspathic sand feed ranged from approximately 24.5% for Quartz 1 to 13.6% for Quartz 3. The overall quartz yields will range from approximately 38% to 42% and will be dependent on the quartz production matrix.

1.5 Mineral Resource Estimate The mineral resource statement is presented in Table 1.5.1.The resource is confined within a Whittle™ pit design according to the parameters listed in Section 16.2. No cut-off grade (CoG) is applied to the resource because all recovered material in the resource estimation contains sufficient sand, kaolinite, or halloysite to be mined for a profit.



Table 1.5.1: Measured and Indicated Mineral Resource Statement (as of April 20, 2014)*

Classification Location Tons

(M) Qtz & K-Spar Sand

(%) Kaolinite

(%) Halloysite

(%) Qtz & K-Spar Sand Tons

(000’s) Kaolinite Tons

(000’s) Halloysite Tons

(000’s)

Measured Kellys Hump 2.3 76.7 13.0 3.9 1,761 297 90 Middle Ridge 1.0 74.8 12.0 5.9 745 120 58 All 3.3 76.1 12.7 4.5 2,505 417 148

Indicated

Kellys Hump 3.8 72.2 18.0 2.8 2,721 680 107 Middle Ridge 2.9 77.0 12.4 3.0 2,208 355 86 WBL Pit 1.3 75.0 15.8 1.7 973 204 22 All 7.9 74.4 15.6 2.7 5,902 1,239 215

M & I


* Note that values presented here have been rounded to reflect the level of accuracy. Source: SRK Consulting, 2014



Mineral Resources that are not mineral reserves do not have demonstrated economic viability. Mineral resource estimates do not account for mineability, selectivity, mining loss and dilution. These mineral resource estimates include mineral resources that are normally considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves. There is also no certainty that these inferred mineral resources will be converted to Measured and Indicated categories through further drilling, or into mineral reserves, once economic considerations are applied.

1.6 Mineral Reserve Estimate Mineral reserves have been estimated in accordance with CIM guidelines. CoGs were not applied since all weathered Thatuna material in the resource estimation contains sufficient sand, kaolinite or halloysite to be mined for a profit

The proven and probable reserves are presented in the Table 1.6.1.

Table 1.6.1: Mineral Reserve Statement, (as of June 14, 2014)

Reserve Mt Halloysite Grade

Kaolin Grade

Qtz & K-Spar Sand (%)

Halloysite Tons

Kaolinite Tons

Qtz & K-Spar Sand (t)

Kelly Hump Proven 1.7 4.8% 13.5% 81.7% 82,000 229,000 1,389,000 Probable 1.0 6.0% 15.4% 78.6% 60,000 154,000 782,000 Kelly South Proven 0 0 0 0 0 0 0 Probable 1.3 1.6% 23.2% 75.3% 20,000 296,000 959,000 Middle Ridge Proven 0.7 6.9% 12.8% 80.3% 48,000 90,000 563,000 Probable 1.4 4.6% 13.1% 82.3% 66,000 187,000 1,179,000 WBL Proven 0 0 0 0 0 0 0 Probable 0.8 2.4% 16.5% 81.1% 18,000 128,000 629,000 Source: SRK Notes:

1. Some numbers may not add up as a function of rounding. 2. Reserves are based on 100% mine recovery and 0% dilution. This is due to the small equipment being utilized and the

selectivity of the material being mined. This will require further review as part of additional studies. 3. Halloysite processing recovery is 90%. 4. Kaolinite processing recovery is 90%. 5. Quartz and K-Spar processing recovery is 68%. 6. Variable selling prices were used depending on supply. 7. There is an overall strip ratio of 0.69:1 (waste: ore).

1.7 Mining Methods Open pit mining methods will be utilized to extract all material. Four deposits have been identified as economic, with individual pits in each deposit.

Each pit pushback design was divided into benches for creating a yearly mine schedule. The mining is controlled by the required mix to the plant. To address the percentages of kaolinite, halloysite, and sand were converted into product tonnages. Those tonnages of all the products were calculated for each of the mining areas in order to get specific product tonnage (which replaces grade in this study) for each subdivision.



Mining for both pits take place on the slopes of low hills at an approximate elevation of 3,000 ft. The pit depth is limited by 75 ft. The pit design for each deposit was divided into several phases for production scheduling with 40 ft wide ramps at an in-pit road grade of 10% or less.

The mining operations will be performed by a mining contractor, and will be open pit mining using a hydraulic excavator and a wheel loader to load haul trucks for waste and ore haulage.

1.8 Project Infrastructure The infrastructure for this project has been designed and costs estimated to incorporate existing conditions along with developmental needs. Some road improvement is needed into the plant and mining areas to facilitate new traffic conditions for mine haul vehicles, product transport vehicles, and general employee vehicles. On-site services will include basic services commensurate to a facility such as designed.

Power and natural gas will be brought into the facilities by a local utilities supplier, Avista Corp. Although most of the water for the processing facility will be recycled, necessary make-up water will be supplied from ground water wells and a reservoir north of the mining areas as described in Section 18.

1.9 Environmental Studies and Permitting The Idaho Surface Mining Act requires the operator of a proposed surface mine to submit an operating and reclamation plan prior to mining. The Idaho Department of Lands (IDL) is the lead state agency overseeing surface mining activity in Idaho and the acquisition of related environmental and operating permits will be coordinated in close cooperation with IDL. A number of major plans and permits are required for the Project including an Idaho Mine Operation and Reclamation Plan (administered through IDL), an Air Quality Permit (administered by Idaho Department of Environmental Quality [IDEQ]), a Clean Water Act (CWA) Section 402 National Pollutant Discharge Elimination System (NPDES) (administered by Environmental Protection Agency) General Permit for Discharges from Construction Activities and the NPDES Multi-Sector General Permit (MSGP) for Stormwater Discharges Associated with Industrial Activity. The MSGP has been issued to I-Minerals. The only remaining federal permit required is the NPDES stormwater construction activities general permit with associated requirements for complying with the Endangered Species Act, Section 106 of the National Historic Preservation Act, and CWA Section 401 Certification. Studies, mitigation plans, impact analyses, and monitoring will be required to secure the major permits. Several other minor permits and approvals are required from County and State agencies.

A wetland and ordinary high water mark boundary delineation of the Project area was conducted in the summer and fall 2012. The delineation focused on areas of potential land disturbance associated with roads, processing facilities, quarries, and tailing storage facilities. The goal of the field study was to map areas of wetlands and waters of the U.S., and then to provide this information to the project design team so that disturbance of these areas could be avoided. As such, if there is no fill or disturbance to wetlands and other waters of the U.S., no Clean Water Act Section 404 permit would be required. A wetland delineation report will be submitted to the U.S. Army Corps of Engineers (USACE) for a formal jurisdictional determination.



1.10 Capital and Operating Costs Estimated life of mine (LoM) capital costs are shown in Table 1.11.1. The initial capital estimate is US$72.7 million. An additional US$18.2 million in sustaining capital will be required. Total LoM capital is estimated US$90.8 million.

Table 1.11.1: LoM Capital Cost Summary (US$000’s)

Description Initial Sustaining LoM Mining 919 467 1,386 Process & Infrastructure 59,538 9,797 69,335 Tailings Storage Facility 9,170 4,900 14,070 Owners Costs 3,056 3,000 6,056 Total Capital $72,682 $18,165 $90,847

Source: SRK

Operating costs are estimated on preliminary mine and process design criteria, engineering, as well as budgetary quotes. LoM operating costs are shown in Table 1.11.2. Over the LoM, operating costs will be about US$70.72/t of product.

Table 1.11.2: LoM Operating Cost Summary

Description US$/t-Product LoM (US$000’s) Contract Mining 8.89 44,340 Mine G&A (Owner) 1.00 5,000 Mine Support (Owner) 0.17 861 Processing 59.18 295,090 G&A 1.43 7,152 Total $70.68 $352,443

Source: SRK

1.11 Economic Analysis The financial analysis results, shown in Table 1.12.1, indicate a NPV 6% of US$212.7 million with an IRR of 30.5% (after estimated taxes). Payback will be in 3 years from the start of production. The following provides the basis of the SRK LoM plan and economics:

• A mine life of 25 years; • Product yields include, 3.8% halloysite, 6.9% kaolin, 7.3% metakaolin, 16.6% K-Spar, and

37.9% quartz, over the LoM; • An average operating cost of US$70.68/t-product; • Capital costs of US$90.8 million, consisting of initial capital costs of US$72.7 million and

sustaining capital over the LoM of US$18.2 million; • Mine closure cost, included in the above estimates is US$5.1 million; and • The analysis does not include provision for salvage value.



Table 1.12.1: Financial Model Results

Description Units US$ Value (000’s)

Unit Cost US$/t-Prod.

Production RoM Ore Processed kt 6,878 - Recovered Products kdt 4,986 - Estimate of Cash Flow Gross Revenue US$000’s 1,270,410 254.79 Freight & Marketing US$000’s (4,986) (1.00) Net Revenue US$000’s 1,265,424 253.79 Royalties US$000’s (63,521) (12.74) Gross Income US$000’s 1,201,904 241.05 Operating Costs Contract Mining US$000’s 44,340 8.89 Mine G&A US$000’s 5,000 1.00 Mine Support US$000’s 861 0.17 Process Opex US$000’s 295,090 59.18 G&A US$000’s 7,152 1.43 Operating Costs US$000’s 352,443 70.68 Operating Margin US$000’s 849,460 170.36 Capital Costs Mining US$000’s 1,394 - Process Capital US$000’s 69,335 - Tailings Capital US$000’s 14,070 - Owners Costs US$000’s 6,056 - Total Capital US$000’s 90,854 - Income Tax US$000’s (262,600) - Mining License Tax US$000’s 0 - Cash Flow US$000’s 496,006 - Present Value 6% 212,767 - IRR 30.5% -

Source: SRK

1.11.1 Sensitivity Analysis Sensitivity analysis for key economic parameters is shown in Table 1.12.1.1. The Project is most sensitive to product prices (revenues). The Project’s sensitivities to operating and capital costs are quite similar.

Table 1.12.1.1: Project NPV Sensitivities (US$ millions)

Unit -15% -10% -5% Base 5% 10% 15% Revenues 159 177 195 213 231 248 266 Operating Costs 228 223 218 213 208 203 198 Capital Costs 221 218 215 213 210 207 205

1.12 Conclusions and Recommendations An update to the 2013 Prefeasibility evaluations have been completed on I-Minerals’ Bovill Kaolin Project in Latah County, Idaho. The results of the study indicate that aluminosilicate mineral products (kaolin, halloysite, metakaolin, K-feldspar) and quartz can be produced in the desired purity, and that market interest exists in a number of fields such as ceramics, personal care and cosmetic products, glass, bricks, and cement products.



Project financial results justify advancing and refining the prefeasibility engineering designs to feasibility level during the next 12 to 18 months. SRK recommends that I-Minerals conduct additional infill drilling within the areas supporting the first 10 years of mine life to increase the confidence of the reserve to a proven classification. Additional elements of the recommended work program include environmental permit application support, additional geotechnical studies, testwork, modeling, and engineering designs for the tailings and waste storage; marketing studies; hydrogeologic modeling; more comprehensive metallurgical testing; and more detailed evaluation of process equipment and infrastructure. The recommended work programs are estimated to cost US$3 million.



2 Introduction 2.1 Terms of Reference and Purpose of the Report

This report was prepared as a National Instrument 43-101 (NI 43-101) Prefeasibility Technical Report for I-Minerals Inc. (I-Minerals) by SRK Consulting (U.S.), Inc. (SRK), Tetra Tech Inc. (Tetra Tech), HDR Engineering, Inc. (HDR) and Roberts & Schaefer Company, a KBR Company (R&S/KBR), or collectively (the Consultants). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in SRK’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by I-Minerals subject to the terms and conditions of its contract with SRK and relevant securities legislation. The contract permits I-Minerals to file this report as a Technical Report with Canadian securities regulatory authorities pursuant to NI 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. The responsibility for this disclosure remains with I-Minerals. The user of this document should ensure that this is the most recent Technical Report for the property as it is not valid if a new Technical Report has been issued.

This report provides mineral resource and mineral reserve estimates, and a classification of resources and reserves in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum Standards on Mineral Resources and Reserves: Definitions and Guidelines, November 27, 2010 (CIM).

Within this document, the term, “K-feldspar” refers collectively to the mineral members of the feldspar group containing a high potassium content, namely orthoclase, microcline and sanidine. In commercial terms, K2O is generally > 10%. The term “Na-feldspar” refers collectively to the mineral members of the plagioclase feldspar group containing high sodium content, namely albite, oligoclase and andesine. In commercial terms, Na2O is generally > 7%. The term “feldspar” refers collectively to a material mixture of either K-feldspar and/or Na-feldspar in undetermined ratios. The term “kaolin” is generally used as the name of a rock that is made up mostly or all of the clay minerals from the kaolinite group. “Kaolinite” and “halloysite” are two clay minerals in the kaolinite group and are hydrated aluminosilicates with platelet morphology. “Halloysite” contains extra water that causes an imperfection in its platelet, causing it to roll up, giving it a tube-like morphology.

2.2 Qualifications of Consultants (SRK) The Consultants preparing this technical report are specialists in the fields of geology, exploration, mineral resource and mineral reserve estimation and classification, mining, geotechnical, environmental, permitting, metallurgical testing, mineral processing, processing design, capital and operating cost estimation, and mineral economics.

None of the Consultants or any associates employed in the preparation of this report has any beneficial interest in I-Minerals. The Consultants are not insiders, associates, or affiliates of I-Minerals. The results of this Technical Report are not dependent upon any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning



any future business dealings between I-Minerals and the Consultants. The Consultants are being paid a fee for their work in accordance with normal professional consulting practice.

The following individuals, by virtue of their education, experience and professional association, are considered Qualified Persons (QP) as defined in the NI 43-101 standard, for this report, and are members in good standing of appropriate professional institutions. The QP’s are responsible for specific sections as follows:

• Bart Stryhas, PhD., CPG., SRK Associate Principal Consultant (Resource Geology), is the QP responsible for Geology and Resource Sections 1.1 through 1.3, 1.5, 2 through 12 except for 4.4 and 5.5, 14, 23, 25.1, 25.2.1, 25.2.2, 26.1, 27 and 28.

• Harry Gatley, P.E., R&S/KBR, is the QP responsible for Process and Metallurgy, Recovery and Infrastructure Sections 1.4, 1.8, 5.5, 13, 17, 18 except 18.5, 21.1.2, 21.2.2, 25.2.4 and 26.5.

• David S. Hallman, PE, Tetra Tech, is the QP responsible for Tailings Sections 18.5, 21.1.3, 25.2.5 and 26.4.

• Manuel Rauhut, PE, HDR Engineering, is the QP responsible for Environmental and Permitting Sections 1.9, 4.4, 20, 25.2.6 and 26.3;

• Valerie Obie, BS Mining, RM-SME, SRK Principal Consultant (Mineral Economics), is the QP responsible for Mining, Reserves and Economic Sections 1.6, 1.7, 1.10, 1.11, 1.12, 15, 16, 19, 21.1.1, 21.1.4, 21.2.1, 21.2.3, 22, 24, 25.2.3, 25.2.7, 26.2 and 26.6.

The Certificate of Author forms are provided in Appendix A.

2.3 Details of Inspection SRK Consulting – Lakewood, Colorado (SRK)

Dr. Bart Stryhas conducted site visits to the Project on May 12, 2010 and September 5, 2013. During the 2010 inspection, Dr. Stryhas was met by Mr. Lamar Long, Project Manager of I-Minerals. Mr. Long led a tour of the Project discussing the regional and local geology, inspection of the historic mining and materials, explanation of the proposed mining and processing plan and reclamation plan. Dr. Stryhas spent the afternoon in the field mapping.

On the 2013 inspection, Dr. Stryhas was met by I-Minerals staff including; Mr. Lamar Long, Project Manager and Mr. Gary Nelson, Metallurgical Operations Manager. The morning was spent observing diamond drilling procedures, sample chain of custody and geologic logging procedures. The afternoon was spent at the University of Idaho’s custom clay testing facility in Moscow Idaho. Staff at the laboratory provided detailed explanations of I-Mineral’s custom testing procedures, data transfer and capture, SEM analysis and internal QA/QC procedures.

SRK Principal Consultant Valerie Obie and other report contributors have not visited the site.

Tetra Tech Inc. – Denver, Colorado (Tetra Tech)

Report contributor Ms. Jessica Spriet, PE, completed a site visit in May of 2011. She toured the site with Mr. Lamar Long of I-Minerals and the previous QP for Tetra Tech, Mr. Justin Knudsen. Potential tailings storage facility locations were visited and potential design issues were identified. Mr. Knudsen has since left Tetra Tech and has been replaced by Mr. Dave Hallman as the QP. Mr. Hallman has not visited the site.



Roberts & Schaefer Company, a KBR Company - Sandy, Utah (R&S/KBR)

Mr. Paul Slade (Project Manager) and Mr. Pete Gagestein (Sr. Mechanical Designer) of R&S/KBR visited the Project site on October 6, 2010. They were met by Mr. Lamar Long and Mr. Gary Nelson of I-Minerals. A tour was provided of the proposed processing plant site, tailings facility, mining areas, and transload facility. In additional to obtaining a basic familiarization of the Project sites, a primary focus of the visit was to investigate the proposed plant site topography for general layout design to be used in future studies.

Mr. Laxmanarao Ravindra Nath, R&S/KBR Principal Process Engineer, visited the Minerals Research Laboratory (MRL) at North Carolina State University in Asheville, NC on November 28, 2007. The results from the MRL tests were the basis for the development of process and plant design criteria used in the PEA.

Mr. Harry Gatley, P.E., has expertise in the review and verification of the process and layout design. His previous experience with process design, water management, solids drying, and many other industry phases led to his new assignment as the QP for R&S/KBR for this study.

HDR Engineering, Inc. - Boise, Idaho (HDR)

Report contributor Michael Murray, PhD, HDR Vice President, visited the site on May 13, July 18, and July 19, 2012, to assess general site setting and conditions including surface drainage patterns, soil and vegetation conditions, wetlands, and wildlife habitat. Report contributor Christine L. Whittaker, LA., HDR Associate Vice President, visited the site July 18 and 19, and August 25 to 27, 2012, to conduct wetland delineations and vegetation survey as well as to assess overall site conditions to support mine permitting. HDR’s Manuel Rauhut and Diane Holloran have not visited the site.

2.4 Sources of Information SRK has used information primarily provided by I-Minerals or its consultants (Tetra Tech, Roberts & Schaefer, and HDR, Inc.) to prepare this pre-feasibility study. The pre-feasibility study builds on the results of prior technical reports and assessments completed by SRK (2010, 2012a, 2012b). Charles Rivers Associates provided market studies information. Operating costs in the technical economic model are based on data provided by I-Minerals and/or their consultants.

2.5 Effective Date The effective date of this report is April 20, 2014.

2.6 Units of Measure The U.S. System for weights and units has been used throughout this report. Tons are reported in short tons of 2,000 lb. All currency is in 1Q 2014 U.S. dollars (US$) unless otherwise stated.



3 Reliance on Other Experts The Consultants’ opinion contained herein is based on information provided by I-Minerals throughout the course of the investigations. The sources of information include data and reports supplied by I-Minerals personnel as well as documents referenced in Section 27.

The Consultants used their experience to determine if the information from previous reports was suitable for inclusion in this Technical Report and adjusted information that required amending. The QPs, Bart Stryhas, Ravi Nath, Dave Hallman, Christine Whittaker, and Valerie Obie, have examined the data for the Project provided by I-Minerals, and have relied upon that basic data to support the statements and opinions presented in this Technical Report. In the opinion of the QPs, the data is present in sufficient detail, is credible and verifiable in the field, and is an accurate representation of the Project.

This report includes technical information that requires subsequent calculations to derive sub-totals, totals, and weighted averages. Such calculations inherently involve a degree of rounding and consequently can introduce a margin of error. Where these rounding errors occur, they are not considered to be material.

The authors have relied upon the work of others to describe the geology, exploration, land tenure and land title in the state of Idaho, referring to Sections 4 – Property Description and Location and 4.2 – Mineral Titles and others. The information contained in these sections was obtained from “Report on the Helmer-Bovill Feldspar, Quartz and Kaolin mineral Leases, Latah County, Idaho” prepared on behalf of I-Minerals by James L Brown, PG.

Additionally, the authors have relied upon the independent market study by Charles River Associates (CRA) dated December 2011 for Section 19 – Market Studies and Contracts. CRA is a leading global consulting firm that offers economic, financial, and business management expertise to major law firms, corporations, accounting firms, and governments around the world. Their marketing expertise is well known in the business world.

The authors reviewed the December 2011 study, as mentioned above, against industry available data and information, and found the results to be within industry standards and adequate for use in this Technical Report at this level of study.

Prices for sand and clay products can vary dramatically depending upon the specifications and quality each product produced. Due to the highly competitive nature of the industrial sand and clay industry, contract prices are highly confidential and are not presented in public documents.

However, the QP for this Technical Report confirms that I-Minerals has, in fact, completed a market study for its products which included preliminary negotiations to supply a series of clay and sand products. The QP also confirms that the process facility is capable of producing these products.



4 Property Description and Location 4.1 Property Location

The Project is a development stage open pit mining operation which will produce quartz sand, K-feldspar sand, kaolinite clay, metakaolin clay and halloysite clay. The Project area has been mined historically for similar products. This section summarizes information related to the property location, mineral titles, royalties and agreements, environmental permits and liabilities, and Project risks.

The Project is located at geographical coordinates 46° 53’ 14.7” N. latitude and 116° 28’ 11.7” W longitude (UTM NAD 27 Zone 11 N, coordinates 5,192,807 N and 440,392 W,) in Latah County, Idaho, USA. (Figure 4.1.1).The property currently totals 5,140.64 ac in area. The mineral leases are not contiguous, but are situated within three surveyed townships near the town of Bovill, Idaho.

Source: HDR Engineering, 2012

Figure 4.1.1: Location Map of the Bovill Kaolin Project

4.2 Mineral Titles The Project area is located on endowment lands owned and administered by the IDL. These and other IDL holdings across the state of Idaho were granted to the state in 1890 by the federal government on the condition they produce maximum long-term financial returns for public schools and other beneficiaries. Therefore, IDL has a mandate for these lands to produce revenue to support the state’s public school system and other state institutions. To achieve this, IDL manages these



properties primarily for profit through the production of timber, livestock grazing, and the extraction of mineable materials.

The State of Idaho endowments lands fall in two categories referred to as Fee Simple (FS) and Minerals Only (MO). The FS lands are where the State owns both mineral and surface rights. The MO lands are where the State owns mineral rights but someone else owns surface rights. The majority of the lands held by I-Minerals are FS. All mineral resources and mineral reserves described in this report are located on FS lands. By way of its mineral leases, I Minerals has surface rights and legal access to the Project provided it meets all permitting and bonding requirements administered by IDL. In the State of Idaho, mineral leases are not required to be physically located in the field. The mineral leases are currently described only on paper by the U.S. Public Land Survey Grid.

In 2002, I-Minerals acquired from Idaho Industrial Minerals (IIM), through its wholly owned subsidiary Alchemy Kaolin Corporation, 16 State of Idaho mineral lease applications in Latah County, Idaho, to cover deposits of feldspar, kaolin, and quartz located near Bovill, Idaho. In 2003, I-Minerals converted these applications to ten mineral leases and subsequently obtained two more mineral leases. The Project then consisted of 12 Idaho State mineral leases. Renewal applications for all 12 leases were filed on April 27, 2012 with a US3,000 application fee. As part of the renewal process, the State converted the 12 mineral leases into 10 revised mineral leases which were issued on February 28, 2013. Subsequently, during 2013 the State of Idaho granted one additional mineral lease to I-Minerals. At this time, I-Minerals holds 11 mineral leases totaling 5,140.64 acres. All these are valid until 2023, at which time; they can be renewed for an additional ten years. All leases are subject to rental fees of US$1.00/acre/y and a production royalty of 5.0% based on gross proceeds. The production royalty is prepaid at a rate of US500/lease for the first five years and increases to US1,000/lease for the second five years of the lease. Details of the mineral leases are listed in Table 4.2.1 and shown in Figure 4.2.1.

Table 4.2.1: Mineral Leases

Source: HDR Engineering 2013



Source: HDR Engineering 2013

Figure 4.2.1: Land Management/Ownership Map

4.3 Royalties, Agreements and Encumbrances I-Minerals has rights to develop the Project through minerals leases issued by the State of Idaho (Leases). These Leases were acquired from IIM and are held by I-Minerals, based upon a certain Assignment Agreement with Contingent Right of Reverter (the Agreement), dated August 12, 2002, between I-Minerals USA (formerly Alchemy Kaolin Corporation) and IIM. The Agreement was subject to several amendments and ratifications between the parties, dated effective August 10, 2005, August 10, 2008, and January 21, 2010. Under the terms of the Agreement, I-Minerals acquired a 100% interest in the property upon final issuance of a total of 1.75 million shares of common stock to IIM. The issuance of these shares occurred on a staged basis following completion of a series of defined work programs conducted during the different phases of Project development, subject to approval by the TSX-V, which was granted by letter dated January 18, 2013. The final block of shares was delivered on January 23, 2013.

The State of Idaho retains a 5% gross production royalty due upon commencement of any mineral production.



4.4 Environmental Liabilities and Permitting

4.4.1 Environmental Liabilities The mineral leases held by I-Minerals cover areas of historic open pit mining. These areas include open pit mines, waste dumps and tailings areas. At this time, there are no known environmental liabilities associated with the exploration work conducted by I-Minerals and all activities to date are covered under general state and federal authorizations for exploratory work.

I-Minerals submitted an original bond of US$750 to the IDL to cover environmental liabilities associated with its exploration work. This bond has remained in place throughout the work programs but it was refunded in December, 2012. On November 1, 2010, the State of Idaho changed its bonding program. I-Minerals has been paying a reclamation bond of US100/lease per year since that time. All reclamation bonding is current through October 31, 2014. The IDL has approved all reclamation conducted to date.

4.4.2 Required Permits and Status I-Minerals is currently permitted for the following activities at the Bovill Kaolin Project site (IDL mineral leased lands):

Exploration activities – I-Minerals has conducted exploration activities in accordance with Idaho Administrative Procedure Act (IDAPA) 20.03.02.060 – Exploration Operations and Required Reclamation. I-Minerals filed an original Notification of Exploration (NOE) to the IDL in 2000, which was subsequently amended for surface exploration and drilling programs. Exploration disturbances have been reclaimed and approved by IDL.

Mining activities – I-Minerals is permitted through an approved Mine Plan of Operations and Reclamation Plan from IDL for the mining of approximately 10 acres of feldspathic sands from June through October for up to 10 years (2012 through 2022). The feldspathic sands were deposited as tailings from clay mining operations that occurred on or near the Company’s mineral leases between 1956 and 1974. These activities are conducted under aNPDES 2008 MSGP for Stormwater Discharges Associated with Industrial Activities (Permit Number IDR05CU73). The stormwater permit became effective on November 8, 2012.

Permits to be Acquired for Project

A review of Project plans identified a range of environmental permits, review processes, and authorizations needed for construction, operation, and closure. Development of the Project would require approval of a Plan of Operations and Reclamation Plan by IDL (IDAPA 20.03.02), a NOI for coverage under the federal NPDES General Permit for Stormwater Discharges from Construction Activities, and a NOI for coverage under the NPDES MSGP for industrial activities (Sector J3: Chemcal and Fertilizer Mineral Mining and Dressing/Clay, Ceramic, and Refractory Materials). In addition, a state air quality permit for emission sources including dryer stacks and fugitive dust will be required. Closure of the mine requires approval through IDL of a detailed mine site reclamation and tailings closure plan. Also, monitoring of certain resources will likely be mandated through the state mine permitting process as well as through the federal NDPES stormwater general permits.

The goal of Project design is to avoid fill in jurisdictional wetlands and other waters of the US, such that no Clean Water Act Section 404 permit would be required. No federal lands or federal permits



(except for the stormwater general permits) are involved in the Project plans, and as such no National Environmental Policy Act (NEPA) environmental review of the proposed Project is anticipated (other than resource information required as part of the stormwater general permits).

A description of permitting requirements, risks, and other important factors are described in Section 20.

4.5 Other Significant Factors and Risks There are no other risks or significant factors known at this time that may affect access, title, or the right or ability to form work on the Project.



5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Topography, Elevation and Vegetation The Project is located on property that belongs to the State of Idaho and is administered by the Idaho Department of Lands (IDL). The average elevation is about 3,000 ft above mean sea level, with a topographic relief of about 200 ft. The area is largely covered with soil, but old workings (pits and trenches) and road cuts provide exposure to the underlying bedrock geology. The Project is located on the west side of the Potlatch River drainage area.

The Project area consists of low foothills and ridges alternating with relatively wide, flat basins. Forested areas occupy the slopes and ridge tops which are managed primarily for timber production. Conifer forest makes up approximately 50% of the overall Project area. Forest stands were observed to be early seral, highly fragmented, and lacking in the ecological functions and values of older, more contiguous forests. Grasslands occur in the basins alongside sinuous intermittent and perennial stream channels. The Project area is currently permitted for livestock grazing. Most of the Project area has been disturbed by previous mining, forestry and grazing activities and, as such, contain predominantly disturbance oriented plant communities. Non-forested meadows or pasture areas are intensively grazed resulting in a proliferation of non-native vegetation and soil compaction and erosion.

Surface waters primarily consist of small, meandering, intermittent stream channels that flow toward the Potlatch River. These channels are typically located in the level “flats” between low hills or ridgelines and dry up by mid or late summer. Most streams are hydrologically altered by high-density road construction, historic mining, and cattle grazing. Grazing has also eliminated much of the woody growth along most stream channels resulting in eroded channels and sedimentation. Other surface waters include several old clay mining pits and small dams that have developed into water catchment basins as well as emergent wetlands flanking the stream channels. Groundwater appears in scattered locations as either springs or seepage discharge along streams or edges of wetlands. Native soils predominate in the area.

5.2 Accessibility and Transportation to the Property The Project is accessed by road from the town of Bovill by following State Highway ID-8 W for 0.4 mi, then turning right on Moose Creek Road/National Forest Road 381 and following for 5.5 mi. ID-8 W is an improved two-lane road, while Moose Creek Road/National Forest Development Road 381 is a dirt/gravel road that provides access to State and Federal lands. In addition, access to specific areas to be mined will require either upgrades to former logging roads or construction of new access roads.

5.2.1 Proximity to Population Center and Transport The nearest, large communities are Moscow, Idaho, which lies about 28 miles WSW from the Project, and Lewiston, Idaho, which lies about 33 miles to the southwest. Transport to the Project would utilize standard over-highway vehicles.



5.3 Climate and Length of Operating Season The climate at the Project site, as described by the Western Regional Climate Centers (WRCC), Clarkia Ranger Station, Idaho, is characterized by an estimated average annual precipitation of 38.82 inches, with the highest values recorded between October and March (71% of the annual precipitation). The average annual minimum and maximum temperatures are 30.1°F and 55.7oF, respectively; with average monthly minimum and maximum temperatures ranging from 18.5°F to 41.7°F and 41.1°F to 83.3°F, respectively.

Available records (February 1950 to February 1975) from the Clarkia Ranger Station weather station indicate an average total snowfall ranging from 0.1 inch in October to 37.3 inches in January, with an annual average of 100.9 inches. Average snow depth ranges from 1.0 inch in November to 23.0 inches in February, with an annual average snow depth of 6.0 inches (WRCC, 2010).

5.4 Sufficiency of Surface Rights The surface ownership of the 11 mineral leases is a mixture of private land owners and the State of Idaho. The surface right of the mineral leases specific to the resource estimation are owned and administered by the State of Idaho. The U.S. Army Corps of Engineers (USACE) owns the surface rights of all waterways located within the mineral leases.

5.5 Infrastructure Availability and Sources Electric power would be provided by Avista Corp. Approximately four miles of power lines would need to be constructed.

Natural gas is available to the Project from a natural gas pipeline that extends from Moscow to Bovill and is available to be utilized for this processing facility. Approximately two miles of pipeline would need to be constructed.

Water needed for processing would come from new wells located at the process site. Groundwater from drilled wells is typically used to serve domestic needs within the vicinity of the Project. Additional water is also available in a small reservoir north of the Project. I-Minerals intends to apply for water rights to this reservoir in the near future.

The region has a long history of clay production, forestry and farming. A labor force skilled in heavy equipment operation, trucking, and general labor exists within the surrounding communities and rural areas. Additional information about the local community is provided in Section 20.3.

There are several suitable locations for potential tailings storage, mining waste disposal, and potential processing plants.



6 History Section 6 summarizes the prior history of the Project.

6.1 Prior Ownership, Exploration, Resource Estimates and Production U.S. Bureau of Mines (USBM) and United States Geological Survey (USGS) (1942-1947)

During WWII, the clays in eastern Washington and northern Idaho were examined as a possible source of alumina and a substitute for foreign bauxite ores. Domestic bauxite reserves were being depleted, and the importation of foreign bauxites was handicapped by transportation difficulties (Hosterman, et al., 1960). Both the USGS and USBM conducted extensive field studies that were followed by the drilling of 650 holes that totaled about 20,252 ft. From this work, over 300 Mt of clay were identified in this region with available alumina greater than 20%. About 90% of this tonnage was found in four deposits in Latah County-Bovill, Olson, Canfield-Rogers and Benson deposits (Hosterman, et al., 1960).

At the Bovill deposit, just west of the town of Bovill, 11 holes were drilled on approximately 1,000 ft spacing in an 8,000 ft x 14,000 ft area. The average overburden in this area, which is about 900 ac in size, is about 10 ft thick and the average thickness of the clay is 21 ft. Using a density of 2.15 g/cm3 for clay in place, Hubbard (1956) calculated an indicated clay resource of over 57 Mt, where the clay has available alumina averaging 21.8% and Fe2O3 4.0%. He also estimated an additional inferred clay resource of 27 Mt for a 650 ac area, adjacent to the aforementioned area, with a clay layer thickness of 20 ft, an available alumina of 20%, and Fe2O3 of 4.0%. Both of these resource calculations are unconfirmed and uncategorized in terms of NI 43-101.

USBM (1953-1963)

In 1953 the USBM continued their search for viable clay deposits. They also investigated the potential of the contained silica sand for the glass industry. The USBM tested the Benson and Olsen clay deposits between Troy and Deary, and then moved on to the Bovill deposits. Ninety-seven samples were collected from 1,325 ft of drilling over an area covering 750 ft x 350 ft that is located 1.5 miles southwest of Bovill near State Highway 8 (Kelly, et al., 1963).

A.P. Green Refractories Company (1956-1993)

In 1956, A.P. Green Refractories Company purchased all the remaining assets of Troy Brick and Clay and acquired a lease on Section 11, T.40N., R.1W., north of Helmer, from which they produced refractory clay. They processed the clay by air flotation to produce two grades of refractory clay. Production continued until the early 1990’s when Hammond Engineering purchased one pit from A.P. Green. This pit produced transported clay for ceramic applications. Total production from the area during this period is estimated to be 250,000 t. Erikson (1986) estimates that 10,000 t/y were produced from 1956 to 1986.

J.R. Simplot Company (1956-1974)

In 1956, the J.R. Simplot Company (Simplot) of Boise, Idaho, acquired leases covering the Bovill deposits. In a cooperative program, Simplot and USBM drilled 240 holes (99 of which were on 50 ft centers) and conducted washing, pyrometric, mineralogical, and beneficiation tests (Kelly, et al., 1963). By 1962, Simplot had built a clay plant, the Miclasil facility, for the production of paper fillers



and specialty ceramics (Hosterman and Prater, 1964). Production initially came from pits in the Bovill deposit as defined by Kelly, et al. (1963), which are in transported clay of the Latah formation directly south of the plant. Simplot shifted production to residual clay deposits in the granodiorite, as this source proved more satisfactory for paper filler (Hosterman and Prater, 1964). The pits exploited by Simplot for residual clays were the WBL north and south pits in Section 23, T41N, R1W and the Moose Creek Clay Mine in Section 28, T41N, R1W in the Moose Meadows area, and the Stanford pit in Section 5, T40N, R3W. Simplot operated their plant until 1974, when it was sub-leased to Clayburn Industries of British Columbia (Rains, 1991). Reportedly, Clayburn operated the property only a few years, calcining clay that was shipped to Canada and processed into super-duty and 70% alumina bricks. In 1994 the plant was dismantled and the property partially reclaimed.

Several Companies (1983-1986)

During the mid-1980’s a number of companies began exploration work in the Helmer-Bovill area to identify clays suitable for use as paper fillers and coaters. The University of Indiana, Nord Resources, Miles Industrial Mineral Research, and Cominco American conducted work on the Helmer-Bovill area deposits. In 1985-1986, the Erikson- Nisbet Partnership formed a consortium of companies to develop new processes for beneficiation of the clays, but the introduction of precipitated calcium carbonate (PCC) fillers for paper reduced the demand for kaolin fillers.



7 Geological Setting and Mineralization This Section summarizes the regional and local geology and mineralization found in the deposit.

7.1 Regional Geology The regional geology is dominated by Precambrian sedimentary rocks of the Belt Supergroup (Belt), which have been strongly deformed and intruded with granitic phases of the Idaho Batholith during the Cretaceous age Sevier Orogeny.

During the Middle Proterozoic, the area was dominated by a large intra-cratonic basin that was subsiding along syn-sedimentary faults. The basin sediments comprise the Belt which range in age from about 1,470 to 1,400 Ma. The oldest units consist of the Lower Belt sequence, these are overlain by the Middle Belt Carbonates and the youngest are the Missoula Group.

The Belt sediments are believed to have remained relatively stable until approximately 1,350 Ma when portions of the basin were affected by compressional tectonics of the East Kootenay Orogeny. This orogeny was followed by rifting of the basin during the late Proterozoic-early Paleozoic when large portions of the sediments were transported away and the western margin of North America was developed.

The next major tectonic event occurred during the Cretaceous Sevier Orogeny. Early compressional tectonics dominated the area forming large-scale folds, reverse and thrust faults. During the late Cretaceous, the Bitterroot Lobe of the Idaho Batholith was emplaced in the region. The intrusive rocks described below were formed during this event.

The most recent, significant, geologic event was the deposition of the Columbia River Basalts (CRB). The CRB consist of a large plateau flow sequence of Miocene age (6 to 17 Ma). The lavas are distributed over an extensive area covering portions of Idaho, Oregon, and Washington. Minor extensional block faulting has resulted in much of the present landscape. Figure 7.1.1 illustrates the regional geology of the Project.



Source: I-Minerals, 2014

Figure 7.1.1: Regional Geology

7.2 Local Geology Belt Series (pm)

The Precambrian metasediments of the Belt series are the oldest rocks in the Bovill-Moscow area and form the basement for the entire area (Tullis, 1944). The Belt series rocks crop out primarily in the northern and eastern sections of the Property. They form a high-grade metamorphic facies assemblage that includes gneiss, schist, and minor meta-quartzite, meta-argillite, and meta-siltite.

Thatuna Granodiorite

Granitoid intrusive rocks of Cretaceous age underlie a large portion of the Helmer-Bovill area and form part of a body referred to by Tullis (1944) as the Thatuna batholith. He believed that this intrusive body was separate from the Idaho batholith, owing to the distance between the two. However, Priebe and Bush (1999) consider the Thatuna granodiorite to be a lobe of the Idaho batholith. Tullis (1944) reported the Thatuna lithologies to consist predominantly of granodiorite with subordinate adamellite, tonalite, and granite. The principal mineral constituents are quartz, plagioclase feldspar, K-feldspar, and biotite with trace to minor amounts of muscovite, garnet, and



epidote. The batholith is medium- to coarse-grained granular, and porphyritic textures are common. Erosion of the Thatuna batholith developed a mature topography where it is exposed in Latah County (Kelly, et al., 1963).

Recent geological mapping done for the benefit of this Project, detailed in an internal company report by Clark (2003a), identified a previously undescribed phase of the Thatuna batholith, referred to as the Kmcp. The Kmcp is interpreted to be a border zone of the intrusion that occurs along the interface between the main-stage, coarse-grained, and porphyritic Thatuna batholith and the Precambrian Belt series roof rocks. Intrusion into cooler roof rocks resulted in a distinctive and texturally diverse unit characterized by dominant granular medium-grained and subordinate coarse-grained and pegmatoid textures, the lack of well-developed porphyritic textures and the presence of Precambrian xenolithic paragneiss, paraschist and metasiltite blocks inherited from the roof rocks. Where unaltered, the Kmcp intrusive rocks contain a primary assemblage of plagioclase, K-feldspar, quartz, biotite, and muscovite, and are predominantly of granodioritic to granitic composition. The porphyritic main body of the Thatuna batholith (Kg, Kgd) does not appear to crop out within the mapped part of the Helmer-Bovill area.

According to Clark (2003a), the Kmcp derives its distinctive character from high-level interaction with the Precambrian metasedimentary roof rocks. More rapid cooling in the contact zone produced a dominant medium-grained, non-porphyritic, granodioritic unit in contrast to the coarser-grained, porphyritic granodiorite lithology that characterizes the deeper main stage of the batholith (Kg, Kgd). In the roof zone, hydrous mineral-bearing xenolithic blocks of the Precambrian Belt series metasediments were entrained by the intruding magma and outgassed of their volatile component. The outgassing contributes to the creation of pockets of hydrous granitic liquid proximal to the Precambrian blocks. These pockets crystallized subsequently into coarse-grained to pegmatoid granite pods that are distributed within the larger body of medium-grained granodiorite. Owing to the physicochemical conditions of crystallization within the hydrous pods of granitic liquid, the resultant solidified rocks show a stronger tendency toward higher proportions of K-feldspar relative to plagioclase and higher K2O/Na2O ratios than does the dominant medium-grained granodiorite.

Weathered Thatuna Granitoid

The exposed Thatuna batholith was subjected to intense weathering in a tropical or near-tropical climate during the Miocene epoch, while the Columbia River basalts were erupted and the Latah formation sediments were deposited (Hosterman, et al., 1960). In response to the strong weathering, much of the feldspar and at least some of the mica in the igneous body were altered to one or more varieties of clay minerals. The depth limit of weathering may initially have been fairly consistent; however, subsequent erosion has left a variable weathering profile with thickness roughly dependent on topography. At present, the depth of weathering may exceed 100 ft along ridges and be less than 3 ft in some valleys.

The weathering profile of the Thatuna granodiorite is currently under study by I-Minerals, owing to its importance in determining the range of mineral products that can potentially be produced from a given area on the Property. Of particular importance is the weathering of the feldspar in the granitoids to halloysitic to kaolinitic clays. It was the presence of kaolinitic clay deposits that provided the initial impetus for economic mineral development in north Idaho. Plagioclase (Na-Ca bearing) feldspar is the least stable phase in the weathering environment, and it alters to form clay well before K-feldspar and muscovite (Murray, et al., 1978). K-feldspar and the micas (biotite and muscovite) are



relatively resistant to alteration during all but the most intense weathering. Quartz is impervious to alteration throughout the weathering cycle. In the Helmer-Bovill area, pits that were mined for kaolin in residual deposits contained mostly quartz, halloysite, kaolinite, and K-feldspar. The waste material is primarily quartz and K-feldspar, with Na-feldspar (plagioclase) accounting for only a minor proportion of the total feldspar. Na-feldspar made up less than 5% of the total feldspar in a tailings sample examined by Clark (2003b). Residual clay deposits in the Helmer-Bovill area reflect this mineral distribution and targeted commodities from strongly-weathered Thatuna granitoid are kaolin, quartz, and K-feldspar.

Potato Hill Volcanics (Tphy, Trdy)

The Potato Hill volcanic rocks were tentatively considered to be of Permian age by Tullis (1944), but Bush, et al. (1999) interpreted field relationships to indicate an Eocene age. The silicic to intermediate volcanic rocks include both lava flow and pyroclastic flow units, as well as hypabyssal intrusive rocks. They form much of the rock along the western edge of the Helmer embayment at Potato Hill, and along the southern edge of the Thatuna. Many of the pyroclastic flows contain abundant xenolithic clasts of older granodiorite and Belt metasediments.

The individual flows are 3 to 50 ft thick and the complete sequence exceeds 900 ft in thickness. The flow units generally contain 3% to 10% phenocrysts of feldspar and quartz distributed in an aphanitic matrix of devitrified volcanic glass. Accessory minerals include magnetite, hornblende, apatite, and zircon. Some lithic-rich pyroclastic flow units carry up to 20% fragments. The saprolitic weathering that is well-developed in the older rocks has not appreciably affected the Potato Hill volcanics.

Columbia River Basalts (Tcrb)

Swanson, et al. (1979) described the stratigraphy of the basaltic units and divided them into 14 members assigned to five formations. Two flow units are interpreted to have reached Latah County by Swanson, et al. (1979), although Priebe and Bush (1999) have mapped at least five distinct flow units. The First Normal member of the Grande Ronde formation, the Priest Rapids member of the Wanapum formation, and the Onaway member of the Saddle Mountain formation (oldest to youngest, respectively) are all Columbia River basalt flows mapped by Priebe and Bush (1999) in the Helmer-Bovill area. The Grande Ronde formation flow occurs in the southern portion of the Helmer-Bovill area and consists of fine-grained to very fine-grained aphyric basalt. The Priest Rapids flow is a medium to course-grained basalt with microphenocrysts of plagioclase and olivine in a groundmass of intergranular pyroxene, ilmenite, and devitrified glass. It crops out in increasing abundance to the southwest toward Deary. Saddle Mountain basalts are found much further to the west. The importance of the Columbia River basalts to the genesis of the Latah formation is that the episodic basaltic extrusion dammed streams and formed lakes into which kaolin-rich sediments eroded from weathered granitoid and Precambrian metasediments were deposited (Kirkham and Johnson, 1929).

Latah Formation (Tsb)

Kirkham and Johnson (1929) described the Latah formation as lake bed sediments that, although local in origin and distributed in disconnected basins, occur over an area 175 miles long and 75 miles wide in eastern Washington and northern Idaho. Episodic flows of the Columbia River basalts blocked streams and formed lakes that collected sediments eroded from surrounding rocks. In the Helmer-Bovill area, a major basin termed the Helmer embayment (Hosterman, et al., 1960) occurs



over an area of approximately 25 to 30 mi2. Latah formation sediments are termed the sediments of Bovill (Tsb) by Bush, et al. (1999) and are described as clay, silt, sand and minor gravel deposits that are laterally equivalent with and overlie flows of Columbia River basalts. The clays are white, yellow, red and brown in color, kaolinite-rich, and range from a few feet to several tens of feet in thickness.

Palouse Formation

The Palouse Formation comprises mixed loess and flood plain sediments of Pleistocene age. It ranges in thickness from 3 to 35 ft in thickness and averages 10 ft thick in the Helmer embayment. The unconsolidated layers also include volcanic ash from the eruption of various Cascade Range volcanoes.

7.3 Significant Mineralized Zones The Project hosts four different deposit types. These include primary Na-feldspar deposits, residual K-feldspar-quartz-kaolinite±halloysite deposits, transported clay deposits and K-feldspar-quartz tailings deposits.

The primary Na-feldspar deposits are hosted within granitic border phases of the Thatuna granodiorite. These deposits are described in detail in a previous I-Minerals report (SRK, 2010).

The transported clay deposits are hosted primarily within the Latah formation. This formation was deposited primarily in shallow lakes dammed by Columbia River Basalts. Extensive weathering of feldspathic source terrains constitutes the provenance of these clays.

The K-feldspar-quartz tailing deposits are the result of previous mining and washing of the residual deposits. Here, the majority if the clay has been removed and the tailings are composed primarily of K-feldspar and quartz. These deposits are described in detail in a previous I-Minerals report (SRK, 2012b).

The residual deposits are derived from saprolitic weathering of the Thatuna granodiorite-granitic phases. In general, the Na-feldspar alters to kaolinite and halloysite. These clays are accompanied by residual K-feldspar and quartz. These deposits are described in detail in a previous I-Minerals report (SRK, 2012a) and are the subject of this report.

The information in the following sections has been cited with minor modifications from “The report on the Helmer-Bovill feldspar, quartz, and kaolin mineral leases, Latah County, Idaho” on behalf of I-Minerals Vancouver, B.C. by James L. Browne, PG March 13, 2006. These citations are intended to inform the reader of the general geologic setting as it pertains to the four deposit types described above.

7.3.1 Feldspars Tullis (1944) described the main lithologies in the Thatuna Batholith as consisting primarily of granodiorite with subordinate adamellite and tonalite and minor granite. Total feldspar content in these intrusive rocks is reported by Tullis (1944) to range between 47.4% and 80.6%, with an average of about 62.7% total feldspar. By definition (Streckeisen, 1976), granodiorite contains an abundance of plagioclase feldspar in excess of 65% of the total feldspar. Thus, the unweathered Thatuna represents a source carrying a high total feldspar abundance, of which a significant proportion is Na-bearing feldspar (sodic plagioclase).



Clark (2003a) collected many samples of Thatuna batholithic rocks from the Moose Meadows portion of the Helmer-Bovill area during his mapping program in 2002 and from core drilled during the 2000-2001 diamond-drilling program. Results from the petrographic work indicate that intrusive lithologies range from granodiorite to quartz monzonite (one sample) to granite, with granodioritic rocks being the most common. Estimated total feldspar abundances for these samples range from 60% to 82% and average about 71.5%. Following the petrographic and cathodoluminescence work, electron microprobe analyses of feldspars and quartz from representative samples were undertaken in order to quantify feldspar compositions and determine potential product quality in terms of alkali abundances and suitably low Fe2O3 contents (Clark 2003a, 2003b). Petrographic analyses of the Kmcp samples show that contained feldspars rarely have inclusions of Fe-bearing minerals (biotite, muscovite, or FeOx; Clark, 2003a, b).

In the strongly weathered Thatuna Batholith rocks plagioclase shows nearly complete alteration to a kaolin mineral, but much of the K-feldspar survives alteration. This is illustrated by sample IK81, collected from the Stanford Pit, about 11 miles WSW of the Moose Meadows area. Plagioclase was not identified by the X-Ray Diffraction (XRD) analysis. These results correspond well with the mineralogy of the material in the tailings impoundment adjacent to the pit. The tailings contain essential quartz and K-feldspar, some clay/mica, and only minor amounts of plagioclase.

7.3.2 Quartz Petrographic examination of 21 granitoid samples from the Moose Meadows area led Clark (2003a) to conclude that quartz in Thatuna batholithic rocks is relatively free of Fe-bearing mica or oxide inclusions. The average quartz composition calculated from electron microprobe analyses of quartz in drill core, surface outcrop, and processed quartz product samples from the Moose Meadows area granitoid (Clark, 2003b) is given in Table 7.3.2.1, along with two analyses of MRL quartz products from Moose Meadows granitoid (WBL pit and an interval from drillhole MC-22) and one of a commercial mid-western U.S. glass sand:

Table 7.3.2.1: Average Quartz Composition Calculated from Electron Microprobe Analyses

Product SiO2 (%) Al2O3 (%) Fe2O3 (%) CaO (%) Na2O (%) K2O (%) Avg. quartz analysis >99.9 0.004 0.003 0.004 0.009 0.007 MRL-P quartz prod 99.8 0.15 <0.01 <0.01 <0.05 0.08 MC-22 quartz prod 99.7 0.19 <0.01 <0.01 0.08 0.05 Mid-west glass sand 99.5 0.15 0.09 0.01 0.01 0.04 Source: Clark, 2003b

The analytical values for the trace elements in the quartz are very near or below detection limits for the electron microprobe and indicate that quartz from the Moose Meadows area is essentially free of impurities. This data suggests that the area has excellent potential to produce a glass-grade product that might be processed further into feed stocks for the high purity quartz market.

7.3.3 Clay Minerals The kaolinite group of clay minerals includes four minerals that are similar chemically, but differ with regard to crystal structure. Kaolinite and halloysite, two of these kaolinite group minerals, are the major clay minerals in the Helmer-Bovill area clay deposits. Crystal structure differences are important and control properties relevant to their commercial applications. Kaolinite occurs as distinct



platelets, whereas halloysite forms tubes and spheroids. Although halloysite also has a plate-like crystal form, imperfections in its crystal lattice cause the crystal to “roll up” into the tubular forms. There are two varieties of halloysite, the four-water variety and the two-water variety. The two-water variety is a dehydrated version of the four-water halloysite and is almost impossible to distinguish from poorly crystallized kaolinite. Both varieties of tubular halloysite and poorly crystallized kaolinite exhibit poor viscosity, and their use is limited to fillers and ceramics. Well-crystallized kaolinite generally exhibits good viscosity properties and is suitable for high quality ceramics and paper coaters.

Most of the mineralogical work (Kelly, et al., 1963; Yuan, 1994; GMT, 2005) completed on the Helmer embayment clays indicates that the transported, sedimentary kaolins consist predominantly of kaolinite, but have a significant halloysite component. Yuan (1994) sampled clays from both the A.P. Green Refractories Company (A.P. Green) pit near Helmer and Simplot's Miclasil pits near Bovill. The main producing clay bed in the A.P. Green pit is 10 ft thick and includes several thin (1 to 6 in) interlayers of white to yellow tonsteins. The clay fraction in the main clay bed contains variable proportions of kaolinite and halloysite. Kaolinite abundance in the clay fraction ranges between 42 and 100%, while halloysite abundance ranges from 58 to 0%, respectively. Ginn Mineral Technology, Inc. (GMT) recognized only minor halloysite in a bulk sample from the same pit (GMT, 2005). The tonstein interlayers are generally all halloysite, the spheroidal halloysites that Yuan (1994) found to have low viscosities. Historically, Simplot mined sedimentary clay from their Miclasil pits west of Bovill for paper filler, but later switched production to residual clay pits. A.P. Green mined sedimentary clay from their pit north of Helmer for refractory brick.

Residual clays developed on weathered granitoid in the Helmer-Bovill area are a mixture of halloysite and kaolinite, with the concentration of each dependent upon the degree of weathering. The halloysite content increases with depth as the effects of weathering diminish (Yuan, 1994). He reported that kaolinite abundance can be as high as 100% of the clay fraction in samples taken near surface, while samples collected deep in old pits reach 100% halloysite. In tests on two samples from the WBL north pit, GMT (2005) demonstrated that there is a significant halloysite fraction in the residual clay. It is difficult to say where the samples discussed by Yuan (1994) occur within the weathering profile in this area. Historically, Simplot produced a filler clay for the paper industry from residual clay mined in the Moose Meadows area (Hosterman and Prater, 1964). The work done by GMT (2005, 2006) indicates that the quality of the residual clay from the WBL pit is high enough to be used in some high-end specialty uses in paper, paint, and ceramic markets. New technologies in kaolin processing have made further research into the high-grade markets worthwhile. However, work done by I-Minerals and further continued by GMT (2008) show that a wet process using cyclones can product a halloysite product that is sufficient to gain attention of halloysite markets.



8 Deposit Type 8.1 Mineral Deposit

The residual deposits consist primarily of K-feldspar, quartz and clays and comprising the mineral deposit of this report. The mineral deposit is underlain by the Thatuna Batholith, composed mainly of Na-feldspar, K-feldspar and quartz. Weathering has created a residual saprolite horizon which directly overlies the bedrock from which it was derived. During the natural processes of weathering, the original plagioclase feldspars have preferentially broken down to produce the clays kaolinite and halloysite. The K-feldspars have resisted weathering to a degree and much of the original component remains as free grains. Similarly, the quartz component of the host rock remains as free grains in the weathered material.

Minerals of economic interest include the following:

• Halloysite clay, an aluminosilicate with hollow tubular morphology in the submicron range; • Kaolinite clay, hydrated aluminum silicate; used in ceramics, rubber, plastics, etc and when

calcined becomes a metakaolin clay, or dehydroxylated kaolin clay, which is reactive (Pozzolan) and enhances the strength, density and durability of concrete and ceramics;

• K-feldspar, uniquely suited to ceramic formulations requiring an alumina source; and • Quartz, SiO2 silicon dioxide.



9 Exploration 9.1 Procedures and Parameters of Surveys and Investigations

During the period from 1999 through the end of 2001 the exploration work included the acquisition of over 6,000 ac of mineral lease applications, the compilation of an extensive file on the results of previous operations, and drilling.

During 2002 and 2003, geologic mapping and petrographic studies were performed. An electron microprobe analytical study was conducted on field samples, quartz products and feldspar products from earlier work. Following petrographic and microprobe studies, select intervals of residual deposits from the 2000-2001 drilling program were sent to Mineral Resource Laboratory (MRL) for process testing.

All exploration work completed on the property since 2003, has been diamond core drilling.

9.2 Sampling Methods and Sample Quality The field sampling described above consists of grab samples collected by digging with a shovel to below the soil horizon and placing the residual clay material into a sample bag. In previously mined locations, samples were collected directly into a sample bag by scraping with a trowel or hammer from freshly exposed residual clay horizons.

9.3 Significant Results and Interpretation The exploration work conducted by I-Minerals meets current industry standards.



10 Drilling 10.1 Type and Extent

During 2000-2001, a 41-hole diamond drill program was completed at the Project, focused on both bedrock feldspar deposits and residual deposits. Approximately 50% of the drillholes penetrated residual deposits at or very near the surface. A total of 4,063 ft. were drilled during this program. All holes were surveyed by Rim Rock Surveying. This work is described in two previous Technical Reports by Hodgson (2000) and Montgomery (2002).

In 2003, a 12-hole, diamond drill program was completed at the Project, testing for residual deposits over a broad area. A total of 1,333 ft. were drilled in this program. The core was split, sampled, and described in detail within a previous Technical Report by Clark (2004a) and in petrographic reports prepared for I-Minerals (Clark, 2003c; 2004b; 2004c; 2004d). All holes were surveyed with a hand held GPS with an accuracy of several meters.

In 2007, a 28-hole, diamond drill program was conducted to further evaluate the residual deposits. Six holes were located in the WBL Pit area on 200 to 600 ft spacing. The remaining holes were spread over the entire property to test those areas believed to be underlain by the weathered Thatuna granodiorite, establishing several new prospective areas. A total of 3,529 ft were drilled during this program. The six holes located at WBL Pit were surveyed by Jamar and Associates and all remaining holes were surveyed by handheld GPS with an accuracy of several meters.

In 2010, a 10-hole, diamond drilling program was completed in the WBL Pit and Middle Ridge areas. Five holes were completed in each area, on 400 to 900 ft spacing. A total of 1,195 ft were drilled in this program. All holes were surveyed by Taylor Engineering with a differential GPS to cm accuracy.

In 2011, a 66-hole, diamond drilling program was conducted in the WBL Pit and Middle Ridge areas. At Middle Ridge, 45 holes were drilled and at WBL, 21 holes were drilled. These holes were mostly located on 200 ft spacing with a few on 400 ft. A total of 7,747 ft were drilled during this program. All holes were surveyed by Taylor Engineering with a differential GPS to cm accuracy.

In 2013, a 167-hole, diamond drilling program was conducted in the Middle Ridge deposit and in two new areas referred to as Kelly’s Hump North and South. At Middle Ridge, 21 additional holes were completed to provide a drill pattern on 100 ft spacing in the area hosting higher halloysite grades. In the Kelly’s Hump area, a phase one program was completed with 17 holes spread though out the elevated area of the north south trending ridge. These were generally spaced at approximately 400-800 ft with all but one, located in the northern area. A phase two program was completed with 113 additional holes in the northern area and 16 in the south. The majority of drilling at Kelly’s Hump North is on 100 ft spacing. The drilling at Kelly’s Hump South is all on 200 ft spacing. A total of 17,811 ft. were drilled during this program. The drill hole locations were first laid out by Taylor Engineering with a differential GPS and then once the drill rig was set up any offsets were measured with a tape measure.

The drillhole database supporting the resource estimation of this report consists of 322 diamond core drillholes totaling 35,909 ft. (Figure 10.1.1). The shallowest hole is 20 ft, the deepest is 260 ft, and the average is 112 ft. All drillholes are oriented vertically and none of the holes have down hole deviation surveys. Since all of the drilling is relatively shallow the lack of down hole deviation survey



has no material impact on the sample location. Since many of the older drillholes are located with a hand held GPS their elevations do not match the current, high resolution topographic surface. For this reason, all drillhole supporting the resource estimation of this report, are draped onto the high resolution topography to provide a uniform basis of elevation control. Typically, the sample recovery was very good ranging from 60 to 100%. The average core recovery is 87%.

Source: SRK, 2013

Figure 10.1.1: Drillhole Locations and Resource Areas



10.2 Interpretation and Results The exploration drilling programs are all of appropriate type, they were well planned and carried out in a prudent and careful manner. All geologic logging and sampling has been done by trained and professional personnel. I-Minerals has made a concerted effort to ensure good sample quality and has maintained a careful chain of custody and ensured sample security from the drill rig to the assay laboratory.

The drilling has been conducted by reputable contractors using industry standard techniques and procedures. This work has defined zones of residual deposits derived from weathered granitoid overlying the Thatuna batholiths. These zones generally are continuous, following topography or lying sub horizontal down to an average depth of 70 ft below surface. The zones are thicker along ridges and thin toward the valleys. The drillholes are all oriented vertical and the deposits are interpreted to be sub-horizontal. Therefore, the drill intercepts do represent an approximate true thickness of the mineralization.

SRK is of the opinion that the drilling operations were conducted by professionals, the core was handled, logged and sampled in an acceptable manner by professional geologists, and the results are suitable for support of a NI 43-101 compliant resource estimation.



11 Sample Preparation, Analysis and Security 11.1 Sampling and Preparation Methods

Three types of samples have been collected from the study area to support this PFS. These include one sample type to support resource estimation and two types to support parts of the metallurgical test work. The resource estimation is supported by diamond drill core. Parts of the metallurgical test work are supported by hand dug channel samples and large bulk samples.

During collection of the drill core samples, the core barrel was removed from the hole and the core was allowed to slide into the core box, with the top of the interval at the top left of the core box. Poorly consolidated core was scraped with a sharp instrument and hard core scrubbed with a brush to remove adherent drilling mud from the core. The core boxes were labeled with hole number and footage interval on tops and bottoms. The core was transported to the I-Minerals' core facility near Moscow, Idaho at the end of every drilling shift. The core is all stored in a locked building prior to sampling. Once it has been logged and sampled it is moved to a locked core shed or a locked storage container. As part of the logging procedure, drill core was described in detail, and the descriptions were recorded on a standardized, hand written drill log form. A knife or chisel was used to split the core in half, and quarter-splits were made from one of the halves. In the 2007 and 2010 programs, one quarter-split in the visually clay-rich zones was bagged as a geochemical sample in intervals of uniform lithology that generally did not exceed 5 ft. The clay in the bag was crushed by hand and the bag was shaken up to thoroughly mix the sample. In general, sample intervals were 5 ft in length for the clay testing and 10 ft in length for whole rock geochemistry unless lithic contacts required a shorter interval. In the 2011 and 2013 programs, the one quarter-split is bagged and saved for clay testing in the laboratory at the University of Idaho. Sample intervals are no thicker than 5 ft down to 50 ft in depth and 10’ below that.

Two hand-dug channel samples of approximately 150 lb each, were collected from the North WBL Pit and a single sample of about 120 lb was collected from a pit in the southern portion of the property. These were collected as channel samples with pick and shovel from the face of the pit after the face was cleaned by scrapping with a hoe. The sample material was shoveled directly into 5 gal buckets lined with plastic bags. Bags were tied, and the buckets were sealed, palletized and shipped directly to the laboratory.

Two large bulk samples of residual deposit were collected from the North WBL Pit. In 2005, a 1.5t sample was taken and in 2007, a 2.0 t sample was taken. Both were collected by a Kobelco 905LC excavator with a 3 ft wide bucket. The pit face was scraped to expose fresh material prior to sampling. The excavator dug across the face, taking as much as possible in both vertical and horizontal directions. Once the bucket was filled, the material was hand shoveled into a 1-t super sack. These sacks are brand-new woven plastic bags that are constructed to handle heavy loads. The sacks were tied shut and shipped directly to the laboratory. In discussions with I-Minerals personnel, it is understood that these samples are not representative of the entire clay deposit, but they should give a good idea of the character of the material, such that a testing program can be designed.



11.2 Laboratories Analysis of the drill core to support the resource estimation was conducted at four laboratories. Whole rock analysis was completed at ALS Global (ALS) and material characterization studies were undertaken at Ginn Mineral Technology (GMT), a laboratory at a commercial clay operation(CCL) and at the University of Idaho (UOI).

Whole rock sample preparation and geochemical analysis was completed at ALS in Vancouver, British Columbia. The sample preparation was standard procedures to support the X-ray fluorescence (XRF) analytical method used. ALS is an ISO 9002 certified, international corporation and its analytical services are highly respected by the mining industry.

GMT is located in Sandersville, Georgia, in the heart of the Georgia kaolin belt. GMT is a technology-based company focusing on industrial mineral and base metal resources, fine particle process and product development, and the commercial application of minerals. GMT is the foremost independent kaolin process testing laboratory in North America. GMT is not ISO certified.

A commercial clay operation’s private laboratory (CCL) was used to determine recoveries of different size fractions and to obtain specific characteristics of the clay fraction. The laboratory itself is not ISO certified.

The University of Idaho’s Geological Engineering Department (UOI), located in Moscow, Idaho was utilized for particle characterization studies and scanning electron microscopy (SEM) for the majority of the samples supporting the resource estimation of this report. Some of the 2013 SEM work was completed at Washington State University in Pullman, Washington. The UOI is not ISO certified.

The laboratories described above are independent of the issuer.

11.3 Analysis Whole rock sample preparation and geochemical analysis was completed at ALS. The sample preparation was standard procedures to support the X-ray fluorescence (XRF) analytical method used.

The material characterization studies provide the primary support for the resource estimation of this PFS. This work involved two general areas of study including particle size analysis and clay characterization.

The particle size analysis is basically a screening/decantation process. First the material is weighed, slurred, run through an attrition scrubber and then washed over 50 and 325 mesh screens. The overflow materials from both represent the sand portion of the sample. The underflow material is then volume adjusted, run through a high speed mixer and washed over 500, 635 mesh screens. After the 325, 500, and 635 mesh screenings, multiple static settlings occur and the clay decants are collected. The decant samples collectively comprise the clay portion of the sample. The residual fractions from the 500 and 635 mesh screenings collectively comprise the waste portion of the sample. The mass balance of the three sample components, sand, clay and waste equals 100%. The specific procedures used by UOI are listed below.

• Obtain a 25g sample then dry to determine % moisture; • Obtain a 750g sample to be attrition scrubbed; • Use dry basis for calculations of material below;



• Add sample below to attrition scrubber vessel of D-12 Unit; o 450 grams of sample (dry basis), o 300 ml of D.I. water, and o 2 lb/t dispersant or 0.001 grams.

• Attrition scrub for 5 minutes; • Transfer contents from vessel to wet screen apparatus double-stacked with a 50 mesh

screen and a 325 mesh screen (placed on 5 gal bucket). Turn on vibration to unit and allow material screen. The 2011 samples were washed into the wet screen using a hand held squeeze bottle containing D.I. water. The 2013 samples were washed into the wet screen using a high pressure shower containing D.I. water;

• When 50 mesh top-screen appears clean of clay, remove it and continue rinse with 325 mesh screen. Stop screening when all clay appears to have passed through 325 mesh screen;

• Combine retained fractions from screens and dry. Determine dry weight; • Pour contents of bucket (-325 mesh fraction) into 2,500 ml beaker – rinse bucket with D.I.

water to insure all -325 mesh fraction has been removed and adjust beaker level to 1,000 ml volume with D.I. water;

• Using bench mixer, mix contents for 5 minutes at a fairly vigorous mix speed; • Pour mixed sample into 2,000 ml cylinder (use D.I. water squeeze bottle to remove all

contents). Allow to settle for 20 minutes; • Using a siphon tube, remove (decant) top layer of suspended clay fraction into a separate

beaker. There will be a distinct settled fraction that contains a predominantly -325+635 mesh material (+20 microns);

• Decanted fraction is adjusted to 1,000 ml volume in 1,000 ml cylinder using D.I. Water and transferred to 2,500 ml beaker for mixing. After 3 minutes draw 14 ml of sample using a syringe, then centrifuge drawn sample for 10 minutes and dry;

• Add some D.I. water to cylinder to remix settled fraction in cylinder and transfer back into a beaker. Wet screen this material over a 500 mesh screen;

• Retain fractions from screen and dry. Determine dry weight; • Pour contents of bucket (-500 mesh fraction) into 2,500 ml beaker, adjust volume to 1,000 ml

with D.I. Water and mix for 2 minutes. Transfer material into 2,000 ml cylinder and static settle for 20 minutes. Repeat decantation process and sampling steps as detailed above with 325 mesh screen;

• Using settled fraction from cylinder, repeat process again using 635 mesh screen; • Determine weights of all retained fractions and record; and • Prep decant samples for Scanning Electron Microscopy (SEM) analyses.

Clay characterization includes the differentiation between kaolinite and halloysite. This work was determined visually using the three clay decants secured from the process described above. The halloysite clay has a tubular shape while the kaolinite has a plate-like or blocky appearance.

The visual determinations were made using SEM technology. A small portion of each retained clay decant was prepped for SEM analysis by mounting each sample onto SEM wafers and coating with carbon. The prepped samples were then placed in the SEM and observed at 800X and 2000X magnifications. Representative photomicrographs were taken of each sample at each magnification.



Visual reviews for each laboratory processed drillhole interval were then performed and assigned a qualitative rating based on the amount of halloysite present in each respective sample. The key sample of interest for each interval was the -325 mesh decant (first decant), with the -500 mesh decant (second decant) of secondary interest. If present, the halloysite was primarily found in these two decants.

The relative ratios of kaolinite versus halloysite are visually estimated in each decant and then the entire sample is coded from 1 to 4. The lowest coding (1) has no halloysite present and the highest coding (4) has approximately 70%+ halloysite. Allocation of the halloysite and kaolinite quantification was then based on the clay coding parameters as described in Table 11.3.1.

Table 11.3.1: Clay Code Assignment Clay Code Clay Assignment

1 100% of all clay is assigned as kaolinite.

2 100% of the -325 mesh clay decant is assigned to halloysite and all remaining clay decant material is assigned to kaolinite.

3 100% of the -325 mesh clay decant is combined with 50% of the -500 mesh clay decant and assigned as halloysite, the remaining clay decant material is assigned as kaolinite.

4 100% of the -325 mesh clay decant is combined with 100% of the -500 mesh clay decant and assigned as halloysite all remaining clay decant material is assigned as kaolinite.

11.4 Security Measures I-Minerals has maintained a careful chain of custody throughout the sampling and transportation process. All samples have been bagged and closed immediately with tamper proof ties. Samples have always been transported by I-Minerals staff or commercial carriers. All sample storage has been within a locked facility.

11.5 QA/QC Procedures and Results I-Minerals has completed a program of QA/QC by ensuring that all samples have been collected using industry best practices, analyses were completed by reputable laboratories, a representative number of the samples have been subject to duplicate analysis at independent laboratories and standard reference material was submitted to the UOI laboratory. Certified reference material for this type of mineralization is currently non-existent so I-Minerals has created a non-certified reference material by using splits from bulk samples analyzed by pilot scale testing at GMT.

The duplicate analysis of the + 325 mesh and -325 mesh sample portions were completed during the 2011 and 2013 test work conducted at UOI. Because these two size fractions total to 100%, all results of the +325 mesh are inverse to the -325 mesh. For simplicity, only the -325 mesh results are presented and discussed. Figure 11.5.1 show a scatter plot of the 2011 UOI analyses versus GMT analyses. Although there is an expected amount of scatter, the duplicate analysis shows good correlation. Figures 11.5.2 shows a scatter plot of the 2011 UOI analyses versus CCL analyses. This plot clearly shows that the CCL has a bias towards the finer size fraction. This bias is not considered a material effect on the resource estimation since the CCL analytical data only represents the widely spaced drilling of the first exploration phase and because parts of this data were replaced by the data from the duplicate analyses. The duplicate analyses from the 2013 UOI testing were all sent to GMT. Figure 11.5.3 shows a scatter plot of the 2013 UOIanalyses versus the GMT analyses. This plot clearly shows that the 2013 UOI tests have a bias towards the finer size fraction.



Source: SRK, 2012

Figure 11.5.1: UOI vs. GMT Analyses for 2011,-325 Mesh Duplicates



Source: SRK, 2012

Figure 11.5.2: UOI vs. CCL Analyses for 2011,-325 Mesh Duplicates



Source: SRK, 2012

Figure 11.5.3: UOI vs. GMT Analyses for 2013, -325 Mesh Duplicates

The UOI also conducted testing of the non-certified reference material. These were run over the entire program of testing, resulting in 51 tests by 18 different lab technicians. The results are shown in Figure 11.5.4. Here again, the UOI tests clearly show a bias toward the finer size fraction.



Source: SRK, 2012

Figure 11.5.4: UOI vs. Non Certified Reference Material for the -325 Mesh, 2013 Analyses

The significant change in UOI’s laboratory practice from 2011 to 2013 was the addition of the high pressure shower used to complete the initial -325 mesh wet screening. This change has clearly impacted the comparison between the two laboratories. Since the entire pilot scale metallurgical test work and associated analyses were completed at GMT and since this data supports the technical economic model, the GMT results were accepted as more accurate.

To overcome UOI’s 2013 bias towards the finer size fraction, all of the 2013 size fraction analyses from UOI were adjusted to remove the bias. This was done by first determining the average bias. This was taken to be the average difference between all of the 2013 UOI -325 mesh samples and the GMT results for the same. A total of 121 tests were averaged to show that UOI reported 6.5922%



more -325 mesh material than GMT. The database of the 2013 UOI samples was factored by this amount, as follows. The relative proportion of the clay waste, halloysite and kaolinite constituting the -325 mesh material was determined. Each of these components was then factored down by its weighted proportion of the 6.5922% bias. The remaining +325 mesh, sand fraction was then factored up by 6.5922 to provide a 100% mass balance in the sample. Figures 11.5.5 and 11.5.6 show the results of the UOI vs. GMT duplicates and reference samples, respectively, with the 2013 bias factored out.

Source: SRK, 2013

Figure 11.5.5: UOI vs. GMT Analyses for 2013, -325 Mesh Duplicates, with Bias Removed



Source: SRK, 2013

Figure 11.5.6: UOI vs. Non Certified Reference Material for the -325 Mesh, 2013 Analyses with Bias Removed

11.6 Opinion on Adequacy SRK is of the opinion that the sampling work conducted by I-Minerals and the analytical work performed by the laboratories discussed above is valid and suitable for use in resource estimation. The sample characterization studies used are industry accepted analytical techniques used to determine particle size distributions in exploration samples. The QA/QC program employed by I-Minerals meets current industry best practices and the results of this work indicate acceptable precision and accuracy of the analytical results.



12 Data Verification 12.1 Procedures

The exploration database delivered to SRK by I-Minerals consisted of an Access database, an Excel spreadsheet and SEM photomicrographs.

The Access database was originally created by Dr. Mark Groszos of Valdosta State University in Georgia and subsequently modified by UOI, The database contained tables describing the drillhole collar locations, drillhole orientations, lithology intervals and descriptions, analytical sample intervals, XRF assay data, and various laboratory’s material size classification data generated prior to 2012.

The Excel spreadsheet was constructed jointly by I-Minerals and UOI. It is titled “Master Data Summary” and contains all drill sample intervals, material size classification data and clay identification information. The Master Data Summary was originally constructed in 2012 and was subsequently updated in 2014 with additional data.

The SEM photomicrographs are arranged in electronic folders by drillhole, sample interval, size fraction decant sequence, SEM magnification and photo number. The SEM photomicrographs are in .tif file format.

The drill collar locations were verified by comparing original layout maps and coordinate sheets with the Access collar tables. The drillhole collars have been surveyed in Idaho State Plane (ISP) coordinate system by a licensed surveyor. The drill collar locations were further verified by comparing original surveyor coordinate data to the coordinates in the Access database. In addition, the original drillhole layout maps were compared to maps derived from the collar locations in the Access database. All drillholes are oriented vertical so verification of drillhole azimuth and inclination is not required.

The material size characterization data in the Master Data Summary was originally verified in 2012, by comparing information from the original, independent laboratory data files to the same records in the Excel spreadsheet. At that time a total of 112 records representing 17% of the total database were checked. A few minor errors were corrected.

UOI provided a Master Data Summary which only includes data generated since the original 2012, verification. Since the 2012 verification, UOI has been entering their results directly into the Master Data Summary. Because of this, UOI does not generate any other assay certificates or data files from which to verify the Master Data Summery. SRK believes that UOI staff has assembled the data with utmost regards to accurate transfer and data entry. SRK did conduct additional verification on the Master Data Summary by verifying from-to intervals and mass balance results. A total of 295 samples were verified representing 19% of the total database. Twenty seven errors were detected and corrected but overall the database represents the actual data very well.

The clay characterization codes were verified by viewing the SEM photomicrographs and visually estimating the proportion of kaolinite versus halloysite clay. This is essentially in a parallel routine to the methods used by I-Minerals. A total of 548 records, representing 35% of the total data were checked. A total of 213 discrepancies were noted and addressed. The discrepancies were related to clay type assignment codes for intervals which did not have SEM photomicrographs or were missing



size fraction data. I-minerals corrected all of these intervals and the database was accepted by SRK as suitable to support the current resource estimation.

12.2 Limitations SRK was not limited in its access to any of the supporting data used for the resource estimation or describing the geology and mineralization in this Technical Report.

The database verification is limited to the procedures described above. All mineral resource data relies on the industry professionalism and integrity of those who collected and handled it.

12.3 Opinion on Data Adequacy SRK is of the opinion that best professional judgment, and appropriate exploration and scientific methods were utilized in the collection and interpretation of the data used in this report. The sampling data is sufficient and spaced appropriately to support the resource estimation. However, users of this report are cautioned that the evaluation methods employed herein are subject to inherent uncertainties.



13 Mineral Processing and Metallurgical Testing I-Minerals retained GMT to perform testing on the Bovill Kaolin deposit as it pertained to the recovery of the clay portion of the deposit. I-Minerals retained the Minerals Research Laboratory (MRL) at North Carolina State University (NCSU) in Asheville, North Carolina, to perform a series of pilot plant tests on K-spar sand tailings as it pertained to the recovery of feldspar and quartz of the deposit. The recovery processes supporting the PFS of the Bovill Kaolin deposit incorporate both of these testing procedures as discussed below.

The GMT facility is located at 150 International Drive, Sandersville, Georgia, USA 31082. GMT is a technology based company focused on industrial minerals and with demonstrated expertise in fine particle processing, separation, and dewatering technologies, surface chemistry, and thermal processing of kaolin clays. GMT provided independent professional testing services to I-Minerals as described below.

The MRL Pilot Plant is located at 180 Cox Avenue, Asheville, North Carolina, USA 28801. The MRL is an extension facility of NCSU College of Engineering focused on industrial minerals and with demonstrated expertise in research, development, and implementation of industrial minerals processing techniques. Due to its North Carolina location, the MRL is particularly knowledgeable in the beneficiation of both feldspar and quartz minerals.

The following information will discuss the primary clay testing first followed by the feldspar and quartz testing.

13.1 Testing and Procedures

13.1.1 Initial Testing of Primary Clay GMT performed testing on two WBL pit primary clay samples (WBL-1 and WBL-2) from the Bovill Kaolin deposit and one sedimentary clay sample (JH05) from another deposit area. The initial work was documented in a report titled Kaolinite/Halloysite Clay Evaluation - July 12, 2005 (GMT, 2005). For purposes of this report, this work will be referred to as Phase I. For reasons of simplicity and relevance, the sedimentary clay sample will be discussed on a very limited basis.

GMT initially performed an evaluation to characterize the clay samples for physical, chemical, and mineralogical properties. Tests were also performed to determine the responsiveness of the clay to commercial processing technology. A second test program, Phase II, was performed by GMT on another WBL primary clay sample as well as the sedimentary clay sample due to the successful results of the initial work. The purpose of this second phase evaluation was:

• Optimize processing parameters as identified in the initial testing phase and evaluate advanced/specialized brightness processing options;

• Produce larger samples for I-Minerals; and • Determine potential end uses for the kaolin products as well as potential co-products from

the +325 mesh separated fraction.

This work was documented in a report titled Kaolinite/Halloysite Clay Evaluation Phase II – February-March, 2006 (GMT, 2006). Subsequent reference to this work in this report will be as Phase II. As in



the Phase I discussion, the sedimentary clay sample will not be discussed in the results of this report.

Primary Clay Test Procedures

Phase I

In the initial evaluation of I-Minerals clay samples, three clay samples were provided by I-Minerals. The samples consisted of two WBL ore body samples and one sedimentary clay sample. Using conventional kaolin beneficiation techniques, approximately 100 lb of each sample were processed and evaluated (Figure 13.1.1.1).




Figure 13.1.1.1: Helmer-Bovill Clay Processing Plant Phase I Testing Block Flow Diagram

WATER + SODIUM POLYACRYLATE DISPERSANT

" AS RECEIVED" KAOLINITE CLAY

+325 MESH FELDSPATHIC SAND

H2SO4 + Na2SO4

H2SO4 + Na2SO4

BLUNGER

HORIZONTAL SWECO SCREEN

COARSE PIGMENT(-325 MESH) CLASSIFIER

FINE PIGMENT(80% -2μ)

SUPERCONDUCTING MAGNETIC SEPARATOR

AGITATED TANKLEACHING


AGITATED TANKLEACHING

FILTRATION

FILTRATIONDRYER

DRYER

HIGH-BRIGHTNESS COARSE PIGMENT

HIGH-BRIGHTNESS FINE PIGMENT



The steps used to process the samples as received included:

• The crude samples were individually blunged at approximately 60% solids with 1 lb/t sodium polyacrylate dispersant;

• The dispersed materials were screened at 325 mesh using a horizontal Sweco screen to remove the +325 mesh feldspathic sand from the clay fraction;

• For each sample a portion of the -325 mesh sample was separated and designated coarse pigment, while the remainder was classified to an 80% -2µ particle size and designated fine pigment;

• Each pigment fraction was then processed through a super conducting magnet and leached (H2SO4 and Na2S2O4). They were then filtered and dried; and

• The various products, including the crude, were then evaluated. Chemical and mineralogical characterization of the clays included X-ray Fluorescence (XRF), X-ray Diffraction (XRD), Scanning Electron Microscopy (SEM), and Cation Exchange Capacity (CEC). The physical and optical characterization of the clays included Particle Size Distribution, Viscosity at low and high shear, Abrasion, Tappi Brightness and Whiteness, and color L-a-b values.

Phase II

The steps used to process the samples as received included:

• The samples were processed according to the diagram shown in Figure 13.1.1.2; • The -325 mesh fractions were classified to 90% -2µ particle size and designated ultra-fine

pigment, while the oversize fractions from the previous classification were delaminated to 80% -2µ particle size using bead milling and designated fine pigment;

• WBL pigment fractions were then processed through a Super Conducting Magnet (SCM), leached, filtered, and dried. The majority of the JH05 sample was thermally processed (calcined at approximately 875°C) to produce a metakaolin sample;

• Additional advanced or specialized processes were also performed and analyzed on the samples after classification to -325 mesh. These processes included froth flotation, selective flocculation, and selective separation, and then followed by SCM treatment; and

• The various products and crude were then evaluated as in Phase I. The +325 mesh co-products of quartz and K-spar were also characterized. These evaluations included mineral analysis (XRF/XRD) and PSD.




Figure 13.1.1.2: Helmer-Bovill Clay Processing Plant Phase II Testing Block Flow Diagram

WATER + SODIUM POLYACRYLATE DISPERSANT

" AS RECEIVED" KAOLINITE CLAY

+325 MESH FELDSPATHIC SAND

-325 MESH OVERSIZE

UNDERSIZE

H2SO4 + Na2SO4 H2SO4 + Na2SO4

BLUNGER

HORIZONTAL SWECO SCREEN

BEAD MILLING(DELAMINATION) CLASSIFIER

ULTRA-FINE PIGMENT(90% -2μ)


(20% SOLIDS)

AGITATED TANK


AGITATED TANK

FILTRATION FILTRATION

DRYER DRYER

METAKAOLIN

HIGH-BRIGHTNESS ULTRA-FINE PIGMENT

FINE PIGMENT(80% -2μ)

PULVERIZER

CALCINER (875 oC)

PULVERIZER


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 52 13.1.2 Mechanical Testing of Hydrocyclones

The results of the GMT primary clay tests demonstrated there was both halloysite and kaolinite present in the WBL primary clays. Based on the clay separation characteristics, gravitational testing was implemented to investigate a satisfactory method for separating the halloysite and kaolinite. Based on experience and research, hydrocycloning was selected as potential separation method. FLSmidth Krebs, an industry leader in hydrocycloning technology, was contacted for testing. They are located at 5505 West Gillette Road, Tucson, Arizona, USA 85743.

In 2008 FLSmidth Krebs tested five 5-gal buckets of -325 mesh clay slurry originated from WBL crude ore. The cyclone tests evaluated three different cyclone sizes for their effect on fine grit and clay separation. One cyclone evaluated fine grit while the other two evaluated separation of the kaolinite and halloysite (FLSmidth Krebs, 2008).

Hydrocyclone Test Procedures

Grit Separation – Test 1

A 3 inch diameter cyclone was mounted in conjunction with a sump, mixer, and pump closed-loop circuit. Sample was transferred into the sump and the circuit operated to evaluate the overflow and underflow splits. Particle size, percent solids, flow rates, and yields were determined.

Clay Separation – Tests 2 and 3

Using the same circuit and procedure described above, 1 inch and 0.5 inch diameter cyclones were evaluated. Particle size, percent solids, flow rates, and yields were determined.

Representative samples of each fraction separated during the three tests were sent to other laboratories for additional analysis. SEM photomicrographs were produced through GMT, and chemical data results were obtained through ALS Minerals Laboratory. All data results were then compiled and evaluated.

13.1.3 Pilot Plant Testing of Primary Clay A sample of primary clay from the north WBL pit of approximately 3,000 lb was provided to GMT for pilot plant testing. The primary focus of the pilot plant evaluations was to separate the clay from the feldspathic sand and then concentrate the halloysite from the kaolin clay. Two tests were performed and reported in July 2008 and July 2010. Three additional smaller scale pilot plant testing followed up the July 2010 testing. The results from these smaller tests were provided to I-Minerals for review, but no formal reports were provided.

Primary Clay Pilot Plant Test Procedures

GMT processed approximately 500 lb of crude clay in 2008 for the purpose of evaluating the efficiency of the cyclone separations (GMT, 2008). The flow sheet included the following:

• The crude clay was initially blunged and dispersed with sodium polyacrylate at about 50% solids and then screened at 325 mesh to remove the feldspathic sand from the clay fraction;

• The screen passing fraction was re-mixed and pumped through the 3 inch cyclone. The underflow fraction was considered fine grit or waste. The overflow clay fraction was captured and used for subsequent fine hydrocycloning separations; and



• The overflow clay fraction was mixed and pumped through a 0.5 inch cyclone. The overflow and underflow fractions from this first pass were collected and a portion of the overflow mixed and fed through the 0.5 inch diameter cyclone in a second pass. The products, consisting of first and second pass overflows and underflows, were then dewatered for product evaluations. The overflow fractions were representative of concentrated halloysite, and the underflows were considered concentrated kaolinite.

GMT processed the remaining 2,500 lb and reported the work in July 2010 (GMT, 2010). The flow sheet was similar except for the following:

• Two 0.5 inch cyclones were used in parallel (vs. one cyclone) for each separation pass and higher feed solids were also used in an effort to expedite the processing time due to the amount of material;

• A small amount of second pass overflow was passed through the 0.5 inch cyclones to produce a third pass product for SEM examination; and

• A majority of the combined underflow fractions from the 0.5 inch cyclone passes, or kaolinite clay, was dewatered and then calcined to produce a metakaolin product.

The three additional small sample programs included:

• A test was performed in August 2010 due to high overflow recovery and less distinctive SEM visual separation in the July 2010 test. Some of the dry second pass material product was slurried at lower solids and cyclone;

• Another small 300 lb of bulk sample was provided and tested in October 2010. Procedures mirrored the July 2008 work; and

• A small 250 lb of core composite sample was provided and tested in October 2010. Procedures mirrored the July 2008 work.

13.1.4 Pilot Plant Testing of WBL K-Spar Tailings The I-Minerals’ Helmer-Bovill Property feldspar deposit contains both residual feldspar resources as well as potassium-based feldspar with clay tailings from a previous operation. I-Minerals retained MRL to perform a series of pilot plant tests on K-spar tailings. Work was also performed on a composite sample of soda feldspar diamond drill cores for feldspar and quartz recovery but is not discussed.

The pilot testing program had two major objectives:

• Demonstrate the technical feasibility of the overall process consisting of an integrated series of unit operations or process steps, and

• Generate scale-up data, process design criteria, and functional equipment duty specifications to assist in the design of a commercial processing facility.

The MRL, under the direction of Mr. John W. Schlanz, Chief Engineer, completed pilot plant testing of I-Minerals’ WBL K-spar tailings and this work was documented in a report titled Pilot Plant Testing of I-Minerals Moose Creek K-spar Tailings – August 7, 2009. These tests successfully proved:

• The technical feasibility of the process on a continuous basis by processing approximately 35 t of K-spar tailings as feed and producing, by optimizing the unit operations, tonnage of feldspar and quartz concentrates.



The secondary objective was also met in support of commercial-scale process plant design.

It should also be noted that prior to the pilot testing program several bench-scale tests were conducted by the MRL. These tests to identify candidate unit operations were conducted to verify initial mineral response and recovery using conventional beneficiation techniques for feldspar and quartz, develop a basic processing flow scheme, and determine operating variables.

Bench tests were performed and reported in 2002 to 2003 for the K-spar using stockpile composite samples and residual ore face samples.

K-Spar Tailings Test Procedures

On June 26, 2006, the MRL took receipt of 35 t of tailings packaged in super sacks. The tailings were from the WBL tailings stockpile and are residues of material previously processed for clay recovery from the WBL pits (Section 6) and required only air drying as feed preparation to the pilot testing program. The configuration of the flowsheet for the beneficiation pilot plant included the following unit operations:

• Automated weigh belt feeder to maintain constant feed rate with a recorder to monitor the weight of feed to a 36 inch diameter Sweco screen to scalp the feed to the pilot plant at 10 mesh;

• The +10 mesh screen oversize reported to an 18 inch diameter x 36 inch long rod mill; • The -10 mesh screen undersize was pumped to a CFS (Classification and Flotation

Systems, Inc.) dense media hydrosizer set at 30 mesh size cut, with the underflow directed to the rod mill, which operated at 50 wt% solids;

• The rod mill discharge was sent to the Sweco screen for a closed-loop grinding operation; • The -30 mesh hydrosizer overflow was pumped to the #1 Cyclone for initial desliming, and

the #1 Cyclone underflow reported to the #1 Screw Classifier for additional desliming and dewatering;

• The Screw Classifier discharge was fed to the Attrition Scrubber equipped with three 3 hp (2.25 kW) motors. The feed to the attrition scrubber was maintained at 60 wt% solids and conditioned with 0.4 to 0.5 lb/t of sodium hexametaphosphate [(NaPO3)6] to assist in slime dispersion;

• After conditioning, the Attrition Scrubber discharge was diluted to approximately 10 wt% solids and pumped to the #2 Cyclone for desliming before the 30 x 200 mesh #2 Cyclone underflow slurry was dewatered in the #2 Screw Classifier;

• The #2 Screw Classifier transported the washed pulp to the agitated mica Conditioning Tank in which the 35 to 40 wt% solids slurry was conditioned with sulfuric acid for pH control, the amine collector, and #2 fuel oil in preparation for mica flotation;

• The conditioned slurry was discharged to the mica flotation circuit, which consisted of two Denver D-7 stainless steel flotation cells for removing mica as the froth product. A small amount of frother was added during flotation to assist in the mica removal;

• Tailings from the mica flotation cells consist of a mixture feldspar and quartz. It was discharged for dewatering in the #3 Screw Classifier before conditioning for feldspar flotation;

• The spar conditioning consisted of mixing the pulp at 40 wt% solids with hydrofluoric acid (HF), amine, and #2 fuel oil and adding small amounts of frother to the spar flotation cells;



• The feldspar Rougher Flotation circuit consisted of a bank of three Denver D-7 Flotation Cells for feldspar float recovery with quartz reporting as rougher tails;

• The spar rougher float product was gravity fed to two Denver D-7 Flotation cells in the Cleaner Flotation circuit. The froth product is the flotation concentrate and the cleaner tails are the “spar middlings;”

• Both spar and quartz were dewatered on 200 mesh Sweco screens and collected in 55 gal drums. In final runs, a 50 mesh 36 inch Sweco screen was added to the circuit to recover additional spar from the middlings stream;

• Feldspar, quartz, and middling products were dried in a stainless steel Rotary Steam Tube Dryer;

• Any remaining mica (magnetic impurities) in the product(s) was removed in three-pass magnetic separation using the triple roll rare earth (permanent) magnetic separator from Outotec, Inc.;

• Four to five sampling rounds were conducted, typically one hour apart, during each test run. Product streams sampled included the feldspar cleaner product, the feldspar cleaner middlings, the feldspar rougher float product, the quartz rougher tails, mica float product, and all slimes products consisting of Cyclone and Screw Classifier overflows;

• At the end of each test run, samples were collected around the grinding circuit to determine the performance of the Hydrosizer and to evaluate the efficiency of the grinding operations (Rod Mill);

• Samples of discharges were also collected from the Attrition Scrubber and the Conditioner to determine flow rates for equipment specifications. These samples, however, were collected at the end of test run to avoid disrupting the material flow and balance of the Pilot Plant operations; and

• All timed samples were dewatered, dried, and weighed for calculating the mass balances for each test run.

13.1.5 Bench-scale Testing of K-Spar Flotation Tailings for Quartz I-Minerals retained MRL to perform a series of bench-scale tests from the Bovill Kaolin deposit to evaluate the potential of high grade quartz. Feedstock for this test was residual K-spar material (feldspathic sand cut) from the Bovill Kaolin deposit that had been separated by GMT earlier.

The testing program included the following requirements and objectives:

• Initiate K-spar recovery using previously determined process schematic, and • Produce value added quartz products of varying grades employing various flotation steps.

The MRL, under the direction of Mr. John W. Schlanz, Chief Engineer, completed this bench testing program and successfully produced three grades of high pure quartz. This work was documented in a report titled Recovery of High Grade Quartz from I-Minerals WBL Ore Body – September 14, 2011 (Schlanz, 2011).

It should also be noted that prior to this testing program a limited series of bench-scale tests were conducted by the MRL in 2006 on Kelly’s Basin Na-spar tailings to investigate the potential for high purity quartz. No formal report was issued on this work but the data results demonstrated a successful production of high purity quartz grades.



Bench-scale Quartz Testing Procedures

Standard beneficiation techniques were used and unit procedural operations were as follows:

• Due to a high amount of clay still present in the sample, a combination of wet and dry 200 mesh screening was used to remove additional -200 mesh fines;

• Grinding was performed in an 8 inch x 12 inch stainless steel (SS) rod mill charged with SS rods. Grinding was performed at 50% solids and done in closed circuit to minimize fines generation. Grinds included 30 mesh for the initial K-spar recovery and 50 mesh for the quartz upgrading;

• Desliming was performed at low solids using a 30 second settling period followed by decantation through a 200 or 140 mesh screen;

• Attrition scrubbing was accomplished using an octagon-shaped cell and a triple-blade impeller. The retention time was set at 5 min using 70 to 75% solids, and was followed by desliming at 200 mesh;

• Conditioning was performed in 2,000 ml beakers at 50 to 60% solids. Flotation reagents were added at predetermined rates using three flotation schemes: (1) iron/mica flotation using sulfuric acid for pH control and amine as the mineral collector (prior to feldspar flotation), (2) iron flotation using sulfuric acid and petroleum sulfonate collector (prior to quartz upgrading flotation), and (3) standard hydrofluoric acid and amine collector (for both feldspar recovery and quartz upgrading). Conditioning was conducted at 2.5 to 2.7 pH and retention times ranged from 3 to 4 minutes; and

• Flotation was performed in a standard Denver lab cell at 20 to 25% solids and included the addition of small amounts of frother reagent in all steps.

13.1.6 Pilot Plant Testing of K-Spar Flotation Tailings for Quartz In order to verify the results of the bench-scale quartz testing on a larger scale, collect data for proper scale-up to production, and obtain sufficient quantities of quartz upgraded products for market and product development, the MRL was again retained to perform additional pilot plant testing. This work was secured to simulate the successful bench-scale work just completed but on a pilot scale. Feedstock for this test was tailings fraction (quartz circuit feed) retained from the pilot plant testing of WBL K-Spar Tailings as described in Section 13.1.4.

This work commenced in late 2011 and was completed in January 2012.

13.1.7 Repeat Bench-Scale Testing of K-Spar Flotation Tailings for Quartz It was determined that additional bench-scale testing for quartz qualitative comparison analysis was needed utilizing both WBL K-spar Tailings material and composite core K-spar material from the Bovill Kaolin deposit. I-Minerals retained MRL to perform a series of bench-scale tests to recover upgraded quartz products from these two feedstocks. Procedures previously documented were utilized to produce three quartz products using single, double, and triple flotation schemes.

The work commenced in May 2012 and was completed in July 2012.



13.2 Relevant Results

13.2.1 Results and Conclusions Based on Primary Clay Tests Primary Clay Test Results

Phase I

Initial primary clay test results from the July 2005 tests:

• A crude clay characterization was performed on the WBL-1 and WBL-2 samples, and results summarized in Table 13.2.1.1. Also shown in the table are the analyses for the WBL-1 and WBL-2 samples after processing;

• Good brightness values were achieved with extensive processing methods; and • SEM images from the crude and processed materials for each sample illustrated the range

in particle size and particle morphology of the minerals and showed the presence of the halloysite with the kaolinite.

Phase II

Results from the February to March 2006 tests:

• Similar characterizations as performed in Phase I were completed and shown in Table 13.2.1.2. The table also shows the analyses for the WBL sample after processing. Results were similar compared to the Phase I work;

• The products produced from the crude samples concentrated the clay (halloysite and kaolinite) to almost 100% and removed the majority of other mineral contaminants; and

• Good brightness values were achieved with extensive processing methods.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 58 Table 13.2.1.1: Phase I Characterization Data

Sample Chemical Analysis Data Physical Properties Brightness/Color Properties

MgO SiO2 Al2O3 K2O Na2O FeO TiO2 LOI PSD @ % µ Passing Viscosity, cps (1) Abrasion (2)

Tappi Brightness L a b 10 5 2 1 Low Shear High Shear % Solids

WBL-1

Crude 0.35 60.43 35.91 2.23 0.25 0.83 5.53 -325mesh Product 86.1 66.4 34.1 21.0 75.31 91.38 0.79 6.84 Classified Product 99.5 98.6 80.6 56.8 77.90 92.60 0.96 6.51 Classified Oversized 78.1 44.9 7.7 3.7 71.13 89.61 0.54 7.74 Product A (3) 14.05 90.3 76.2 51.3 39.0 160 @ 10 1540 @ 18 63.0 6.7 84.85 94.79 -0.25 4.27 Product B (4) 56.70 43.30 15.06 99.5 99.0 82.9 63.6 390 @ 20 620 @ 18 63.0 0.8 93.01 97.27 -0.20 1.73 Product C (5) 0.03 55.76 39.19 0.75 0.26 3.22 0.80 12.52

WBL-2

Crude 60.30 34.75 2.91 2.01 0.03 6.39 -325mesh Product 85.2 71.3 49.7 37.8 67.09 87.47 1.41 8.12 Classified Product 99.2 96.8 79.6 65.5 70.65 89.56 3.06 8.09 Classified Oversized 68.7 33.8 8.3 5.4 61.77 85.39 2.05 9.69 Product A (3) 14.17 92.4 82.3 66.4 56.1 210 @15 1200 @ 15 63.0 9.4 85.42 95.08 -0.11 4.26 Product B (4) 57.43 41.88 0.69 14.80 99.5 98.6 84.9 73.4 480 @ 14 520 @ 18 63.0 0.6 90.51 96.67 -0.17 2.70 Product C (5) 55.36 38.29 2.99 3.35 15.84

Source: I-Minerals 1. Viscosity: Low Shear = cps @ dispersant lb/t and High Shear = rpm @ dynes 2. Abrasion @ 43,500 revolutions 3. Final Product (-325 mesh), magnet and leach processed 4. Final Product (Classified to 80% -2µ), magnet and leach processed 5. Magnet Rejects from Product B


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 59 Table 13.2.1.2: Phase II Characterization Data

Sample

Chemical Analysis Data Physical Properties Brightness/Color Properties

MgO SiO2 Al2O3 K2O Na2O FeO TiO2 LOI PSD @ % µ Passing Viscosity, cps (1)

Abr (2) Tappi

Brightness L a b 10 5 2 1 Low

Shear High

Shear % Solids

WBL

Crude 0.24 60.32 35.39 2.54 1.35 0.16 3.95 -325mesh Product 86.4 73.0 50.8 36.4 64.09 86.03 0.15 8.65 Ultrafine Product (90%-2µ)

Classified Product 73.34 90.71 1.86 7.46 Magnet Product 84.30 94.92 0.35 4.86 Leached Product 56.61 42.38 0.19 0.82 14.65 99.3 97.0 89.0 73.5 [email protected] 320@18 63.0 91.18 96.75 -0.25 2.32

Classified Oversized 67.03 88.15 1.12 9.04

Fine Product (80%-2µ)

Grinder Product 67.12 88.33 0.87 9.20 Magnet Product 78.14 93.06 -0.09 6.99 Leached Product 57.33 39.30 2.10 1.26 13.37 99.0 97.4 80.0 54.2 [email protected] 220@18 61.3 82.10 94.15 -0.64 5.48

SSP (3) SSP Product 71.21 89.58 1.06 7.63 SSP Magnet Product 84.06 94.87 0.41 4.96

SFP (4) SFP Product 67.46 88.12 0.67 8.67 SFP Magnet Product 73.47 90.90 0.73 7.64

FP (5) FP Product 67.05 87.97 0.69 8.80 FP Magnet Product 78.07 92.80 0.39 6.66

JH05

Crude 0.23 55.39 39.39 0.89 1.50 2.62 12.97 -325mesh Product 91.9 81.3 64.6 52.8 Classified Product 64.08 85.30 0.91 7.61 Classified Oversized 59.52 84.18 0.91 9.94 Metakaolin 56.19 40.13 0.52 1.48 1.68 1.10 85.3 68.5 41.5 27.3 [email protected] 4400@8 62.8 78.79 92.87 0.55 6.30

Source: I-Minerals 1. Viscosity: Low Shear = cps @ dispersant lb/t and High Shear = rpm @ dynes 2. Abrasion @ 43,500 revolutions 3. Selective Separation Processing 4. Selective Flocculation Processing 5. Flotation Processing



Primary Clay Test Conclusions

Phase I

• The WBL samples contained predominately quartz with varying amounts of potassium feldspar, kaolinite, halloysite, and muscovite mica;

• The processing of these samples was consistent with processing other primary clay deposits and technically proved the feasibility; and

• The clay generated from these two samples has excellent brightness and color indices, thus lending itself applicable to the paper, paint, and ceramic industries.

Phase II

• The WBL sample is a combination of kaolinite and halloysite clays once the +325 mesh material is removed. The primary discoloring impurities are iron and titanium oxides within the dispersed clay water system;

• The processed clay from the sample responded well to conventional brightness improvement processes such as magnetic separation and leaching;

• The final clay products produced from the sample resulted in a unique combination of kaolinite and halloysite clay, and are relatively pure in composition. Additional product applications may include plastics, pharmaceuticals, and nano-composites;

• The majority of the +325 mesh fraction from the sample is quartz sand with a significant amount of potassium feldspar. The quartz and K-spar could be processed as co-products, diversifying the WBL clay resource; and

• The JH05 sample processed to a metakaolin was successful and noteworthy as future work will include metakaolin processing of the WBL ore.

As noted above, a significant discovery during the evaluations was evidence of the presence of halloysite in the samples. Scanning Electron Microscopy photomicrographs confirmed the presence of the halloysite clay. This unique form of kaolin was found as a film floating on the surface of a kaolin slurry sample, indicating a gravimetric difference in the physical attributes of the two clays. I-Minerals subsequently concluded, based on the gravimetric separation characteristics of the kaolinite and halloysite clays, which specific gravimetric separation testing techniques needed to be evaluated.

13.2.2 Results and Conclusions Based on Mechanical Testing of Hydrocyclones Hydrocyclone Test Results

Particle size data for the Test 1 grit separation revealed excellent separation of fine grit in the underflow and contained primarily -325+635 mesh particle size material. Chemistry data also confirmed the overflow fraction contained principally clay. Test 2 did not produce successful clay separations based on process data and SEM’s. Test 3 did produce clay separations based on the process data and SEM’s.

Hydrocyclone Test Conclusions

The conclusions from the test work performed at the FLSmidth Krebs facility were developed after all the analytical work was received from Krebs, GMT, and ACT Laboratory. The 3 inch cyclone proved to be suitable for the removal of fine grit. The 1 inch cyclone was not shown to provide an adequate



separation of the halloysite and kaolinite clays. The 0.5 inch cyclone demonstrated that a shape separation was indeed possible using cyclones. Due to these test results, one 3 inch and three 0.5 inch diameter cyclones were purchased and sent to the GMT facilities for additional testing.

13.2.3 Results and Conclusions Based on Pilot Plant Testing of Primary Clay Primary Clay Pilot Plant Test Results

Discussion of results is centered on the recovery and separation of the halloysite and kaolinite clays. Additional results, such as clay brightness data and chemistry, are not discussed.

Hydrocyclone Evaluation July 2008

The screened -325 mesh clay fraction yield was 28.88% and considered average for this deposit (average at 30%). The 3 inch cyclone successfully split the grit fraction from the clay. SEM’s shown in Figures 13.2.3.1 and 13.2.3.2 demonstrate the separation. Recovery splits through the 0.5 inch cyclone were very similar and deemed acceptable.




Figure 13.2.3.1: SEM of 3" Hydrocyclone Overflow Showing Clays




Figure 13.2.3.2: SEM of 3: Hydrocyclone Underflow Showing Feldspathic Sand Grit

Of most importance, the SEM’s illustrated excellent separation and concentration of the halloysite in the second pass overflow fraction using the 0.5 inch diameter cyclone. The underflow fractions contained the highest concentration of kaolinite clay. SEM’s shown in Figures 13.2.3.3 and 13.2.3.4 illustrate the clay fraction separations.




Figure 13.2.3.3: SEM of 0.5" Hydrocyclone 2nd Pass Overflow (Predominantly Halloysite)




Figure 13.2.3.4: SEM of 0.5" Hydrocyclone 2nd Pass Underflow (Predominantly Kaolinite)


The screened -325 mesh clay fraction yield was 31.4%, slightly higher than the previous test. The 3 inch cyclone underflow yielded a lower percentage of grit than the July 2008 test. Recovery splits for each pass through the 0.5 inch cyclone were again similar.

The SEM’s showed some separation and concentration of the halloysite in the second and third overflows, but lacked the degree of concentration and separation demonstrated in the previous test results. It was determined that higher percent solids ranges used in this test were detrimental in achieving the results from the previous test.



The combined underflows from the first and second passes through the 0.5 inch diameter cyclones were processed into a metakaolin. The SEM’s for the metakaolin product showed good structure formation that is considered typical of thermally processed pigments. Further testing of the metakaolin product was left for I-Minerals to pursue with a concrete laboratory. This is discussed in the marketing section.

August 2010 Testing

The August 2010 results did not provide results that could be used for further development. The retesting of the second pass overflow was performed in an effort to understand if the higher percent solids used during the initial two 0.5 inch diameter cyclone passes was detrimental to the clay separations. The results were not conclusive, but all indications from published data point to a need for low percent solids in this process application. The results from this test were not considered necessary for further investigation.

October 2010 Testing

The October 2010 test provided clay separation results similar to the July 2010 test. Several recovery deficiencies occurred that were not consistent with known and previously achieved data. Feldspathic sand fraction recovery should be in the range of 70%, but the feldspathic sand recovery in this test exceeded 80% and indicated an inadequate separation of the clay fraction. This was attributed to less than optimum dispersant blunging due to the smaller sample size. The results demonstrated the need for minimum ore volumes for processing through GMT’s pilot plant (also confirmed in the November 2010 work). The presence of the halloysite in the clay fraction was significant.

November 2010 Testing

The November 2010 test also provided clay separation results similar to the July 2010 test. The recovery results closely mirrored the data obtained in the October 2010 test. As noted in the October 2010 test results, the presence of halloysite in the clay fraction was significant. The circuit yields and recoveries for the October 2010 and November 2010 tests are shown in the Table 13.2.3.1.

Specific process variable requirements for GMT’s pilot plant equipment were confirmed with this test. Adequate dispersant blunging of the crude, considered key for successful clay separation, was compromised due to the smaller sample provided by I-Minerals. The pilot plant blunging equipment designated for this service could not be used due to the sample quantity, and an alternative mixing arrangement was used in its place.

2011 Testing

Testing performed in 2011 was primarily performed to generate additional quantities of halloysite product for market development work. The principal company involved in this work is DURTEC GmbH located in Neubrandenburg, Germany and headed by Dr. J. Schomburg. Two major test runs were performed using GMT’s pilot plant – one completed in September 2011 utilizing approximately 2,000 lb of ore and another in December utilizing approximately 6,500 lb of ore. One significant equipment modification was incorporated into the clay processing schematic – a solid bowl decanter centrifuge was used in place of the 0.5 inch diameter cyclone for both passes. The halloysite and kaolin product yields and quality closely matched the previous results, and the centrifuge equipment will be utilized in all future work.



2012 Testing

Testing performed in 2012 was again use to generate additional quantities of halloysite product for market development work. The August 2012 run utilized the approximately 550 lb of ore remaining in GMT’s inventory.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 68 Table 13.2.3.1: Primary Clay Pilot Plant Trials Compilation Recovery Data (1)

Description Pounds, Approximate % Yield Thru Circuit % Recovery "Normalized" Data (2)

Jun-08 Jul-10 Oct-10 Nov-10 Jun-08 Jul-10 Oct-10 Nov-10 Jun-08 Jul-10 Oct-10 Nov-10 Average Pounds % Yield % Recovery Ore Feed 500.00 2,500.00 299.4 250.00 250.00 +325 Mesh Screen Fraction 355.60 1,715.00 243.0 203.00 71.12 68.60 81.16 81.20 71.12 68.60% 81.16% 81.20% 75.52% 175.00 70.00 70.00 -325 Mesh Screen Fraction 144.40 785.00 56.40 47.00 28.88 31.40 18.84 18.80 75.00 30.00 3 inch Cyclone Overflow 131.07 751.45 47.15 35.40 26.21 30.06 15.75 14.16 56.49 22.60 3 inch Cyclone Underflow 13.33 33.55 9.25 11.60 2.67 1.34 3.09 4.64 2.67 1.34% 3.09% 4.64% 2.93% 18.51 7.40 7.40 First Pass Overflow 82.65 509.41 25.80 16.40 16.53 20.38 8.62 6.56 26.17 10.47 First Pass Underflow 48.42 242.04 21.25 19.00 9.68 9.68 7.10 7.60 9.68 9.68% 7.10% 7.60% 8.52% 30.32 12.13 12.13 Second Pass Overflow 52.20 314.82 17.12 12.15 10.44 12.59 5.72 4.86 10.44 12.59% 5.72% 4.86% 8.40% 19.39 7.76 7.76 Second Pass Underflow 30.45 194.59 8.78 4.25 6.09 7.78 2.93 1.70 6.09 7.78% 2.93% 1.70% 4.63% 6.78 2.71 2.71 Total 100.00 100.00 100.00 100.00 100.00 100.00 Feldspathic Sand Recovery 71.12 68.60 81.16 81.20 75.52 70.00 Losses to Waste 2.67 1.34 3.09 4.64 2.93 7.40 Halloysite Recovery (Second Pass Overflow) 10.44 12.59 5.72 4.86 8.40 7.76 Kaolin Recovery (First and Second Pass Underflows) 15.77 17.47 10.03 9.30 13.14 14.84 Total Clay Recovery 26.21 30.06 15.75 14.16 21.55 22.60 Source: I-Minerals

(1) Processing includes 325 mesh screen for degritting, 3 inch cyclone for fine degritting, and 0.5 inch cyclones for clay separation. (2) Using 70% recovery of feldspathic sand and November 2010 yield data.



Primary Clay Pilot Plant Test Conclusions


The 3 inch cyclone was very efficient at separating and concentrating the grit fraction from the clays. Of most importance, the SEM’s illustrated excellent separation and concentration of the halloysite in the second pass overflow fraction using the 0.5 inch diameter cyclone. The underflow fractions contained the highest concentration of kaolinite clay.


The 3 inch cyclone underflow yielded a lower percentage of grit than the July 2008 test. Recovery splits for each pass through the 0.5 inch cyclone were again similar. While the SEM’s showed significant amounts of halloysite, the concentration aspect of the clay separation was not as pronounced. It was determined that higher percent solids ranges used in this test were detrimental in achieving the results from the previous test.

The combined clay cyclone underflows were processed into a metakaolin. Further testing of the metakaolin product is for I-Minerals to pursue with a concrete laboratory. This work is discussed in the marketing section.

Compilation of October 2010 and November 2010 Testing

The results demonstrated that specific attention must be placed on the process dispersion and blunging steps prior to cyclone classification. It is apparent that this area of process preparation is critical to achieving classification results that mirror the excellent data achieved in the July 2008 testing.

Compilation of Primary Clay Pilot Plant Test Data

A compromise recovery table was developed to normalize the available data and develop current product recoveries that best represents the knowledge base achieved to-date. Recovery data was used from the July 2008, July 2010, October 2010, and November 2010 tests. The July 2008 data were excellent and representative of the process capabilities. However, subsequent testing did not replicate these results. Therefore, all recoveries were reviewed to consider obtainable results once circuit optimization is achieved and the following concluded:

• Key consideration was given to the separation of the feldspathic sand and clay fractions. The large sample tests (July 2008 and July 2010) achieved splits of about 70/30 sand to clay that were consistent with core analysis data. The small sample tests (October 2010 and November 2010) achieved splits of only about 81/19 sand to clay – not considered representative. Therefore, a feldspathic sand fraction recovery of 70% was used for the normalized recovery data;

• The 3 inch cyclone grit separation step was adjusted to a very conservative recovery (loss) of 7.40% in the normalized recovery data. The actual results ranged from only 2.67 to 4.64% loss of grit material. However, it was assumed that once optimum feldspathic sand and clay separation is achieved the grit fraction will increase. The conservative 7.40% value used was determined using the normalized clay fraction recovery determinations discussed below. Any grit reduction change would increase the feldspathic sand recovery; and

• The November 2010 test was the only test that used WBL core composite material and therefore the clay separation data from this test were used for the determination of the



normalized clay recoveries. The clay yields were used in conjunction with the sand/clay data above to determine normalized recoveries for the halloysite and kaolinite clay fractions. This test was not considered optimal for use in these calculations, but the results were representative of composited core and viewed as conservative data that did not bias towards the best clay recovery data from the July 2008 work. By comparison, the July 2008 data results were exceptional but not adequately representative of the potential WBL ore body. Also, the July 2010 results were negatively affected in quality by negative processing practices discussed earlier. The remaining tests and data, including November 2010, were incorrectly biased with low clay recoveries due to inadequate feldspathic sand and clay separation. The main comparable distinction in all the data results was the definite presence of halloysite in the clay fraction. Table 13.2.3.1 shows all of the pertinent data.

Hydrocyclone Evaluation December 2011

The solid bowl decanter centrifuge used in place of the 0.5 inch diameter cyclone for both passes resulted in similar product recoveries and quality when compared to previous testing. The circuit is somewhat simplified with this change as a single centrifuge at each pass stage of the process can handle the entire flow stream compared to the high number of 0.5 inch cyclones that would be required for production operation (at 1.3 gpm feed rate per cyclone). The centrifuge overflow and underflow splits were set to match the yields obtained by the 0.5 inch cyclone and the product results obtained were similar.

13.2.4 Results and Conclusions for Testing of WBL K-spar Tailings K-Spar Tailings Test Results

The Pilot Plant testing of K-Spar Tailings consisted of six shakedown runs to test the mechanical integrity of equipment in each process step followed by 25 production runs to evaluate performance as well as to produce feldspar.

• The MRL conducted six shakedown runs, the duration of which lasted from 1 to 8 hours and consumed almost 5 t of feed that is not included in the production figures;

• The feed rates varied from 400 to 600 lb/hr during initial production runs (Tests PP1 through PP5) for adjusting the grind via feed rate and the hydrosizer performance. Fines generated accounted for over 30 wt% indicating that the slimes result in decreased flotation loads and increased reagents consumption to recover acceptable grade feldspar;

• ACME Labs assays results for initial production runs show that alumina values for the spar concentrates ranged from 15.5 to 18.2% Al2O3;

• Re-calibration of the hydrosizer and increasing the feed rate to 500 lb/hr reduced the amine dosage to the spar flotation (Test PP11) and the product grade, as reported by Activation Labs, increased to 17.86 wt% Al2O3, which is very close to the targeted 18 wt% with a yield of 15.8%. ACME Labs reported 18.52% Al2O3 for this run;

• A feed rate of 650 lb/hr was found to be too high as the rod mill began to overload at the end of the run; and

• Table 13.2.4.1 shows that grade increases significantly as hydrofluoric acid (HF) dosage increases.



Table 13.2.4.1: Grade vs. Hydrofluoric Acid Dosage Test No. % Yield % Al2O3 HF, lb/t PP16A 17.3 17.82/18.38(1) 1.00 PP16B 16.0 17.99/18.36 1.20 PP16C 17.1 18.08/18.65 1.40 PP16D 16.7 18.26/19.10 1.55 Source: I-Minerals, 2010 (1) Results reported by Activation Labs/ACME Labs

Table 13.2.4.2 shows that keeping the HF dosage constant, yield increases with increased amine dosage while producing acceptable grades at all collector levels.

Table 13.2.4.2: Grade vs. Collector Amine Dosage Test No. % Yield % Al2O3 Amine, lb/t PP17A 17.7 18.27/18.25(1) 0.30 PP17B 18.0 18.31/18.98 0.40 PP17C 20.4 18.18/18.64 0.50 Source: I-Minerals, 2010 (1) Results reported by Activation Lab/ACME Lab

Conclusions based on K-Spar tailings tests are:

• Over 71,000 lb of K-Spar Tailings as feed were processed in a continuous mode of operation of the Pilot Plant, proving the process and producing bulk samples of approximately 9,000 lb of feldspar are of acceptable grade with 17,000 lb of quartz and about 3,000 lb of middlings;

• The process consistently produced product grades ranging from 18.0 to 18.5% Al2O3 and 13.0 to 13.5% K2O with yield from 17 to 20 wt%; and

• Middling evaluation based on further pilot tests show additional spar product of satisfactory grade (Table 13.2.4.3) can be recovered by simple classification at 50 mesh, with yield increased by up to 15%.

Table 13.2.4.3: Feldspar Middlings Evaluation Test No. Product Wt % % Al2O3 % K2O % Na2O

7088-PP21 Spar Conc. 18.4 17.89 13.59 1.19 +50-mesh Middlings 1.2 17.46 13.21 1.25 Total 19.6 17.86 13.57 1.19

7088-PP22 Spar Conc. 16.3 18.33 14.26 1.11 +50-mesh Middlings 1.2 16.87 12.87 0.99 Total 17.5 18.23 14.16 1.10

7088-P23 Spar Conc. 15.6 18.31 14.06 1.11 +50-mesh Middlings 3.0 16.72 12.19 1.08 Total 18.6 18.05 13.76 1.11


The relatively large amount of feed tested demonstrated that the second scrubbing step can be eliminated, with attendant positive results on reducing both capital and operating costs.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 72 13.2.5 Results and Conclusions for Bench-scale Testing of K-spar Flotation for

Quartz Bench-scale Quartz Testing Results

• Feed characterization was performed and contained 33.9% +30 mesh and 14.9% -200 mesh. Major oxide chemistry in the feed was as follows: 0.23% Fe2O3, 7.55% Al2O3, and 4.23% K2O.

• Since only WBL stockpile tailings had been used in all previous K-spar feed test work, initial tests focused on the recovery of satisfactory K-spar concentrates. Acceptable K-spar chemistry data and recoveries were achieved and duplicated by the fourth test (<0.01% Fe2O3, 18.19% Al2O3, and 14.48% K2O).

• Tests 5 through 8 were conducted to recover upgraded quartz products. Test 5 recovered a 30 mesh quartz (no regrind to 50 mesh) using a spar scavenger float only; Test 6 produced a 50 mesh quartz with a single spar scavenger float; Test 7 produced a 50 mesh quartz product consisting of both iron and spar scavenger floats (double float); and Test 8 produced a 50 mesh quartz product consisting of iron and two spar scavenger floats (triple float).

• Feldspar products collected during the quartz testing continued to produce good grades and consistent yields. Alumina values ran close to 18% Al2O3, iron remained low at 0.01% Fe2O3, and potassium stayed high at 14.6-14.7% K2O. K-spar yields ranged 17-19%.

Table 13.2.5.1: Quartz Products Evaluation Test Wt % Fe (ppm) Al (ppm) K (ppm) Na (ppm) Ca (ppm) Ti (ppm) Li (ppm)

5 53.6 41 838 774 44 11.0 16 0.43 6 43.2 27 176 100 16 9.0 15 0.48 7 38.3 23 145 88 14 8.5 14 0.49 8 32.5 22 131 81 13 8.2 14 0.49


Conclusions Based on Bench-scale Quartz Testing

The following conclusions were noted:

• Residual clay content of the as-received material was estimated at 11.6%. • The initial testing for primary feldspar recovery produced higher grade K-spar products than

those recovered from previously tested WBL tailings stockpile. Iron values were lower (0.01% vs. 0.04-0.10% Fe2O3) while alumina and potassium values were in the anticipated ranges of 17.75-18.25% Al2O3 and 14.5% K2O. Yields also compared favorably at 17-20%.

• Quartz grades improved progressively as the beneficiation was extended. The total impurities were measured at 1750 ppm in the 30 mesh quartz, 355 ppm in the 50 mesh single float quartz, 300 ppm in the 50 mesh double float quartz, and 280 ppm in the triple float quartz.

• The quartz yields decreased as the flowsheet became more complex. The yields were as follows: 53.6% for the 30 mesh quartz, 43.2% for the 50 mesh single float quartz, 38.3% for the 50 mesh double float quartz, and 32.5% for the triple float quartz.

• In the double and triple float schemes, an iron flotation step was used to remove significant amounts of stained quartz. This was observed through microscopic examination.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 73 13.2.6 Results and Conclusions for Pilot Plant Testing of K-Spar Flotation Tailings

for Quartz Pilot-scale Quartz Testing Results

• Equipment schematic set-up dictated that actual production was performed in reverse where the triple float product was produced first, followed by the double float and single float runs. This was necessary to maximize equipment operation efficiency for changeovers and minimize time requirements. While necessary for logistics, this scheme was not desirable for producing optimum products.

• Due to time requirements and costs associated with product chemistry analyses (outside lab required), in-process control of key variables such as reagent optimizations and retention times were not possible. Process variable settings were relegated to data based on previously performed bench flotation studies, and not amenable to actual operating conditions.

• Each flotation run was somewhat limited to the quantity of material available for the testing. • Finished product results were not consistent with previous bench studies as determined by

follow-up chemical analyses.

Pilot-scale Quartz Testing Conclusions

• In order to perform pilot-scale testing for production of high quality quartz, adequate real-time chemistry analysis must be available so that in-process variable adjustments (primarily reagents) can be performed.

• In conjunction with the previous conclusion, significant feedstock must be available to allow for process optimization of each flotation run.

13.2.7 Results and Conclusions for Repeat Bench-scale Testing of K-Spar Flotation Tailings for Quartz Bench-scale Quartz Testing Results

• The number of tests actually performed to produce single, double, and triple float products for each of the two feedstocks (WBL K-spar Tailings material and composite core K-spar material from the Bovill Kaolin deposit) was somewhat limited due to time, material, and cost restrictions associated with the testing, Therefore, complete optimization was not achieved during these tests.

• Even without complete optimization, the tests demonstrated successful results in producing comparable high quality quartz products for each feedstock. The results are shown in the table below.



Table 13.2.7.1: Quartz Products Evaluation Feedstock Product Test Fe Al K Na Ca Ti Li

WBL Tailings Single Float 1 59.5 205 91 17 18.3 33 0.90 Double Float 2 34.8 127 54 17 13.9 26 0.81 Triple Float 3 37.8 131 55 12 13.3 0.86

Composite Core Single Float 4 34.3 187 128 14 14.4 22 0.74 Double Float 5 34.8 140 88 13 14.2 21 0.71 Triple Float 6 30.2 151 86 13 14.5 21 0.70

Source: I-Minerals, 2010 Elements expressed in ppm. Results reported by Cerium Laboratories.

Bench-scale Quartz Testing Conclusions

Favorable comparisons between the two feedstocks were achieved in these tests. This was important to demonstrate the ore body core composite would produce similar quality product results when compared to the WBL Tailings.

13.3 Recovery Estimate Assumptions Assumptions for recovery estimates for each of the products were based on the consistency of the data obtained through the extensive testing performed by GMT. The most variable recovery results were associated with the recovery of the kaolinite and halloysite product fractions, and as a result conservative results were used in order to insure the economic modeling would be representative. This was not the case for the K-spar and quartz product data results. Since such a significant quantity of pilot plant tests were conducted and allowed for specific process variable optimization, the resulting recovery data results were considered to be highly justifiable.

13.4 Sample Representativeness The clay samples used for all of the work performed to-date have come from the exposed face of the north WBL pit within the Bovill Kaolin deposit. Comparisons to core samples that have been extracted from different areas of the same deposit are similar in chemistry and character as discussed in previous sections of this document.

The material used for the K-spar and quartz evaluation work was obtained from the WBL tailings stockpile located adjacent to the WBL pits (Section 4). This material came from the same Bovill Kaolin deposit through previously mined operations, where the clay was washed out and the remaining feldspathic sand fraction was stockpiled. This material is considered to be representative of the Bovill Kaolin deposit as it represents a significant footprint of mined surface area and vertical depth.

13.5 Significant Factors The processing and metallurgical techniques and testing parameters used for both the clay and feldspathic sand processing are considered common practice in the industry. Each product was produced by either bench or pilot-scale processing utilized metallurgical and/or physical separation techniques that are normal for these minerals.



Bench-scale laboratory testing was performed initially to verify quantitative and qualitative parameters for each mineral being investigated. This was then followed by extensive pilot-scale testing to confirm previous results, optimize processing parameters, and provide data for production level scale-up. Detailed specifics for each of the minerals recovered for this Project are described in various sections within this document.

A significant effort was performed to develop a processing scheme for the separation of the two clays – halloysite and kaolinite. It was initially observed that the halloysite naturally separated from the kaolinite by simple gravity separation – allowing the halloysite to remain on the surface of the clay slurry medium. Therefore, effort was placed on accelerating this separation by increasing the physical separation process through mechanical means. Hydrocyclones have proved initially successful and further work is being performed and evaluated on centrifugation techniques. Please reference Sections 13.1.2 and 13.2.2 for more details.

Considerable attention was given to the feldspar flotation processing circuit optimization. This optimization work included attrition scrubbing retention time requirements optimum conditioning determinations, and detailed reagent investigations to optimize combinations and dosage requirements. This work was important to maximize both K-spar product recovery and chemistry. Please reference Sections 13.1.4 and 13.2.4.

It was determined early in the evaluation process that the quartz fraction of the material was of high quality. Considerable time and effort was placed on developing high purity grade quartz products for the marketplace. Ultimately, three grades of quartz product have been developed and integrated into the Project. Please reference Sections 13.1.5, 13.1.6, and 13.2.5.



14 Mineral Resource Estimate Dr. Bart Stryhas constructed the geologic and resource models discussed below. He is responsible for the resource estimation methodology and the resource statement. Dr. Stryhas is independent of the issuer applying all of the tests in Section 3 of NI 43-101.

There are no known material impacts that could negatively affect mineral resource.

14.1 Geology of the Resource Estimation Host material of the resource products is the weathered profile of the granitic phase of the Thatuna Batholith. The Thatuna Batholith is composed mainly of Na-feldspar, K-feldspar and quartz. Weathering has created a residual saprolite type horizon which directly overlies the bedrock from which it was derived. During the natural processes of weathering, the original plagioclase feldspars have preferentially broken down to produce the clays kaolinite and halloysite. The K-feldspars have resisted weathering to a degree and much of the original component remains as free grains. Similarly, the quartz component of the host rock remains as free grains in the weathered material

The geologic model was constructed from the drillhole lithologic descriptions. An upper soil horizon was modeled, by constructing a 3-D base of soil profile. All model blocks located above the base of soil and below topography were coded as un-mineralized soil. Typically, the soil horizon is 10 to 20 ft deep. Directly below the soil horizon, the saprolitic weathered zone of Thatuna is approximately 50 to 125 ft thick. This material hosts the resource products. This zone transitions downward into regolith and un-weathered batholith. The base of saprolitic weathering was modeled based on relative concentrations of clay mineral and geologic descriptions from the drill logs. All blocks located above the base of weathering and below the base of soil were coded as potentially resource bearing. The batholith also contains widely spaced, flat lying roof pendants of un-mineralized Precambrian gneiss. All pendants were modeled and excluded from the potential resource material. Miocene age, basalt dikes typically 10 to 25 ft wide, cut all the other lithologies. They strike at azimuth 140° and dip steeply east approximately 70-75°. These were also modeled and excluded from the potential resource material.

14.2 Drillhole Database The drillhole database supporting the resource estimation consists of 322 diamond core drillholes totaling 35,909 ft. The shallowest hole is 20 ft, the deepest is 260 ft and the average is 111 ft. All the drillholes are oriented vertically and spaced on approximate 100 or 200 ft centers.

Each sample within the drillhole database is characterized by the relative proportions of sand, kaolinite clay, halloysite clay and waste. The sum of these four components equals 100% of each sample. These four variables were estimated as the resource material of this report.

14.3 Capping and Compositing The raw data for sand, kaolinite, halloysite and waste concentrations were plotted on separate histograms and log normal cumulative distribution plots to assess data characteristics and appropriate capping levels. The histograms of all four variables are nearly identical showing a near normal distribution with a slight negative bias. The cumulative distribution plots generally show a continuous, linear distribution up to a point where the data becomes discontinuous and irregular.



Sand, halloysite and kaolinite were all capped at the value where the log normal cumulative distribution plots becomes discontinuous and irregular. All capping was completed prior to compositing. All material capped from the sand, halloysite or kaolinite was added to the clay waste fraction to maintain a 100% mass balance.

The original assay sample lengths generally range from 5 to 10 ft with an average of 5.8 ft. For the modeling, these were composited into 10 ft run length composites. This length was chosen mainly so that approximately two average samples would be composited and the composite length would match the model block height of 10 ft. The composites were broke at the lithologic contacts. Table 14.3.2 lists the results of the capping and compositing.

Table 14.3.1: Capping and Compositing Results

Total Number of Samples Product Capping

Level (%) Number of

Samples Capped

Minimum Capped

Maximum Capped

CV of Capped Composites

1,243

Sand 95 13 95.4 97.8 0.36 Clay Waste None 0 None None 0.51

Halloysite 24 6 25.6 30.9 1.56 Kaolinite 38 20 40.8 52.5 0.61

Source: SRK

14.4 Variogram Analysis Variogram analysis was conducted on the capped bench composites from within the resource material. Semi variograms were constructed for the four variables in all horizontal directions and as omni-directional. The sand, omni-directional semi variogram showed a very crude structure with a large amount of scatter. The kaolinite, omni-directional semi variogram showed a weakly defined structure with a range of about 200 ft, equal to the average drillhole spacing. The halloysite, omni-directional semi variogram showed a pure nugget structure. The waste, omni-directional semi variogram showed a reasonable structure with a range of 200 ft, equal to the average drillhole spacing. Due to the poor or marginal quality of the variograms the grade estimation for all four variables was completed using an inverse distance weighting squared (IDW) algorithm.

14.5 Density I-Minerals conducted density testing on three samples collected from the WBL Pit area that represent the residual deposit material. The density testing was completed by ALS Minerals and showed a mean density of 1.7 g/cm3 for the resource material within the >10% total clay grade shell. A density of 2.0 g/cm3 is used for the material with total clay content of 1-10%. Standard density values were assigned to soil and basalt dikes according to Table 14.5.1.

Table 14.5.1 Block Model Material Densities Material Density g/cm3 Density t/ft3 Soil 1.2 0.0375 Weathered Thatuna 1.7 0.0531 Regolith/sub-crop Thatuna & Pendants 2.0 0.0624 Basalt Dikes 2.9 0.0905 Source: I-Minerals, 2013



14.6 Block Model and Topography Three block models were constructed within the ISP coordinate system parameters listed in Tables 14.6.1 and 14.6.2. A 20 ft x 20 ft x 10 ft (x,y,z) block size was chosen as an appropriate dimension based on the current drillhole spacing and a potential open pit, smallest mining unit. Topography was provided by I-Minerals as a digital map covering the entire resource area. T The topographic surface was created by AeroGeometricc in 2006. The survey was completed using 1:10,000 scale aerial photography and processed to 2.0 ft elevation precision.

Table 14.6.1: Block Model Limits WBL and Middle Ridge Areas Orientation Minimum Maximum Block Size (ft) Easting (ISP) 2,441,680 2,444,720 20 Northing (ISP) 1,903,300 1,908,500 20 Elevation 2,840 3,120 5 Source: SRK

Table 14.6.2: Block Model Limits Kelly’ Hump Areas Orientation Minimum Maximum Block Size (ft) Easting (ISP) 2,445,200 2,447,300 20 Northing (ISP) 1,902,900 1,908,900 20 Elevation 2,840 3,100 10 Source: SRK

14.7 Resource Modeling The two block models described above were each subidived into two model areas based primarily on the sample support represented by the average drill spacing. The WBL area is drilled mainly on 200 ft centers. The Middle Ridge area has an inner portion drilled on 100 ft spacing which is flanked by drilling on 200 ft spacing. The Kelly’s Hump North area is mainly drilled on 100 ft spacing with one area drilled on 200 ft spacing. The Kelly’s Hump South area is all drilled on 200 ft spacing.

The resource estimation is confined within two nested hard boundaries defined by the percentage of total clay which reflects the extent of weathering within the Thatuna granodiorite. The upper/inner, higher grade clay shell was constructed based on a combined halloysite and kaolinite content of 10% or more. This boundary was allowed to extend up to 100 ft from unconfined drillholes. Below or external to the to the 10% clay shell, a lower grade clay shell was constructed based on clay content threshold of1% or more. This lower grade boundary is located below and lateral to the higher grade shell, representing less weathered material. The external grade shell was also allowed to extend 100 ft from unconfined drillholes. Figure 14.7.1 shows the locations of drillholes and the limits of the nested clay shells. Table 14.7.1 lists the number of blocks within each clay shell in each of the four model areas.



Source: SRK, 2013

Figure 14.7.1: Drillhole Locations (Black Dots), Clay Shell >=10% (Blue), Clay Shell >1-<10% (Teal)

Four variables are estimated including, sand, kaolinite, halloysite and waste. The estimations are run independently within each clay shell using only samples within that shell. An inverse distance squared algorithm was used to estimate all variables. The grade estimation utilized a three pass method according to the parameters listed in Table 14.7.2. A varied search orientation was used for the second and third passes based on the strike and dip of the base of soil profile. This profile is interpreted to reflect the pattern of weathering which has created the residual deposits. The varied search orientation is controlled by an anisotropy model which is created by the modeling software. An octant restriction was used to select samples from multiple drill holes. Length weighting was used to account for any short composites at the bottom of drillholes. The number of samples, number of



drillholes and average distance to all samples was stored for each block to be used in the model validation. Once the ID2 estimation was run, all four variables in each block were normalized so they would sum to 100%. As part of the grade estimation, model validation is conducted as an interactive process. To achieve proper validation, higher grade composites were limited by the distance they could be interpolated. A high-grade composite restriction, as listed in Table 14.7.2, means that any sample above the listed grade could only be interpolated over the listed distance. Figures 14.7.2 through 14.7.5 present typical cross-sections showing the estimated block grades for halloysite, kaolinite, sand, and waste, respectively, for each of the model areas.

Table 14.7.1 Percentage of Model Blocks in Clay Shell Model Area Higher Grade Shell (% of Blocks) Lower Grade Shell (% of Blocks) WBL 67 33 Middle Ridge 67 33 Kelly’s Hump North 79 24 Kelly’s Hump South 89 11 Source: SRK

Table 14.7.2 Resource Estimation Parameters

Estimation Area

Clay Shell Estimation

Pass Search Range

(x,y,z) ft Min/Max # Samples

Octant Restriction

High Grade Composite Restriction

WBL

Higher Grade

1 10,10,5 (Box) 1/3 None None 2 250,250,10 3/8 2 Samp/Oct 3 300,300,20 3/8 None

Lower Grade


Middle Ridge

Higher Grade

1 10,10,5 (Box) 1/3 None None 2 175,175,10 3/8 2 Samp/Oct None

3

400,400,20 3/8 None Kaolinite>16% <225ft,225ft,10ft Waste >9% <225ft,225ft,10ft

Lower Grade

1 10,10,5 (Box) 1/3 None 2 200,200,15 3/8 2 Samp/Oct

3 300,300,15 3/8 None Halloysite>4.5% <225ft,225ft,10ft

Kelly’s Hump North

Higher Grade

1 10,10,5 (Box) 1/3 None None 2 125,125,10 3/8 2 Samp/Oct None

3

300,300,20 3/8 None Kaolinite>20% <150ft,150ft,10ft Halloysite>8.5% <150ft,150ft,10ft Sand>75% <200ft,200ft,10ft Waste>7% <200ft,200ft,10ft

Lower Grade

1 10,10,5 (Box) 1/3 None None 2 150,150,15 3/8 2 Samp/Oct None 3 300,300,25 3/8 None None

Kelly’s Hump South

Higher Grade


Lower Grade


Source: SRK



Source: SRK

Figure 14.7.2 WBL, East West Cross Section 1,904,800N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste



Source: SRK

Figure 14.7.3 Middle Ridge, East West Cross Section 1,906490N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste



Source: SRK

Figure 14.7.4 Kelly’s Hump North, East West Cross Section 1,907,200N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste



Source: SRK

Figure 14.7.5 Kelly’s Hump South, East West Cross Section 1,904,200N Viewing North; Composite and Estimated Block Grades, From Top to Bottom-Sand, Kaolinite, Halloysite and Waste



14.8 Model Validation Four techniques were used to evaluate the validity of the block model. First, during the grade estimation; the estimation pass, the number of samples used, the number of drillholes used and the average distance to samples was stored. This data was checked to evaluate the performance of the sample selection parameters discussed above. The results of each estimation are listed in Table 14.8.1. Second, the interpolated block grades were visually checked on sections and bench plans for comparison to the composite grades. Third, statistical analyses were made comparing the estimated block grades to the composite sample data supporting the estimation. The results in Table 14.8.2 show good relations for all variables within the higher grade clay shell which is supported by greater data density. Within the lower grade clay shell halloysite and kaolinite block grades do vary from composite grades primarily due to the paucity of data in certain parts of the grade shell. The fourth validation is a nearest neighbor estimation comparison. The total contained material, at a zero CoG in the NN models were compared to the IDW grade models at the same CoG. The results are listed in Table 14.8.3. These show that no significant material is being manufactured during the modeling process. All four-model validation tests described above provided good confidence in the resource estimation.



Table 14.8.1: Grade Estimation Performance Parameters Estimation Area Clay Shell Criteria Result

WBL

Higher Grade

% Blocks Estimated in First Pass 1 % Blocks Estimated in Second Pass 46 % Blocks Estimated in Third Pass 53 Average Number of Samples Used Per Block 4 Average Number of Drillholes Used Per Block 2.3 Average Distance to Samples 120

Lower Grade

% Blocks Estimated in First Pass 1 % Blocks Estimated in Second Pass 30 % Blocks Estimated in Third Pass 69 Average Number of Samples Used Per Block 4 Average Number of Drillholes Used Per Block 2.1 Average Distance to Samples (ft) 169

Middle Ridge

Higher Grade


Lower Grade



Higher Grade


Lower Grade

% Blocks Estimated in First Pass 2 % Blocks Estimated in Second Pass 64 % Blocks Estimated in Third Pass 34 Average Number of Samples Used Per Block 5 Average Number of Drillholes Used Per Block 3 Average Distance to Samples (ft) 99


Higher Grade


Lower Grade


Source: SRK



Table 14.8.2: Statistical Model Validation Estimation Area

Clay Shell Variable Average Composite

Value (%) Average Block

Value (%) % Difference

Comps to Blocks

WBL

Higher Grade

Sand 73.93 73.87 0.09 Kaolinite 15.87 15.95 -0.53 Halloysite 1.80 1.77 1.95 Waste 7.63 7.66 -0.33

Lower Grade

Sand 72.74 72.48 0.36 Kaolinite 8.13 8.14 -0.09 Halloysite 0.0 0.0 0.0 Waste 4.46 4.31 3.22

Middle Ridge

Higher Grade

Sand 71.41 71.46 -0.08 Kaolinite 11.96 11.80 1.34 Halloysite 4.67 4.05 13.3 Waste 7.25 7.21 0.46

Lower Grade

Sand 53.90 53.54 0.67 Kaolinite 7.24 7.04 2.85 Halloysite 0.56 0.59 -0.80 Waste 4.63 4.46 3.22


Higher Grade

Sand 71.38 69.47 2.67 Kaolinite 12.48 12.57 -0.67 Halloysite 3.80 3.82 -0.62 Waste 6.16 6.25 0.38

Lower Grade

Sand 77.56 77.97 -0.54 Kaolinite 7.56 7.01 7.25 Halloysite 0.96 0.45 52.86 Waste 3.86 3.72 3.62


Higher Grade

Sand 65.63 65.59 0.05 Kaolinite 19.10 18.72 1.88 Halloysite 1.56 1.55 0.45 Waste 6.49 6.39 1.67

Lower Grade


Source: SRK



Table 14.8.3: Nearest Neighbor Model Validation

Estimation Area Clay Shell IDS/NN

Tons (M) Variable IDW

Normalized Grade (%)

NN Normalized Grade (%)

% Diff Contained Material IDS to

NN

WBL

Higher Grade 1.359


Lower Grade 0.620

Sand 84.79 81.03 4.43 Kaolinite 9.94 9.63 3.12 Halloysite 0 0 0.0 Waste 5.24 5.07 3.24

Middle Ridge

Higher Grade 3.249


Lower Grade 1.520



Higher Grade 3.746

Sand 75.14 74.58 0.75 Kaolinite 13.92 13.85 0.50 Halloysite 4.15 4.44 -6.99 Waste 6.78 6.81 -0.44

Lower Grade 1.036



Higher Grade 2.085


Lower Grade 0.241


Source: SRK



14.9 Resource Classification The Mineral Resources are classified under the categories of Measured, Indicated and Inferred according to CIM guidelines. Classification of the resources reflects the relative confidence of the grade estimates and the continuity of the mineralization. This classification is based on several factors including; sample spacing relative to geological and geo-statistical observations regarding the continuity of mineralization, data verification to original sources, specific gravity determinations, accuracy of drill collar locations, accuracy of topographic surface, quality of the assay data and many other factors, which influence the confidence of the mineral estimation. No single factor controls the resource classification rather each factor influences the result.

The resources are all classified as indicated based on the drillhole spacing, the sufficiency of the density testing and the lack of precise QA/QC monitoring of the various analytical techniques.

14.10 Mineral Resource Statement The mineral resource statement in Table 14.10.1 is confined within a Whittle™ pit design according to the parameters listed in Section 16.2. No CoG is applied to the resource because all recovered material in the resource estimation contains sufficient sand, kaolinite or halloysite to be mined for a profit.



Table 14.10.1: Indicated Mineral Resource Statement, (as of 20 April 2014) Classification Location Tons

(M) Qtz & K-Spar Sand

(%) Kaolinite

(%) Halloysite

(%) Qtz & K-Spar and Tons

(000’s) Kaolinite Tons

(000’s) Halloysite Tons

(000’s)

Measured Kellys Hump 2.3 76.7 13.0 3.9 1,761 297 90 Middle Ridge 1.0 74.8 12.0 5.9 745 120 58 All 3.3 76.1 12.7 4.5 2,505 417 148

Indicated


M & I


Source: SRK



14.11 Relevant Factors There are no known legal, political, environmental, or other risks that could materially affect the potential development of the mineral resources



15 Mineral Reserve Estimate Mineral reserves have been estimated in accordance with CIM guidelines. CoG’s were not applied since all weathered Thatuna material in the resource estimation contains sufficient sand, kaolinite or halloysite to be mined for a profit.

The proven and probable reserves are presented in the Table 15.1.

Table 15.1: Mineral Reserve Statement, (as of June 14, 2014)

Reserve Mt Halloysite Grade

Kaolin Grade

Qtz & K-Spar Sand (%)

Halloysite Tons

Kaolinite Tons

Qtz & K-Spar Sand (t)

Kelly Hump Proven 1.7 4.8% 13.5% 81.7% 82,000 229,000 1,389,000 Probable 1.0 6.0% 15.4% 78.6% 60,000 154,000 782,000 Kelly South Proven 0 0 0 0 0 0 0 Probable 1.3 1.6% 23.2% 75.3% 20,000 296,000 959,000 Middle Ridge Proven 0.7 6.9% 12.8% 80.3% 48,000 90,000 563,000 Probable 1.4 4.6% 13.1% 82.3% 66,000 187,000 1,179,000 WBL Proven 0 0 0 0 0 0 0 Probable 0.8 2.4% 16.5% 81.1% 18,000 128,000 629,000 Source:SRK Notes:

1. Some numbers may not add up as a function of rounding. 2. Reserves are based on 100% mine recovery and 0% dilution. This is due to the small equipment being utilized and the

selectivity of the material being mined. This will require further review as part of additional studies. 3. Halloysite processing recovery is 90%. 4. Kaolinite processing recovery is 90%. 5. Quartz and K-Spar processing recovery is 68%. 6. Variable selling prices were used depending on supply. 7. There is an overall strip ratio of 0.69:1 (waste: ore).

15.1 Conversion Assumptions, Parameters and Methods The conversion of mineral resources to open pit ore reserves is accomplished through the pit optimization, pit design, and associated modifying parameters.

15.2 Relevant Factors There are no known legal, political, environmental, or other risks that could materially affect the potential development of the mineral resources.



16 Mine Design Section 1 addresses the mine design for the Bovill mine operation. The deposit and surrounding areas had historically been open-pit mined for clays. This is still the preferred mining method due to the proximity of the mineralized material to the surface.

As mentioned previously, the operation will produce quartz sand, K-feldspar sand, kaolinite clay, metakaolin clay, and halloysite clay. The deposit consists of four separate mining areas: WBL Pit, Middle Ridge (MR), Kelly South (KS) and Kelly’s Hump (KH). The current area of mining is limited by lease and wetlands boundaries.

16.1 Geotechnical Conditions A professional services corporation STRATA of Pullman, WA, performed a field investigation to evaluate the properties of the materials in the deposit area and to identify engineering characteristics. The results of a field investigation are described in a geotechnical exploration report on June 27, 2012 on behalf I-Minerals (STRATA, 2012). A pit slope analysis produced by STRATA for the WBL and MR areas assumed 20 ft high benches for both areas, a 10 ft berm for the WBL pit area, and a 13 ft berm for MR area. STRATA recommended a bench face angle to be excavated at a slope 1.5H:1V and near-surface slope angle (first 7 to 10 ft under the soil cover) at a slope 0.5H:1V to minimize raveling. The overall slope angles for subsurface material and for mineralized are shown in the Table 16.1.1.

Table 16.1.1: Slope Angles from Geotechnical Report by STRATA (2012) Parameter WBL Pit Area Middle Ridge Area Overall slope angle in overburden (soil) 34 34 Overall slope angle (in mineralized area) 45 43 Bench face angle 63 63 Source: STRATA, 2012

Due to the shallow nature of the deposit the stability of the open pit slopes was not considered to be a significant issue. Proposed overall slope angles of 45 for WBL pit area and 43 for Middle Ridge area deemed to be achievable over the mine life.

No specific geotechnical report has been commissioned for the Kelly South and Kelly’s Hump deposits. As the geology in the area is considered relatively homogenous, the geotechnical parameters for the Middle Ridge area have been applied to the Kelly South and Kelly’s Hump deposits.

16.2 Pit Optimization Pit optimization is based on preliminary economic estimations of mining, processing and selling related costs, slope angles and mineral recoveries. These pit optimization factors are likely to vary from those reported in the final economic analysis, which are based in the final pit design and production schedule. The pit optimization software considered mineral percentages and tonnages in the model along with estimated recoveries, mining and processing factors, and costs to determine what material could be economically extracted through the use of the Lerchs-Grossman algorithm.



The integral part of pit optimization is the generation of the nested economical pit shells calculated using economical parameters with the range of multipliers (from 0.1 to 1.1), also called revenue factors. The resulting economical pit shells were used as the guide to create mineable pit phases. The economic pit shell calculated with the multiplier equal to 1 defined the ultimate pit boundaries. The ultimate pit defines what is economically mineable for the specific deposit given certain mining and economic parameters.

16.2.1 Block Models Pit optimization was performed using the block models received on April 21, 2014.

A single block model which incorporated the Middle Ridge and WBL deposits was provided (Middle Ridge Block Model), and a single block model which incorporated the Kelly’s Hump and Kelly South deposits was provided (Kelly’s Hump).

The block models were provided in Vulcan format.

The block models were constrained by the lease boundaries and the delineated wetland areas, both which were provided by HDR Engineering Inc. Blocks within the block model but outside the designated “available” areas were coded within the block model and were unavailable to be mined as part of the optimization process.

16.2.2 Parameters Whittle™ optimization was completed on the block models received April 21, 2014.

Resulting nested economical shells were used as a guide for the pit creation and phase design.

Whittle™ model input consisted of the following:

• Percentages of halloysite, kaolinite and sand, and Geological confidence classification of Measured, Indicated and Inferred were used as the input for the Whittle™ optimization;

• Only material with a Measured and Indicated confidence were used to drive the revenue of the optimization;

• Percentages were converted to the units of material and post-recovery tons of products were calculated in Vulcan®;

• Units of materials for sand, kaolinite and halloysite along with products tons were exported to Whittle™ from Vulcan®;

• Costs and revenues were unitized to tons; • For the purpose of the optimization, kaolinite products were combined into one product in

Whittle™ with single weighted selling price; • For the purpose of the optimization, sand products were combined into one product in

Whittle™ with single weighted selling price for sand products; • 5% gross revenue royalty for the state of Idaho was calculated for all three Whittle™

products: combined sand product, combined clay product, and halloysite product. • The optimization revenue was driven by Halloysite and Kaolin. If the optimization were

allowed to run and include the full revenue for the sand products, the pit shells produced encompass almost all the mineralization, however due to market conditions this is not a feasible option. Therefore, a selling price of US$20/t was allocated to sand products to ensure some value of the sand was captured.



The economic parameters used at the time of the pit optimization contain some variance from those stated in the economic model but do not introduce a material change.

Table 16.2.2.1: Whittle™ Kelly’s Hump Block Model Import Parameters Whittle™ Parameters Type Units Value Data units metric or imperial imperial Block Model Parameters Block Model date April 21, 2014 BM Restrictions Lease polygon and

wetland boundaries Categories Grade A Halloysite percent Grade B Kaolin percent Grade C Sand percent BM Dimensions block size (X - Y - Z ) - 20 - 20 - 10 Number of blocks X - Y - Z - 105 - 300 - 26 Reblock in Whittle™ no reblock Origin of BM X 2,445,200 Y 1,902,900 Z 2,840 Slope Top Soil degrees 34 Material other than Top Soil degrees 65 Source: SRK

Table 16.2.2.2: Whittle™ Middle Ridge Block Model Import Parameters Whittle™ Parameters Type Units Value Data units metric or imperial imperial Block Model Parameters Block Model date April 21, 2014 BM Restrictions Lease polygon and

wetland boundaries Categories Grade A Halloysite percent Grade B Kaolin percent Grade C Sand percent BM Dimensions block size (X - Y - Z ) - 20 - 20 - 10 Number of blocks X - Y - Z - 152 - 260 - 28 Reblock in Whittle™ no reblock Origin of BM X 2,441,680 Y 1,903,300 Z 2,840 Slope Top Soil degrees 34 Material other than Top Soil degrees 65 Source: SRK

Within the block models, a single kaolinite mineral has been modelled. This is used to produce two products, kaolin and metakaolin, with different selling prices. An estimated weighted average (depending on recovery and plant capacity) has been used to generate a selling price for the single mineral Kaolin. The weighted average selling price is US$148/t.



The same situation applies to the modelled variable Sand. It is used to produce k-spar and quartz products. As the selling price for both K-Spar and Quartz is the same, the selling price of US$220/t has been used.

The recovery split per product from each variable is fixed and a combined recovery and selling prices were used for kaolinite products and sand final products within Whittle™. Table 16.2.2.3 shows the total recoveries per variable (halloysite mineral is not included as it is used to produce single product).

Table 16.2.2.3: Combined Recoveries for Kaolinite and Sand Minerals

Variable Product Recovery (%) Kaolinite Kaolin 45.3 Kaolinite Metakaolin 49.7 Total kaolinite recovery

95

Sand K-spar 22 Sand Quartz 55 Total sand recovery

77

Source: SRK

The combined selling prices are displayed in the Table 16.2.2.4.

Table 16.2.2.4: Combined Selling Prices for Kaolinite and Sand Minerals Mineral Product Sell Price (US$/t) Kaolinite Kaolin 95 Kaolinite Metakaolin 225 Kaolinite Recovered Ton of products 148 Sand K-spar 220 Sand Quartz 220 Sand Recovered ton of products 220 Source: I-Minerals, 2014



The economic parameters are listed in the Table 16.2.2.5.

Table 16.2.2.5: Whittle™ Economic Parameters Whittle™ Parameters Type Units Value Mining Costs Waste Mining Cost US$/t mined 2.54 Ore Mining Cost US$/t mined 3.96 Mining recovery fraction 1.00 Mining dilution factor 1.00 Processing Process Name combined clay-sand plant Mill Rock types Resource Confidence t Measured t Indicated t Inferred t Waste Process Cost Ore selection method Cash_Flow Process cost US$/t of ore 30.32 General and administration US$/t of ore 1.00 Transportation cost US$/t of ore 0.00 Recoveries (by final product) Halloysite % 95.00 Kaolin % 95.00 Sand % 77.00 Revenue and Selling Cost selling price (revenue) Halloysite US$/t product 1200.00 selling price (revenue) Kaolin US$/t product 148.00 selling price (revenue) Sand US$/t product 20.00 State of Idaho royalty percentage of gross revenue % 5.00 Sales surcharge US$/t ore 0.66 Optimization Revenue factor range 0.3-2.0 Operational Scenario - Time Costs Initial Capital Cost US$ 0 Discount Rate Per period % 0% Operational Scenario - Limits Mining limit t 0 Process limit (PRC) US% 0 Source: SRK

16.2.3 Results As part of the Whittle™ optimization, revenue factors are used as the multiplier for all economic parameters to produce a multitude of the intermediate shells. Figure 16.2.3.1 and Figure 16.2.3.2 show a pit by pit graph produced by Whittle™. This graphs shows how each deposit reacts to different revenue factors ranging from 0.6 to 1.08 based on the assumptions listed in Tables 16.2.3.1 and 16.2.3.2.



Table 16.2.3.1: Whittle™ Economic Pits Middle Ridge Whittle

Pit Revenue

Factor Total Tons

Ore Tons

Waste Tons

Strip Ratio

Halloysite (tons)

Halloysite Grade (%)

Kaolinite (tons)

Kaolinite Grade (%)

Sand (tons)

Sand Grade (%)

16 0.6 2,461,978 1,654,342 807,636 0.49 119,007 7.19 193,407 11.69 1,221,450 73.83 17 0.62 2,581,022 1,736,318 844,704 0.49 122,311 7.04 203,012 11.69 1,284,804 74.00 18 0.64 2,698,918 1,803,322 895,596 0.50 124,979 6.93 210,651 11.68 1,336,835 74.13 19 0.66 2,811,668 1,870,864 940,804 0.50 127,554 6.82 218,356 11.67 1,389,453 74.27 20 0.68 2,906,782 1,925,324 981,458 0.51 129,524 6.73 224,884 11.68 1,431,609 74.36 21 0.7 2,979,554 1,975,674 1,003,880 0.51 131,236 6.64 230,996 11.69 1,470,528 74.43 22 0.72 3,056,370 2,027,384 1,028,986 0.51 132,875 6.55 237,451 11.71 1,510,348 74.50 23 0.74 3,144,202 2,077,218 1,066,984 0.51 134,446 6.47 243,423 11.72 1,549,079 74.57 24 0.76 3,236,578 2,130,988 1,105,590 0.52 135,995 6.38 250,434 11.75 1,590,459 74.63 25 0.78 3,295,224 2,165,956 1,129,268 0.52 136,960 6.32 255,078 11.78 1,617,263 74.67 26 0.8 3,381,892 2,206,000 1,175,892 0.53 138,140 6.26 259,834 11.78 1,648,436 74.73 27 0.82 3,443,428 2,245,132 1,198,296 0.53 139,095 6.20 265,202 11.81 1,678,427 74.76 28 0.84 3,505,716 2,278,654 1,227,062 0.54 139,950 6.14 269,439 11.82 1,704,488 74.80 29 0.86 3,558,582 2,306,550 1,252,032 0.54 140,621 6.10 273,114 11.84 1,725,968 74.83 30 0.88 3,599,874 2,332,440 1,267,434 0.54 141,208 6.05 276,429 11.85 1,746,102 74.86 31 0.9 3,641,424 2,354,748 1,286,676 0.55 141,699 6.02 279,388 11.86 1,763,329 74.88 32 0.92 3,684,234 2,379,144 1,305,090 0.55 142,235 5.98 282,435 11.87 1,782,362 74.92 33 0.94 3,753,702 2,406,626 1,347,076 0.56 142,821 5.93 286,239 11.89 1,803,492 74.94 34 0.96 3,816,294 2,436,142 1,380,152 0.57 143,396 5.89 290,332 11.92 1,826,183 74.96 35 0.98 3,874,606 2,467,732 1,406,874 0.57 144,032 5.84 294,183 11.92 1,851,085 75.01 36 1 3,927,594 2,497,486 1,430,108 0.57 144,582 5.79 297,975 11.93 1,874,326 75.05 37 1.02 3,994,428 2,532,126 1,462,302 0.58 145,172 5.73 302,718 11.96 1,901,181 75.08 38 1.04 4,057,572 2,564,152 1,493,420 0.58 145,675 5.68 307,283 11.98 1,925,760 75.10 39 1.06 4,160,714 2,597,902 1,562,812 0.60 146,267 5.63 312,016 12.01 1,951,674 75.13 40 1.08 4,235,972 2,631,174 1,604,798 0.61 146,593 5.57 318,043 12.09 1,976,083 75.10

Source: SRK



Table 16.2.3.2: Whittle™ Economic Pits Kelly’s Hump

Whittle™ Pit

Revenue Factor

Total Tons

Ore Tons

Waste Tons

Strip Ratio

Halloysite (tons)

Halloysite Grade (%)

Kaolinite (tons)

Kaolinite Grade (%)

Sand (tons)

Sand Grade

(%) 16 0.6 3,325,228 2,153,786 1,171,442 0.54 149,128 6.92 281,872 13.09 1,570,687 72.93 17 0.62 3,488,486 2,241,472 1,247,014 0.56 151,849 6.77 299,033 13.34 1,631,859 72.80 18 0.64 3,605,170 2,301,440 1,303,730 0.57 153,819 6.68 309,544 13.45 1,675,018 72.78 19 0.66 3,694,188 2,362,392 1,331,796 0.56 155,677 6.59 319,979 13.54 1,719,434 72.78 20 0.68 3,810,970 2,429,840 1,381,130 0.57 157,597 6.49 331,870 13.66 1,768,124 72.77 21 0.7 3,923,244 2,483,318 1,439,926 0.58 159,052 6.40 342,025 13.77 1,806,045 72.73 22 0.72 4,038,640 2,556,430 1,482,210 0.58 160,927 6.30 354,932 13.88 1,859,376 72.73 23 0.74 4,118,438 2,612,240 1,506,198 0.58 162,267 6.21 364,731 13.96 1,900,271 72.74 24 0.76 4,220,052 2,675,040 1,545,012 0.58 163,528 6.11 377,258 14.10 1,945,131 72.71 25 0.78 4,276,838 2,716,820 1,560,018 0.57 164,378 6.05 385,039 14.17 1,975,424 72.71 26 0.8 4,344,444 2,759,698 1,584,746 0.57 165,237 5.99 392,857 14.24 2,006,640 72.71 27 0.82 4,414,464 2,806,952 1,607,512 0.57 166,073 5.92 401,925 14.32 2,040,772 72.70 28 0.84 4,479,748 2,852,146 1,627,602 0.57 166,699 5.84 411,548 14.43 2,072,672 72.67 29 0.86 4,553,232 2,901,096 1,652,136 0.57 167,448 5.77 421,062 14.51 2,108,003 72.66 30 0.88 4,647,648 2,957,346 1,690,302 0.57 168,181 5.69 432,858 14.64 2,147,985 72.63 31 0.9 4,732,866 3,004,562 1,728,304 0.58 168,688 5.61 443,431 14.76 2,180,654 72.58 32 0.92 4,846,762 3,060,296 1,786,466 0.58 169,335 5.53 455,369 14.88 2,219,838 72.54 33 0.94 4,946,684 3,115,780 1,830,904 0.59 169,844 5.45 467,707 15.01 2,258,408 72.48 34 0.96 5,046,174 3,169,818 1,876,356 0.59 170,381 5.38 479,118 15.12 2,296,581 72.45 35 0.98 5,102,606 3,207,358 1,895,248 0.59 170,776 5.32 486,405 15.17 2,323,891 72.46 36 1 5,188,300 3,255,172 1,933,128 0.59 171,277 5.26 495,626 15.23 2,358,844 72.46 37 1.02 5,268,256 3,300,866 1,967,390 0.60 171,652 5.20 505,006 15.30 2,391,794 72.46 38 1.04 5,336,458 3,345,038 1,991,420 0.60 171,955 5.14 514,142 15.37 2,423,604 72.45 39 1.06 5,405,300 3,388,476 2,016,824 0.60 172,223 5.08 523,157 15.44 2,454,829 72.45 40 1.08 5,465,842 3,413,530 2,052,312 0.60 172,376 5.05 528,343 15.48 2,472,901 72.44

Source: SRK



Source: SRK

Figure 16.2.3.1: Kelly’s Hump Results



Source: SRK

Figure 16.2.3.2: Middle Ridge Results



16.3 Pit Design The operation is planned to be mined via open pit mining methods. To determine an achievable mining sequence, a series of pit designs have been created for each deposit. Each deposit consists of several Phases which allow greater control over the schedule.

A total of nine phases have been designed.

16.3.1 Parameters The physical parameters used in the design are detailed in Table 16.3.1.1.

Table 16.3.1.1: Pit Design Parameters Parameter Units WBL Pit MR Pits Kelly’s Hump Pits Kelly South Pits Bench height ft 10 10 10 10 Face angle, soil degrees 34 34 34 34 Face angle, degrees 65 65 65 65 Berm width ft 10 10 10 10 Road width ft 40 40 40 40 Road grade % 10 10 10 10 Source: STRATA. 2012

16.3.2 Results Nine phases were designed in order to achieve the required clay/sand blend. The phases were based on the distinct pits formed from the Whittle™ optimization. The pits were reviewed and additional phases were not required to meet the schedule requirements.



Source: SRK

Figure 16.3.2.1: Kelly’s Hump Final Design

Phase 1

Phase 2

Phase 3



Source: SRK

Figure 16.3.2.2: Kelly South Final Design

Phase 1

Phase 2



Source: SRK

Figure 16.3.2.3: Middle Ridge Final Design

Phase 1

Phase 2



Source: SRK

Figure 16.3.2.4: WBL Final Design

Phase 2

Phase 3



Table 16.3.2.1: Kelly’s Hump Phases Tonnages Item Phase 1 Phase 2 Phase 3 Total Measured 0 1,111,318 753,676 1,864,994 Indicated 858,948 150,445 91,749 1,101,143 Waste 353,352 629,427 358,880 1,341,658 Strip Ratio 0.41 0.50 0.42 0.45 Source: SRK

Table 16.3.2.2: Kelly South Phases Tonnages Item Phase 1 Phase 2 Total Measured 0 0 0 Indicated 811,991 595,130 1,407,121 Waste 417,755 392,650 810,405 Strip Ratio 0.51 0.66 0.58 Source: SRK

Table 16.3.2.3: Middle Ridge Phases Tonnages Item Phase 1 Phase 2 Total Measured 774,378 0 774,378 Indicated 1,450,853 138,988 1,450,853 Waste 1,025,338 102,464 1,025,338 Strip Ratio 0.46 0.74 0.48 Source: SRK

Table 16.3.2.4: WBL Phases Tonnages Item Phase 2 Phase 3 Total Measured 0 0 0 Indicated 644,313 213,292 857,604 Waste 550,686 196,925 747,611 Strip Ratio 0.85 0.92 0.87 Source: SRK

16.3.3 Pit Design versus Optimized Shell Figures 16.3.3.1 and 16.3.3.2 show the pit designs (translucent blue) compared with the optimized shells (orange).

Table 16.3.3.1 and Table 16.3.3.2 show the differences between the designed pit volumes and the optimized pits.

Table 16.3.3.1: Kelly’s Hump and Kelly South Pit Design versus Optimized Shells Pit Designed Pits Optimized Shell Difference (%) Measured & Indicated Tons 4,373,258 4,614,657 95% Waste Tons 2,152,064 1,968,142 109% Total 6,525,322 6,582,798 99% Source: SRK



Table 16.3.3.2: Middle Ridge and WBL Pit Design versus Optimized Shells Pit Designed Pits Optimized Shell Difference (%) Measured & Indicated Tons 3,221,823 3,408,611 95% Waste Tons 1,875,412 1,637,941 114% Total 5,097,235 5,046,552 101% Source: SRK



Source: SRK

Figure 16.3.3.1: Kelly’s Hump and Kelly South Pit Designs and Optimized Shells



Source: SRK

Figure 16.3.3.2: Middle Ridge and WBL Pit Designs and Optimized Shells

Figure 16.3.3.3 shows an indicative cross section showing the optimized shell (orange) and the designed shell with topography (blue).



Source: SRK

Figure 16.3.3.3: Cross Section (Designed pit versus Whittle™ shell)

16.3.4 Waste Dumps Two waste dumps have been designed. The waste contained within each deposit is assumed to be benign and not require any specific handling.

The waste dumps were designed in 20 ft lifts providing an overall slope angle of 26⁰. This design will allow easy slope contouring during mine closure. The waste dump design parameters are shown in Table 16.3.4.1.

Table 16.3.4.1: Waste Dump Design Parameters Parameter Units Waste Dump Bench height ft 20 Face angle degrees 34 Berm width ft 10 Overall slope degrees 26 Source: STRATA, 2012

The Kelly’s Hump waste dump has approximately 1.4 Mm3 capacity. The Kelly’s Hump and Kelly South deposits require approximately 1.3 Mm3 to store the waste material. A 30% swell factor has been applied to calculate volume requirements.

Figure 16.3.4.1 shows a plan view of the Kelly’s Hump Dump design.



Source: SRK

Figure 16.3.4.1: Kelly’s Hump Waste Dump (Plan View)

The Middle Ridge waste dump has approximately 1.9 Mm3 capacity. The Kelly’s Hump and Kelly South deposits require approximately 1.5 Mm3 to store the waste material. A 30% swell factor has been applied to calculate volume requirements.

Figure 16.3.4.2 shows a plan view of the Middle Ridge Dump design.



Source: SRK

Figure 16.3.4.2: Middle Ridge Waste Dump (Plan View)

The waste storage requirements for both dumps have a 10% multiplier on requirements to ensure that any additional unplanned waste can be accommodated in the designed dumps.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 114 16.3.5 Pit and Waste Dump Layout

Source: SRK

Figure 16.3.5.1: Pit and Waste Dump Layout



Source: HDR Engineering, 2012; SRK

Figure 16.3.5.2: Pit and Waste Dump Layout with Constraints

16.4 Mine Schedule As part of the scheduling process, the client identified the following criteria as critical to the success of the project:

• Minimum production of 7,500 t Halloysite per year; • Maximize the production of the sand plant (230.4 kt/y); • Maximum material moved per year of 500 kt;

For the purposes of the following graphs, Halloysite and Kaolin are to be fed to the “Clay Plant”, and all Sand material is to be fed to the “Sand Plant”.

Identified Wetlands

Id tifi d

License Boundary (pale grey)

Li B d



Table 16.4.1 shows the mine schedule.

Table 16.4.1: Mine Schedule Year Halloysite Kaolin Sand Plant Waste TMM

1 12,195 31,579 172,840 263,384 479,998 2 8,753 49,328 229,454 192,457 479,992 3 6,486 42,338 229,454 218,517 496,795 4 6,200 64,148 229,454 197,198 497,000 5 15,000 55,760 230,454 195,786 497,000 6 15,000 50,117 230,404 184,480 480,001 7 15,000 47,099 229,454 188,446 479,999 8 13,342 46,308 230,454 189,897 480,001 9 14,692 36,792 230,454 198,060 479,998

10 15,000 34,917 230,454 199,626 479,997 11 15,000 34,437 230,454 200,106 479,997 12 15,000 30,810 200,244 233,944 479,998 13 13,281 36,047 212,000 218,672 479,999 14 14,773 36,794 230,454 197,873 479,895 15 14,890 36,833 230,454 191,587 473,765 16 15,000 37,990 230,454 196,555 479,999 17 15,000 41,468 230,454 151,806 438,729 18 15,000 47,790 230,454 138,754 431,998 19 15,000 50,362 230,454 135,906 431,722 20 15,000 54,597 230,454 131,325 431,376 21 6,520 58,692 230,454 184,315 479,980 22 5,000 46,311 218,931 209,757 480,000 23 5,000 37,063 218,931 218,957 479,951 24 5,000 50,661 221,327 202,733 479,722 25 2,052 25,061 113,053 102,804 242,969

Source: SRK



Figure 16.4.1 displays the total material movement of the operation on an annual basis.

Source: SRK

Figure 16.4.1: Total Material Movement

Figure 16.4.2 displays the material feed to the Clay plant on an annual basis.

Source: SRK

Figure 16.4.2: Clay Material Feed



Figure 16.4.3 displays the material feed to the Sand plant on an annual basis.

Source: SRK

Figure 16.4.3: Sand Material Feed Figure 16.4.4 displays the material feed by Resource Classification on an annual basis.

Source: SRK

Figure 16.4.4: Resource Classification Feed



16.5 Mine Operations Each pit design was divided into benches for creating an annual mine schedule. Due to the distribution of mineralized material within each of the pits, further mine design was not required to achieve the mine schedule.

The mine schedule is controlled by the required blend to the plant.

The primary target of the mine schedule was to meet Halloysite production. This was followed by achieving maximum throughput through the sand plant.

Maximizing the throughput through the sand plant and meeting the Halloysite production target resulted in a stockpiling requirement of Halloysite material.

16.5.1 Mining Operations The mining operation is to be completed via open pit mining methods. Conventional truck and shovel mining equipment will be used to extract both ore and waste. Due to the nature of the material, there is not expected to be any requirement for drill and blast operations.

All waste material from the pits is forecast to be benign, and as such not require any specific handling procedures. All waste is planned to be dumped ex-pit at external waste dumps, within the boundaries of the mining lease. At a further stage of planning, an opportunity exists to analyze the potential to backfill some of pits with waste material.

A contractor (Debco) is currently planned to be used for all earthmoving activities. The contractor is expected to supply all equipment and operators. The client will supply all fuel required to perform all activities.

The site is expected to operate 7 days per week, 8 hours per day, during daylight hours only.

The mine life is estimated at 25 years.

16.5.2 Mining Operations Parameters The final major mine equipment fleet requirements will be based on the annual mine production schedule, the mine work schedule and shift pattern.

The final fleet will be determined by a third party contractor; however, SRK has estimated the type of fleet equipment likely to be employed by the contractor.

Indicative fleet equipment is listed in Table 16.5.2.1. SRK notes the equipment brand and manufacturer is indicative and will be interchanged with the fleet available from the contractor.

SRK has based Table 16.5.2.1 equipment models on previous conversations with the planned contractor.



Table 16.5.2.1: SRK Estimate of Mining Equipment

Equipment Units Used Make Model Size

Drilling Blasthole Drill Atlas Copco DM25-SP 5.5 in Loading Hydraulic Excavator Caterpillar 330CL 7 yd3 Hauling Haul Trucks Komatsu HM-300 21 yd3 Other Mine Equipment Track Dozer Caterpillar D8N 305 hp Motor Grader Caterpillar 140G 259 hp Water Truck Freightliner Model 2700 gal Support Equipment Integrated Tool Handler Caterpillar IT62H 210 hp Mechanic's Field Truck Manufacturer Model 2 t Flatbed with Crane Manufacturer Model 2 t Pumps Manufacturer Model 8 hp Generators Manufacturer Model 27 hp Service Pickups Manufacturer 4x4 3/4 t Light Plants Manufacturer Trailer 11 hp Blasting ANFO/Emulsion Truck Manufacturer Model 180 hp Blasters Crew Truck Manufacturer 4x4 1 t Blasthole Stemmer Caterpillar 226B 56 hp Source: SRK

16.5.3 Drilling and Blasting It is anticipated that the majority of the mined material will be free-dig. Drilling and blasting will be employed only when necessary.

16.5.4 Loading (SRK Estimation) For this study, loading equipment selection by SRK focused on having a hydraulic shovel for primary loading. Haul trucks need to be matched to the loading equipment units, and SRK was informed that Debco planned to use Komatsu HM-300 ADTs.

The hydraulic excavator was estimated to be able to free-dig the majority of the material within the planned open pit. Table 16.5.4.1 shows SRK’s estimate of selected loading statistics for the existing and planned loading units.

Table 16.5.4.1: SRK Estimate of Material Loading Statistics by Loading Unit Type

Equipment Type

Bucket Size and Fill Factor

(yd3) & (%)

Matched Truck Size

(t)

Number of Passes (units)

Total Truck Load Time (1)

(min)

Moving and Delay Time (min/op hr)

Prod. per Unit (2) (100% Avail)

(dt/op hr) Hyd Ex 330CL 3.0 - 86% 30 6 2.50 7.50 120 Source: SRK

1. Includes truck spotting time. 2. Average 10% moisture assumed.

SRK’s planned excavator will be able to load 3.0 yd3 per bucket to the 30 t trucks (assuming loaded rock swell factor of 35% and 10% moisture). Total truck loading time for ore is estimated to be 2.5



minutes (including initial positioning of the trucks for loading). Allowance was made in the loading productivities for cleaning up working faces, and for moving the loading equipment to new working areas.

It is expected approximately 10 trucks will be required to meet the haulage requirements.

As part of the operation, there will be some stockpiling required. The contractor will be responsible for any rehandle requirements to meet a smooth grade blend for production.

16.5.5 Haulage (SRK Estimation) Existing access to the Mine site is via Forest Dev. Road 381. The existing road will be used as an access road between Idaho State Highway 3 and the processing plant. One to two miles of existing road will require work to accommodate highway trucks from the plant and local traffic.

Approximately three miles of roadway will require some work between the existing haul road and the proposed processing facility. Mile 0 to 1.0 will require blade work, pull ditches, and the addition of 4 in. of gravel for the 24 ft wide road. Mile 1.0 to 2.3 will have excavations of cut and fill to establish a 24 ft road width. Mile 2.3 to 3.3 will require excavations of cut and fill to establish a 24 ft road width along with additional gravel for the widened road sections and additional drainage structures.

Stream crossings for the new haul road will be limited to intermittent drainage crossings with culvert installations. It is planned to install a 6 inch water line within the bed of the existing and new access road. This water line will be used to transport water from the water wells and reservoir to the processing plant for use in processing.

The truck selection was influenced by Debco’s planned equipment, loading unit/truck matching, and with respect to meeting production requirements. Ore will be hauled either to the plant ore crusher, or plant crusher stockpiles. The haul distance from the pit to the plant is approximately three miles one-way. Waste will generally be placed in a dump west of the pit, approximately 0.7 miles one-way.

16.5.6 Mine Contractor Labor (SRK Estimation) The mining contractor will have salaried staff for production supervision, and hourly employees for mining production, mining support, and maintenance positions.

SRK estimated a salaried staff allocation of 25% of a senior superintendent to the project, and a full time general foreman position, and assisted by a secretary/clerk.

One mine production and maintenance crew will be necessary. Equipment operator labor allocations were based on the assumption that some of the operators will be cross-trained, and when required to replace another person on leave will be able to operate another type of heavy equipment unit. The mining equipment operators were estimated to number approximately 13 to 14 hourly positions, including one operator for an integrated tool handler, and one operator for a flatbed truck (with crane). A mining equipment maintenance department will be staffed with six mechanics. Hourly maintenance staffing was estimated based on an average of 45% of major mining equipment operations man-hours required.

Additional positions would include a blaster, blaster helper, and surveyor. These positions would not require full-time work allocations to those specific roles, so that a variety of general work duties could be performed by these employees.



An allowance was included for a 25% burden in the wage rates. Given the basic shift hours, the total base paid hours per hourly employee per year were estimated to be 2,080 hours (260 days x 8 hours). Based on an annual vacation allowance of 10 days, six statutory holidays, and assuming six sick/absent days per year the total worked hours per hourly employee per year was estimated to be 1,904 hours [2,080 hours – (10 days x 8 hours) – (2 x 6 days x 8 hours)].

16.5.7 Mining Contractor Infrastructure (SRK Estimation) The mining contractor will provide the following mining infrastructure items:

• Mining contractor offices and mine dry; • Mining contractor maintenance facility; • Mining contractor equipment wash pad and water treatment (permit by I-Minerals); • Mining contractor sewage/septic tanks; and • Diesel fuel storage and dispensing facilities.

I-Minerals will provide power and industrial water for use at the mining contractor’s facilities (for normal usage levels to be established as part of the final mining contract).

16.5.8 I-Minerals (Owner) Mining Functions I-Minerals staff will be responsible for grade control sample collection and delivery to the plant assay laboratory. I-Minerals staff will also be responsible for preparing ore grade control maps for the mining contractor and laying out flagging (markers) in the pit. Geological, geotechnical, mine planning and other similar mining technical functions will be the responsibility of I-Minerals. Technical staff will include a mining engineer, a mining technician/surveyor, and grade control technician. It has been assumed that senior level geological direction for the project would come from I-Minerals corporate level. Three pickup trucks will be available for use. Computer hardware and software will be installed for use by the technical staff.

I-Minerals staff will perform ore feed blending using its own wheel loader from the RoM (run-of-mine) ore stockpiles into the crusher, with two operators covering different shifts (as required). Diesel generators will power sump pumps required in the pit. Two mobile light plants have been included.

I-Minerals will perform its own routine light vehicle equipment maintenance, while other maintenance will be out-sourced through a contract with a local equipment dealership.



17 Recovery Methods Section 17 summarizes operation results, processing methods, a conceptual flowsheet, plant design, and plant consumables.

17.1 Operation Results Operational results for the recoveries of the Bovill Kaolin clays, potassium feldspar, and quartz fractions are summarized below. A material balance for the process was developed to determine product recoveries and material losses. This material balance was based on the extensive laboratory and pilot plant scale testing studies that were performed and described in detail in Section 13 of this document. A simplified material balance for the process circuit is shown in Figure 17.1



Source: R&S/KBR, 2012

Figure 17.1: Simplified Material Balance Block Diagram



17.1.1 Yield/Recovery Results The data criteria used to develop a material balance was based primarily on the pilot-scale testing performed at GMT and MRL. This information was discussed in detail within Section 13.

Clay Recovery

It was determined that approximately 30% of the feed reported to the clay separation circuit and the remaining material reported to the feldspathic sand circuit. Following grit removal and fractionation of the two clays, about 22.6% of overall recovery was attributed to the clay (based on overall feed). Of this 22 .6% clay recovery, almost 22% yielded halloysite product (about 7% overall recovery) and the rest yielded kaolinite product (about 15.6% overall recovery).

K-spar and Quartz Recovery

It was determined that of the material feeding the feldspathic sand circuit approximately 18% was recovered as K-spar product (meeting acceptable product quality specifications). Over 50% of the feldspathic sand circuit feed was recovered as quartz products (Quartz 1, 2, and 3 based on acceptable product quality specifications). Incorporating a marketing product matrix for the desired three quartz products, the yields for each were as follows: Quartz 1 at 25.2%, Quartz 2 at 14.4%, and Quartz 3 at 11.1%. Overall recoveries based on ore were 17.6%, 10.1%, and 7.8%, respectively. It should be noted these yield and recovery values were based on yearly operating criteria developed through the desired product matrix – not hourly – due to the fact that actual hourly production operating criteria allows only for either Quartz 1 and Quartz 2 production or Quartz 1 and Quartz 3 production.

17.1.2 Material Balance Criteria Tonnage rates for the material balance were determined by incorporating several basic design criteria factors. These factors included setting fixed operating time parameters, setting desired annual tonnage requirements for the products being produced, and incorporating the average yield/recovery data discussed in Section 17.1.1.

The basic fixed operating parameters used for the development of the material balance were as follows: 24 hours per day, 7 days per week, 50 weeks per year, and a 93% operating efficiency. This equated to a conservative 7,812 hours per year.

The finished product matrix originally developed for the design of the material balance and subsequent circuit/equipment sizing utilized a couple of specific product “drivers” (halloysite and K-spar). Halloysite production was set at a minimum of 15,000 t/y for the clay circuit, while a minimum 28,000 t/y was used for the feldspathic sand circuit. The remaining product production rates were then determined by the resulting material balance criteria.

17.2 Processing Methods Processing of the Bovill Kaolin clays, potassium feldspar, and quartz fractions, as envisioned by I-Minerals, for the Bovill Kaolin deposit is described in this section. A description of the proposed major unit operations, pictorial flow diagrams (Figures 17.2.1 and 17.2.2), and equipment with functional duty specifications are also included for the proposed Bovill Kaolin processing operations.




Figure 17.2.1: WBL Primary Clay Process Simplified Flow Diagram




Figure 17.2.2: Feldspar and Quartz Process Simplified Flow Diagram



17.2.1 General Considerations The mineral leases held by I-Minerals are underlain by the Thatuna granodiorite, a source of feldspar with significant sodium oxide (Na2O) component. The weathering of granitoid ore bodies formed residual (or primary) deposits of kaolin, which also contain K-feldspar and quartz. The clay contains both kaolinite and halloysite (aluminosilicate clay minerals) with the empirical formula Al2Si2O5(OH)4.

17.2.2 Unit Operations Major unit operations in the proposed integrated Bovill Kaolin Processing circuit are summarized below:

Ore Processing

• Stockpile area for run-of-mine (RoM) ore; • Crusher for size reduction of RoM ore; • Attrition Scrubbing of crushed ore with dosed dispersant; • Classification of Attrition Scrubber product in Screw Classifiers;

• Derrick Hi-Cap dewatering screen for Screw Classifier underflow;

• Feldspathic Sands recovery as Hi-Cap dewatering screen oversize;

• Derrick Linear motion Sizing Screen for overflow from Screw Classifier; and

• Tailings (+325 mesh) recovery as Linear Motion Sizing Screen oversize.

Halloysite/Kaolin Clay Processing

• Diluted linear motion Sizing Screen underflow to 3 inch diameter Cyclones; • Fine Grit or 3 inch Cyclone underflow to tailings; • 3 inch Cyclone overflow to fractionation Solid Bowl Decanter Centrifuges (two units, double-

pass); • Fractionation Centrifuge second pass overflow (Halloysite) to dewatering Solid Bowl

Decanter Centrifuge; • Halloysite dewatering Centrifuge effluent (decant) to process for reuse; • Halloysite dewatering Centrifuge cake slurry (concentrate) to Filter; • Halloysite Clay filter cake from Filter to Dryer; • Dried Halloysite product to packaging; • Underflows (Kaolin) from both fractionation Centrifuge passes to dewatering Solid Bowl

Decanter Centrifuge; • Kaolin Centrifuge effluent (decant) to process for reuse; • Kaolin Centrifuge cake slurry (concentrate) to Filter; • Kaolin Filter Cake from Filter to Dryer; • Dried Kaolin Clay product to packaging; and/or • Dried Kaolin Clay to Calciner; and • Metakaolin Clay product (dehydroxylated kaolin clay) to packaging.

Feldspathic Sand Processing

• Stockpile area for dewatered Feldspathic Sands; • Rod Mill grinding of Feldspathic Sands;



• Classification of ground Feldspathic Sands using Hydrosizer; • Attrition Scrubbing of classified Feldspathic Sands with dosed dispersant; • Desliming of Feldspathic Sands using Hydrocyclone; • Conditioning of Feldspathic Sands with flotation reagents; • Iron Flotation with flotation reagents; and • Desliming of Iron Flotation Tailings using Hydrocyclone.

Potassium Feldspar Processing

• Conditioning of Iron Flotation Tailings with flotation reagents; • Potassium Feldspar (K-spar) Flotation with reagents; • Dewatering of K-spar product using Horizontal Vacuum Belt Filtration and Fluid Bed Dryer; • K-spar product final upgrading using Rare Earth Magnetic Separators; and • K-spar product to fine grinding using Air-swept Stirred Media Mill/Air Classifier and

subsequent packaging.

Quartz 1 Grade Processing

• Desliming of Feldspar Flotation Tailings using Hydrocyclone; • Regrinding of Feldspar Flotation Tailings using Rod Mill; • Classification of ground Feldspar Flotation Tailings using Hydrosizer; • Attrition Scrubbing of classified Feldspar Flotation Tailings with dosed dispersant; • Conditioning of Feldspar Flotation Tailings with flotation reagents; • Quartz 1 Grade Flotation with reagents; • Dewatering of Quartz 1 Grade Product using Horizontal Vacuum Belt Filtration and Fluid Bed

Dryer; • Quartz 1 Grade Product final upgrading using Rare Earth Magnetic Separators; and • Quartz 1 Grade Product to packaging or fine grinding using Air-swept Stirred Media Mill/Air

Classifier and packaging.

Quartz 2/3 Grade Processing

• Conditioning of Quartz 1 Grade Product with flotation reagents; • Iron Flotation with reagents; • Desliming of Iron Flotation Tailings using Hydrocyclone; • Conditioning of Iron Flotation Tailings with flotation reagents; • Quartz 2 Grade Flotation with reagents; • Dewatering of Quartz 2 Grade Flotation tailings product using Horizontal Vacuum Belt

Filtration and Fluid Bed Dryer; • Quartz 2 Grade Product final upgrading using Rare Earth Magnetic Separators; • Quartz 2 Grade Product to packaging or fine grinding using Air-swept Stirred Media Mill/Air

Classifier and packaging, or • Desliming of Quartz 2 Grade Flotation tailings using Hydrocyclone; • Conditioning of Quartz 2 Grade Flotation tailings with flotation reagents; • Quartz 3 Grade Flotation with reagents; • Dewatering of Quartz 3 Grade Flotation tailings product using Horizontal Vacuum Belt

Filtration and Fluid Bed Dryer;



• Quartz 3 Grade Product final upgrading using Rare Earth Magnetic Separators; and • Quartz 3 Grade Product to packaging or fine grinding using Air-swept Stirred Media Mill/Air

Classifier and packaging.

17.2.3 Additional Tests & Recommendations I-Minerals, in support of process optimization, are engaged in a number of activities:

• Continue evaluating alternative equipment options to improve the clay circuit both by product quality and reduce overall capital equipment costs. These options include centrifuge optimization and alternative separation techniques to improve halloysite quality, pH adjustment and flocculation for clay filtration, and dryer optimization;

• Methods of packaging the products are dictated by the markets. The most common form of packaged kaolin is the 50 lb (Kraft) bag. Kaolin clay is also supplied in dry bead and dry powder form in supersacks (bulk bags) and fiberboard containers. Halloysite nanotubes will be packaged per customer requirements yet to be determined; and

• Continue product development of halloysite, metakaolin, K-spar, and quartz products per market development requirements.

17.3 Flowsheet Two pictorial flowsheets are provided for the Project as shown in Figures 15-2 and 15-3.

17.4 Plant Design and Equipment Characteristics The basics of the plant design and pertinent equipment characteristics are discussed in this section. Throughput for the plant is shown in the preliminary material balance block flow diagram (Figure 15-1).

17.4.1 Ore Preparation The primary clay from the Bovill Kaolin mine area, which is projected to be the ore going to the processing facility, will be transported and stockpiled at the plant site and fed through a size reduction circuit utilizing a crusher system. The majority of the primary clay contains a large amount of fines and is quite friable. The material will then be stockpiled for classification.

17.4.2 Clay/Feldspathic Sands Classification The classification circuit to separate the feldspathic sands and clay consists of two-stage blunging (attrition scrubbing) and two-stage screw classification. Dispersant is added to each blunger to aid in the clay separation. The feldspathic sands underflow from the first screw goes to a second pass blunger and screw classifier to remove additional clay. The overflows proceed to a continuation of the clay separation circuit while the second pass screw classifier underflow is prepared as feed for the feldspar/quartz mill. This underflow fraction is fed to a Derrick Hi-Cap screen for dewatering to approximately 90% solids and is stockpiled for feeding the mill.

17.4.3 Degritting The screw overflows are transferred to the Derrick Linear Motion sizing screens to make a 325 mesh cut. The +325 mesh retains report to the Derrick Hi-Cap screen, while the underflow passing fraction



reports to a holding/mix tank. The -325 mesh material is intensely mixed and then fed through a bank of Krebs 3 inch cyclone circuit.

Each 3 inch cyclone is capable of a feed rate of approximately 31 gpm at an inlet pressure of 8 to 10 psi. A distribution feed manifold system is used to uniformly maintain constant and consistent feed variables. The 3 inch cyclones effectively make about a 20 μm cut. The overflow fraction (-44+20 μ) reports to waste, while the underflow clay fraction reports to a holding/mix tank.

17.4.4 Halloysite/Kaolin Clay Separation The clay-bearing cyclone overflow is again intensely mixed and fed to a two-stage solid bowl decanter centrifuge circuit for clay fractionation or separation into separate halloysite and kaolin fractions. The overflow from the first stage centrifuge reports to another holding/mix tank and then passed through the second stage centrifuge. The underflows from both stages are combined and report to a holding/mix tank for processing as kaolin clay. The overflow from the second stage reports to a separate holding/mix tank for processing as halloysite clay. Each stage centrifuge is capable of handling the entire circuit flow.

17.4.5 Halloysite Clay Product The halloysite clay slurry is fed to a centrifuge for initial dewatering. The effluent is recycled as process water while the underflow reports to a mix tank where sulfuric acid (H2SO4) is added to lower the pH and flocculate the solids. The flocked solids are then filtered to produce high-solids cake for feeding the dryer. The filter effluent reports to a separate recycle tank system for pH adjustment and reuse as process water. The filter cake reports via a feed conveyor to a dryer. The dryer discharge is pneumatically conveyed to a holding bin for packaging. Packaging will probably consist of supersacks, fiberboard drums, or Kraft bags.

17.4.6 Kaolin Clay/Metakaolin Products Similar to the halloysite clay circuit, the kaolinite clay material is fed to a centrifuge for initial dewatering. Again, the effluent is recycled as process water while the underflow reports to a mix tank where H2SO4 is added to lower the pH and flocculate the solids. The flocked solids are then fed to filters to produce high-solids cake for feeding the dryer. The filter effluent reports to the same recycle tank system for pH adjustment and reuse as process water. The filter cake reports via a feed conveyor to a dryer. The dryer discharge is pneumatically conveyed to a holding bin where it is either packaged as kaolin clay product (assuming similar packaging requirements as the halloysite) or conveyed to a calciner for further product processing.

The calciner operates at a temperature of approximately 900°C to produce a pozzolanic clay material known as a metakaolin. The metakaolin product is discharged and pneumatically conveyed to a holding bin, packaged in supersacks or Kraft bags.

The circuit handling dried kaolin will have built-in flexibility to handle either kaolin as a product for packaging and/or to calcine and package a portion of the kaolin clay into metakaolin product.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 132 17.4.7 Feldspathic Sand Grinding, Sizing, and Attrition Scrubbing

The stockpiled feldspathic sand is fed onto a transfer conveyor, sampled, weighed on a double-idler belt weigh scale, and fed to a rod mill. The ¼ inch rod mill trommel discharge is fed to a hydrosizer. The -30 mesh overflow from the hydrosizer is classified through a hydrocyclone.

The hydrocyclone overflow is pumped to the pre-flotation recycle water tank/clarifier system and the underflow is mixed with dispersant sodium hexametaphosphate [Na2(PO3)4] in a two-stage attrition scrubber. The scrubbed product is deslimed in a hydrocyclone.

17.4.8 Iron and K-spar Flotation The diluted hydrocyclone underflow at 20% solids is conditioned with sulfuric acid, amine, fuel oil, and frother as feed to a bank of four iron (mica) flotation cells. The “concentrate” floated from the iron flotation circuit consists predominantly of biotite (an iron-rich silicate mineral within the mica group), which is pumped to the tailings tank.

The feldspar and quartz-rich “tailings” are floated in a bank of six feldspar rougher flotation cells followed by a bank of four feldspar cleaner flotation cells. The tailings fractions from both banks are the feed for the quartz circuits, while the concentrate from the cleaner cells is the K-spar fraction.

17.4.9 K-spar Product Processing The feldspar concentrate is dewatered on a horizontal belt vacuum filter and the filter cake is dried in a fluid bed dryer. The dried product is passed through triple-pass rare earth magnetic (REM) separators and the magnetic residues and tailings are pumped to the tailings thickener. The magnetics-free K-spar product is stored in bins for fine grinding or packaging for shipment to markets. Fine grinding is accomplished using an air-swept stirred media mill and air classifier system in closed-loop. The fine ground K-spar product is then packaged for shipment to markets.

17.4.10 Quartz Regrinding, Sizing, and Attrition Scrubbing The tailings from the feldspar flotation circuit are sent to another rod mill and re-ground. A hydrosizer is again used and the -50 mesh overflow is classified through a hydrocyclone. The hydrocyclone overflow is again dispersed in a two-stage attrition scrubber.

17.4.11 Quartz Flotation The scrubbed product is deslimed through a hydrocyclone and the hydrocyclone underflow conditioned with hydrofluoric acid, amine, fuel oil, and frother and floated in a bank of six flotation cells to recover a tailings fraction considered to be Quartz 1 grade material. The feldspar-rich concentrate is recycled to the feldspar flotation circuit.

A portion of the Quartz 1 grade fraction is processed further into either Quartz 2 grade or Quartz 3 grade materials. The Quartz 1 grade fraction is conditioned with hydrofluoric acid, amine, fuel oil, and frother, and refloated in a bank of two flotation cells to complete another iron flotation step. The iron rich concentrate is pumped to the tailings tank, while the tailings are again conditioned and “scavenger” floated in a bank of six quartz flotation cells. The concentrate is recycled to the feldspar flotation circuit while the tailings are processed as Quartz 2 grade material or subjected to further flotation. The Quartz 3 grade material is achieved by performing one more conditioning and scavenger flotation step as described for the Quartz 2 grade.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 133 17.4.12 Quartz Products Processing

Quartz 1 grade material is dewatered on a horizontal belt vacuum filter and the filter cake is dried in a fluid bed dryer. The dried product is passed through triple-pass REM separators and the magnetic residues and tailings are pumped to the tailings thickener. The magnetics-free Quartz 1 product is stored in bins for fine grinding or packaging for shipment to markets. Fine grinding is accomplished in the same manner as the K-spar product.

Quartz 2 and Quartz 3 grade materials are individually processed through a separate circuit in the same manner as described for the Quartz 1 grade.

17.5 Consumable Requirements Principal consumables for the Bovill Kaolin project include electricity, natural gas, process reagents, and equipment maintenance parts. The primary consumer of fuel is the mining contractor and is contained within the contractor’s budgeting.

Both electricity and natural gas will be provided by Avista Utilities. Avista has been informed and kept updated on the project for quite some time, and has provided current cost data for the economic modeling.

Reagents for the project include specific flotation reagents for feldspar and quartz, dispersing reagents, flocculation reagents, and pH adjustment reagents. Suppliers will be regionally located with the probable exception of the flotation reagents. The flotation reagents are very specific to the flotation of feldspar and quartz, and the domestic supplier may not be regionally based.

The equipment maintenance parts are variable and will be dependent on detailed analysis of inventory and replacement requirements moving forward with the project. Accountability at this time has been accessed by applying a percentage to the purchased equipment cost estimate contained in the capital cost estimate.



18 Project Infrastructure 18.1 Site Access Road

Primary access to the Project will be via the Moose Creek Road. The mine site is approximately 5 miles north and west from the access point on Idaho State Highway 3, just west of Bovill. Access to the plant site is also via Moose Creek Road approximately 2.3 miles from State Highway 3. In order to eliminate traffic near Moose Creek Reservoir, a “reservoir bypass road” of approximately one mile will be improved north of Moose Creek Reservoir.

18.2 On-Site Services The Project, once completed, will include basic services commensurate to a facility of this type and size and will include the following:

• Administration offices facilitating human resources, purchasing, technical services; • Warehousing for parts and supplies; • Mine and plant repair shops (the repair shop will be constructed and operated by the mining

contractor); • Laboratory; • First aid facility; • Electrical substation; • Process and firewater storage tanks; and • Water treatment facilities.

The processing facility will encompass approximately 4.0 ac of storage areas for pre-processing ore. Post-primary crushing material will cover approximately 1.5 ac. Post-processing storage of products will be in silos and/or warehouses that will collectively cover approximately 10,000 ft2 (0.25 ac). Pre-shipment storage of packaged product in warehouse facilities will cover approximately 0.75 ac. Offices, facilities, and workstations for support services and staff will cover approximately 5.0 ac.

18.3 Energy Supply

18.3.1 Electric Power Power is available to the Property from hydroelectric power lines that extend from Moscow to Bovill and will be available for mining and processing requirements.

The Project complex will have a connected load of 6.5 MW. The largest loads required will be 0.5 MW for the Primary Rod Mill and 0.3 MW for the Quartz Fine Grinding System.

This power will be purchased from Avista Corp and will be received at its substation on transmission lines owned by Avista Corp. Substation transformers will have a capacity of 11.0 MVA, and reduce voltage from 138 kV to 13.8 kV for on-site distribution.

18.3.2 Natural Gas Natural gas is available to the Property from a natural gas pipeline that extends from Moscow to Bovill and is available to be utilized for this processing facility.



Natural gas is primarily used for Rotary Dryers and the Calciner for the Clay circuits, Fluid Bed Dryers for the feldspar circuits, and multiple HVAC units for general facility heating during colder temperatures at the plant. Natural gas is purchased from Avista Corp and supplied to the Project via a gas pipeline owned and operated by Avista Corp. The annual consumption of natural gas is calculated to be 3,300,000 therms per year based on the equipment needs for the plant design. Unit costs were given directly by Avista Corp for this Project.

18.4 Water Water needed at the processing plant can be provided by ground water wells at the processing plant and a reservoir north of the mine site. Despite a considerable amount of in-circuit flow, the majority of the water is recycled within the various plant circuits. The amount of water consumed in the processing plant is approximately 170 gpm. All the water within the circuits is recycled (including collection ponds) or lost due to drying or evaporation, and therefore no water is discharged to waters of the U.S. or State. The water balance within the plant is shown in Figure 18.4.1




Figure 18.4.1: Simplified Water Balance Block Diagram



18.5 Tailings Storage Facility (TSF) The tailings management plan, which includes construction of a new Boyle Tailings Storage Facility (TSF), was designed by Tetra Tech. The details of the Boyle TSF design along with estimated construction and operating costs are contained in the report prepared by Tetra Tech which presented a detailed tailings management plan in support of a Canadian National Instrument NI43-101 compliant Pre-feasibility Study for the proposed Bovill Kaolin project (Tetra Tech, 2014).

The proposed Boyle TSF is designed to accommodate approximately 2.4 Mt of thickened tailings from the processing plant and will handle tailings up to a nominal rate of 87,000 t/y with an assumed average tailings dry density of 90 pcf. Currently the tailings facility will operate for approximately 25 years and will be required to handle tailings at an annual rate of 78,000t/y or a total of 1.9 Mt of thickened tailings. Table 16.5.1 summarizes the design criteria adopted for the TSF design.

Table 18.5.1: Boyle TSF Design Criteria Design Parameter Value Design Tailings Storage Capacity 2.4 Mt Average Tailings Solids Content 57% Average Tailings Dry Density 90 pcf Tailings Production Rate 78,000 t/y Design Life 25 years Source: TetraTech, 2012

The ultimate TSF embankment rises at a 2H:1V (2 Horizontal units to 1 Vertical unit) slope to a maximum elevation of approximately 3,040 ft above mean sea level. The construction of the embankment will occur in three stages utilizing a modified centerline method. The tailings characteristics and deposition plan are assumed to result in a beach slope of approximately 0.5 percent for the Boyle TSF. The Boyle TSF will be lined with a compacted 3 ft clay liner to inhibit seepage from the facility. The TSF includes an overliner drain system to enhance tailings dewatering and consolidation and reduce hydraulic head on the liner system. The TSF design includes an underdrain system to reduce uplift groundwater seepage and a toe drain system to enhance stability.

Three collection ponds are present immediately downstream of the TSF embankment to collect water from the overliner drain and toe drain systems. This water will be carried through a reclaim pipeline to the plant for reuse. Underdrain water will be collected at a sump, checked for water quality, and released downstream. An access road is also designed to the east of the facility for the slurry pipeline and for the reclaim water pipeline to the plant.

18.6 Product Transportation Products produced at the plant will primarily be packaged into 2,000 lb bulk bags or smaller 50 to 100 lb Kraft bags per customer requirements and palletized. The products will then be loaded onto tractor-trailer trucks for transport to customers. Some product will be loaded into bulk tanker trucks for transport to customers. All material will leave from the plant site.



19 Market Studies and Contracts 19.1 Product Markets

Products recovered by the Bovill Kaolin process operations will include:

• Halloysite clay, an aluminosilicate with hollow tubular morphology in the submicron range; • Kaolinite clay, hydrated aluminum silicate; • Metakaolin clay, refined kaolin clay (aluminum silicate) or dehydroxylated kaolin clay

calcined to create amorphous aluminosilicate, which is reactive (Pozzolan) and enhances the strength, density and durability of concrete;

• K-Spar, also known as potash feldspar uniquely suited ceramic formulations requiring an alumina source; and

• Quartz, SiO2 silicon dioxide.

The following discussion is supported through an independent market study by Charles River Associates of Boston, MA, (CRA) dated December 2011. Furthermore, I-Minerals continues to study potential markets for its products as is common in the industrial minerals industry.

19.1.1 Halloysite Halloysite has historically been used in the manufacture of porcelain, bone china, and fine china due to purity of the clay and the low iron and titanium content resulting in ceramic ware with exceptional whiteness and translucency. Fine particle size enables halloysite to also be used as a suspension agent in glaze preparations as well as in filters and inkjets, and as an ingredient in special paints applied to ships to prevent barnacles from growing on the ships' hulls.

The largest supplier of commercial halloysite product available at present is located in Matauri Bay, New Zealand. Limited production is reported from Poland, Turkey and China, and a development stage project is located in Utah with negligible commercial production. I-Minerals’ halloysite is differentiated from those known halloysite deposits because of its high aspect ratio and minimal trace elements.

Halloysite nanotubes (HNTs) are considered a high value product. The high aspect ratio (tube length to diameter) of the tubes allows the tubes to be filled or coated with substances to achieve a wide variety of electrical, chemical, and physical properties. The markets interested in filling halloysite hollow tubes with active ingredients include cosmetics, household and personal care products, pesticides, paint, pharmaceuticals and the life sciences market.

I-Minerals has received inquiries from the following industries: personal care products, Nano-composites, fire retardants, biocides, plastic fillers, animal feed, paint, ceramics, two undisclosed R & D projects and several universities.

19.1.2 Kaolin I-Minerals kaolin characterization is in development stage. Market analysis indicates several high-end applications such as metakaolin (Section 19.1.3), plastics and ceramics which are of interest to the western North American market. Also within the western North America market mid- and low-end applications will be evaluated such as paint, fiber glass, catalysts, refractories, roofing granules, and



bricks. Initial testing has determined the company’s kaolin product fires white. Interest from end-users will be followed with samples which are highly anticipated by ceramic, paint, and brick companies whose geographic proximity to the deposit will allow transportation savings as opposed to receiving kaolin from southeastern United States.

19.1.3 Metakaolin Metakaolin is considered a premium material adding extra strength and durability. The market requires a consistent product and the ability to meet prescribed ASTM standards. For clarification, ASTM C 618 is the gateway to the metakaolin market. The company’s metakaolin was determined to be a pozzolan metakaolin product through meeting ASTM C 618-08: Chemical and Physical Analysis of a Natural Pozzolan and Chemical and Physical Analysis of Fly Ash. The company’s metakaolin also shows significant ASR mitigation capabilities meeting ASTM C 1567 Alkali Silica Reaction standard.

Direct contact with western states’ Departments of Transportation and Ready Mix companies reveal a market for metakaolin both in paving and bridge cap repairs. The company’s geographic location will afford significant transportation savings and an opportunity for use of a once-too-expensive product in the Western United States and Canada. Metakaolin is valued for ASR mitigation, particle size, and low water requirement which mean higher quality—all descriptive of the company’s metakaolin test results.

19.1.4 Feldspar Potassium feldspar (K-feldspar) is primarily used in ceramic bodies and glazes. The company notes strongest market potential for its K-feldspar in Western North American markets; however, Midwest and Eastern North American ceramic, tableware, tile, and sanitary ware manufacturers have expressed an interest in the low iron, high alumina chemical composition of MC K-Feldspar. The chemical quality of the company’s product is not available in the North American market today by other feldspar producers. Samples have been distributed widely; the product performs to expectations in testing by end users.

K-feldspar is also used in the coating industry as functional filler; samples have been submitted to a major coatings manufacturer. K-feldspar is valued for the physical properties of dispersability, weatherability, scrubability, and mildew resistance in industrial and marine coatings.

19.1.5 Quartz The chemistry of I-Minerals’ quartz products ranging from SiO2 of 99.8% to 99.97+% (high purity and low iron) gained immediate attention within the glass industry as the results were announced in news releases and the website. Manufacturers expressing an interest are from North America and the Pacific Rim region from several high-end markets: LCD glass, solar glass, sodium silicate, and lighting manufacturers to low-end container and float glass markets. Melt tests in second quarter 2014 are arranged. Manufactures of glass fiber, art glass, ceramic applications, and paint add to the interested users of 99.8% SiO2 product. High-end quartz product of 99.97+% commands an attractive price within the specialty quartz market.



19.2 Commodity Price Projections Prices for sand and clay products can vary dramatically depending upon the specifications and quality each product produced. Due to the highly competitive nature of the industrial sand and clay industry, contract prices are highly confidential and are not presented in public documents.

However, the QP for this PFS confirms that I-Minerals has, in fact, completed a market study for its products which included preliminary negotiations to supply a series of clay and sand products. The QP also confirms that the process facility is capable of producing these products. While negotiated prices are confidential, composite market prices are presented in Table 19.2.1.

Table 19.2.1: Commodity Prices

Description Price (US$/t-product) Halloysite 1,200.00 Halloysite - incremental 300.00 Kaolin 95.00 Metakaolin 225.00 K-Spar (sliding scale first 4 years) 220-250 Quartz 248.00 Quartz-incremental 87.50 Source: I-Minerals

19.3 Contracts and Status I-Minerals has received preliminary inquiries for clay and sand products from interested buyers; no contracts are currently in place.

19.3.1 Terms Mineral products will be sold FOB at the Bovill Processing Plant along with product shipped and delivered to buyers in the area.



20 Environmental Studies, Permitting and Social or Community Impact This section summarizes the environmental setting and status of environmental permits and studies required for construction, operation, and closure permits and plans. A number of state permits and approvals are required for operations. Plan of Operations and Reclamation Plan to address mitigation plans, impact analyses, and monitoring will be required to secure the major permits. Several other minor permits and approvals are required from County and State agencies. One year is estimated to secure the major permits after application materials are prepared.

20.1 Environmental Setting The Project area is located in the western foothills section of the Bitterroot Mountains Physiographic Province, Latah County, in north central Idaho. The Project area consists of low foothills and ridges alternating with relatively wide, flat basins. The average elevation of the mineral lease area is about 914 m (3,000 ft) above mean sea level, with a topographic relief of about 60 m (200 ft). The Project is located on the west side of the Potlatch River drainage area. The Potlatch River drains south and southwest to the Clearwater River and is part of the greater Columbia River system that flows to the Pacific Ocean at the Oregon-Washington border. In Bovill, approximately 3.2 km (2 miles) southeast from the Project area, the average annual precipitation is 95 cm (37.4 inches), and the average total annual snowfall is 262 cm (103 inches). The average annual maximum temperature is 13.7°C (56.6°F) and the average annual minimum temperature is -0.4°C (31.2°F).

Soils are shallow to moderately deep with loamy to sandy textures and usually contains volcanic ash. Most of the Project area has been used for mining, grazing, and/or timber harvest and therefore a relatively disturbed landscape is present. Evidence of past mining is found in borrow pits, tailings piles, as well as mine pits. Forested areas primarily occur on slopes and ridge tops and are intensively managed for timber production. Forests in the Project area are composed of Douglas fir, grand fir, and western red cedar. Open wet meadows, occurring primarily in the basins along intermittent and perennial stream channels, are dominated by grassland with intermittent shrubs.

The Project area is located near the headwaters of Moose Creek, in the Potlatch River watershed. The Potlatch River originates northeast of Bovill, ID in the Beals Butte area and runs for 90 km (56 miles) in a southwesterly direction through the southern half of Latah County with roughly 3,000 km (1,900 miles) of tributary streams. The Project area is drained by several branching ephemeral streams. Springs in the area are both ephemeral and perennial.

20.2 Environmental Studies and Permitting The following describes the current results of environmental studies, remaining studies and/or analyses, and permitting requirements.

Clean Water Act Section 404 – Wetlands

A wetland and ordinary high water mark (OHWM) delineation of the Project area was conducted in the summer and fall 2012. The delineation efforts focused on areas of potential land disturbance associated with roads, processing facilities, pits, and tailing areas. The goal of the delineation was to map the boundaries of wetlands and other waters of the U.S., and to provide this information to the



project design team so that disturbance of these areas could be avoided. As such, if there is no fill or disturbance to wetlands or other waters of the U.S., no CWA Section 404 permit is required. A wetland delineation report will be submitted to the USACE for a formal jurisdictional determination. Following this determination, additional changes to Project facility layout may be required in order to avoid wetlands and other waters of the U.S.

Idaho Mine Operation and Reclamation Plan

The Idaho Surface Mining Act, Title 47, Chapter 15, Idaho Code requires the operator of a surface mine to obtain an approved reclamation plan and bond. The Project is on state leased land and as such no federal mine permit approval is required.

The IDL is the lead agency for surface mine activity in Idaho and requires an approved operating plan and reclamation plan prior to mining. In addition, IDL is the lead agency for implementing the anti-degradation policy for surface mining and may solicit comments from other resource agencies in reviewing and approving the permit including Idaho Department of Fish and Game (IDFG), Idaho Department of Water Resources (IDWR), and Idaho Department of Environmental Quality (IDEQ). It is anticipated that these agencies will be requested by IDL to participate in the mine permitting process.

Work conducted to date includes pre-feasibility level design of the tailings facility, processing plant, and quarry areas that includes standard water management control features.

Remaining tasks include completing the analysis of potential impacts to ground and surface waters and development of a plan to maintain compliance with surface and groundwater quality standards, as needed. In addition, a reclamation plan will be required that identifies, evaluates, and provides mitigation for areas disturbed by mining. This information, along with the rest of the Project plans will be submitted to IDL as the Bovill Kaolin Mine Operation and Reclamation Plan Application. IDL, with support from resource agencies, will review and provide subsequent comment on the application and, once any outstanding issues have been resolved, will provide mining authorization.

Idaho Air Quality Permit

In Idaho there are three types of air quality permits: Permits to Construct (PTC), Tier I Operating Permits, and Tier II Operating Permits. Air quality permit requirements are addressed under IDAPA 58.01.01 – Rules for the Control of Air Pollution in Idaho. The PTC program is required for new or modified sources. PTC permits do not expire unless construction has not begun within two years of its issue date or if construction is suspended for one year. Tier II Operating permits are issued to facilities when IDEQ has determined that a facility needs an air quality permit to comply with applicable rules, or when an applicant has specifically requested one. The most common type of Tier II operating permit that IDEQ issues are those which the applicants have requested in order to establish synthetic minor emission limits. For the construction of a new source it is a PTC that is required. A Tier II operating permit is generally developed after site startup (often one year after operations begin). A Tier I operating permit (also known as Title V permit) is required for all major sources of air pollution. Tier I Operating permits are required for major sources even if the facility already has a PTC. Tier I permits expire within five years.

An assessment of emission sources (emissions inventory) for roads, tailings, and the processing facility must be completed. Limited stack emissions are typically associated with this type of project and can be controlled through air pollution control devices. Tasks remaining to be completed include



an emissions inventory for the Project, modeling these emissions, and preparation of the PTC that includes expected impacts, minimization measures, and any monitoring, as needed.

No significant air quality issues have been identified to date. This is primarily due to the limited quantity and composition of stack emissions associated with a project of this type. Emissions sources include emissions from gas-fired dryers, generators, diesel truck operation, and fugitive dust. Also, while fugitive dust can be a significant concern, use of mitigation measures (e.g. proposed dust collectors, material wetting, and other controls) and presence of few receptors in the Project area reduce this concern.

Clean Water Act Section 402 NPDES

Construction and operation activities associated with the Project will be covered by the 2012 NPDES General Permit for Discharges from Construction Activities and the NPDES MSGP for Stormwater Discharges Associated with Industrial Activity. Both are general permits that require a NOI to be filed and the development of a SWPPP. The general construction permit covers temporary activities such as facility structure construction, while the MSGP covers on-going mine related activities such as overburden removal and general mining activities.

A preliminary assessment of stormwater discharges has been completed. Issues identified include construction stormwater runoff, and potential discharge of contaminants from runoff from waste dumps and tailings. Based on the processing plant operation plan, no direct effluent discharges to waters of the US would occur so an individual NPDES permit for process water will not be needed. Remaining tasks to be completed include final identification of water balance and management plans for all runoff sources. This will include development of a site-wide SWPPP for construction stormwater and for the MSGP.

The NPDES stormwater general permits described above are the only identified environmental federal permits associated with the Project. As a federal action, these general permits have requirements for complying with the Endangered Species Act, Section 106 of the National Historic Preservation Act, and Clean Water Act Section 401 Certification. Each of the items is addressed below.

Endangered Species Act - In order to obtain coverage under both stormwater general permits, the applicant must avoid adverse effects on federally listed threatened and endangered species from the stormwater discharges and discharge related activities. An updated evaluation of endangered species will need to be conducted at the time of submitting the NOI for the stormwater permits. Threatened and endangered species or critical habitat identified for Latah County based on review of the current list (June 10, 2014) from the U.S. Fish and Wildlife Service (USFWS) are:

• Canada lynx (lynx canadensis) – Threatened; • Spalding’s catchfly (silene spaldingii) – Threatened; and • Water howellia (howellia aquatilis) – Threatened.

A formal hair snare survey was conducted from March 29, 2011 through April 27, 2011 to determine the presence or absence of the Canada lynx with the general Project area (Tetra Tech, 2011). A combination of remote cameras and hair snares were used to monitor lynx activity within and around the general Project area. Hair snare surveys were performed in general accordance with the survey methods used during the National Interagency Canada Lynx Detection Survey in Minnesota, Wisconsin, and Michigan (Burdett, et al., 2006). Due to the distance from the nearest Canada lynx



critical habitat, large amount of human disturbance and year round use, relatively small amount and poor quality of usable habitat, and results of the 2011 hair snare surveys, the Project area does not appear to support Canada lynx or its critical habitat.

The Spalding’s catchfly prefer Palouse prairie habitat, which does not exist in the Project area. During wetlands surveys of the project area in 2008 and 2012, no habitat for water howellia individuals was found.

Several federally listed fish species and essential fish habitat occurs in the Potlatch River drainage, which is over 4 miles from the Project site. Impacts from proposed site activities on the Potlach River is unlikely due to the distance between the Project and the occupied habitat, as well as the fact that best management practices through implementation of the SWPPP would help eliminate or substantially reduce any indirect effects from water quality impacts.

Section 106 of the National Historic Preservation Act – Under a federal stormwater general permit, the applicant is required to make certain certifications regarding the potential effects of stormwater discharge, and discharge related activities on properties listed or eligible for listing on the National Register of Historic Places. A study of the Project disturbance areas (mine quarries, access roads, processing facility, and tailings/dump areas) will be completed to assess for the presence/- absence of cultural resources. A study in the general Project area (but not at the exact locations of Project features) was completed by Dr. Lee Sappington of the University of Idaho in 2007. Findings from that study indicated no effect on listed or eligible historic sites. The State Historic Preservation Office (SHPO) concurred with this finding. Based on the similar conditions found in the Project area it is anticipated that there will be little, if any impact to historic resources associated with stormwater discharge and related activities. A cultural survey of the Project area will be conducted to support the NOI application for the general stormwater permits.

Section 401 Certification – The Clean Water Act requires state certification for any permit or license issued by a federal agency for an activity that may result in a discharge into waters of the U.S. The IDEQ issued Section 401 certification for construction general permit on April 3, 2012 and MSGP on December 5, 2008. The certifications identify monitoring requirements as part of the permit if stormwater is discharged to impaired waters. Monitoring and reporting requirements will be incorporated into the SWPPP.

Other Federal Permits and Requirements

As described above, the only identified federal permit is related to NDPES general stormwater permits and their requirements associated for the Endangered Species Act, Section 106 Historic Resources, and Clean Water Act Section 401 Certification. The goal for the Project is to avoid fill in wetlands and waters of the U.S., and as such, no Section 404 permit would be required. With no federal actions other than stormwater general permits, no National Environmental Policy Act (NEPA) evaluation, Section 7 consultation, or Section 106 consultation are necessary. If it is determined following more detailed mine design that fill in wetlands cannot be avoided, then a Section 404 permit and supporting NEPA evaluation would be required.

20.2.1 Required Permits and Status Table 20.2.1.1 summarizes the major permits for the project, as well as the required studies, impact analysis, potential mitigation and monitoring requirements, and permit status. It is premature to



develop a complete list of permits until the project design is more fully developed, however, the major permits are included in this table.

Table 20.2.1.1: Required Environmental Studies by Permit

Permit/Authorization Studies Impact Analysis

Mitigation Requirements Monitoring Status

Idaho Mine Operation and Reclamation Authorization (IDL)

Description of all mine operation and reclamation plans including tailings storage facilities and water quality management measures.

Analysis of engineering and potential water quality concerns.

Development of water quality management plans including any necessary engineering controls

As required in the Plan of Operations and Reclamation Plan

Final analysis and full mine, tailings, and site reclamation plans to be completed.

Air Quality (IDEQ)

An emissions inventory - Identification of emission sources including location and applied emissions factors for the Project

Modeling of emissions and identification of expected impacts

Preparation of Permit to Construct including any identified mitigation measures and monitoring needs

Monitor air as a condition of Permit to Construct

Emissions inventory, modeling, and permit preparation to be completed

Section 402 NPDES (EPA)(1)

Identification of all stormwater discharge sources. As part of NPDES permit, Engendered Species Act, Section 106 Cultural Resources, and Section 401 Certification are evaluated for stormwater discharges and stormwater related activities

Identification of material impacts from discharge sources to specific waters

Development of SWPPP including mitigation and monitoring plan

Monitor stormwater, site inspection, and implementation of best management practices for Project life

Identification of stormwater discharges, mitigation, and monitoring to be completed. NOI to be completed. SWPPP to be completed.

Source: HDR Engineering, 2012

(1) Except for general NPDES stormwater permits, no other environmental federal permits requirements have been identified.

In addition to the permit/authorizations described in Table 20.2.1.1, several other permits and approval will be required. A preliminary list includes:

• Water rights – obtained through the Idaho Department of Water Resources • Sanitary waste disposal – generally permitted through the North Central Health District and

coordinated with the IDEQ. • Conditional Use Permit – permitted through Latah County. • Drinking Water Supply – permitted through North Central Health District in coordination with

IDWR and IDEQ.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 146 20.2.2 Post-Performance or Reclamations Bonds

The Idaho Surface Mining Act requires I-Minerals to post financial assurances for reclamation with IDL prior to conducting any surface mining operations. Financial assurances will be in an amount sufficient to complete reclamation as described in the reclamation plan to be submitted by I-Minerals to IDL as part of the permitting process. When the IDL approves the reclamation plan, the agency will determine the amount at the time of permit approval. Mining may commence upon posting of financial assurances.

20.3 Social and Community Latah County had a population of 37,704 in 2011, with the population density of 35 people per square mile (U.S. Census Bureau, 2013). The county’s population grew 7% between 2001 and 2011, while Idaho’s population grew 20% and the U.S. population grew 9% (Idaho Department of Labor, 2013). Relatively slow job growth is the main reason for Latah County’s slow population growth.

According to a report from Idaho Department of Labor, the 2011 per capita income for Latah County was US$31,809 compared to a state average of US$32,881. In Latah County, approximately 10.2% of Latah County families are below poverty level between 2007 and 2011 (U.S. Census Bureau, 2013). Approximately 12% of the Bovill residents were reported to have an income below the poverty level in the year 2011.

Construction of the mine could create 30 to 50 new jobs and provide an economic boost to Latah County. The economic stability of the communities in Latah County would benefit by having the current workforce living in the communities and employed at the mine. In addition to the direct employment, indirect and induced employment will also be generated. The indirect and induced employment is that of suppliers to the mine and employment due to spending by employees of the mining operations.

Based on the current employment and wage report, Latah County’s average hourly wage for a construction laborer is US$14.41 generating yearly salary of approximately US$29,979 (Idaho Department of Labor, 2013). Hence, the direct income may increase at peak construction in the Bovill region. Additional temporary indirect income will be generated in construction supporting industries in nearby Bovill, ID.

The current housing market and community service facilities of the region would likely be sufficient to absorb the increase in the local population. Therefore, no increased demand on the housing market and on the community services is expected as a result of the temporary increases in population during construction and operational phase of the extraction pit.

Operation of the mine would result in the long-term employment opportunities as the mine is expected to be operable for 26 years. In addition to the direct jobs created by the operation of the mine, additional indirect jobs would be created in other supporting industries.

20.4 Mine Closure Mine closure will follow the IDL-approved mine operation and reclamation plan.



21 Capital and Operating Costs Pre-feasibility (PFS) capital and operating costs are based on preliminary mine and process design criteria, engineering, as well as budgetary quotes. Owner costs are estimated by SRK. Capital costs for process and project infrastructure are estimated by R&S/KBR, tailings impoundment costs are provided by Tetra Tech.

21.1 Capital Cost Estimates The capital costs presented are to a PFS-level of accuracy and are expected to be within ±25%. Contingency of has been applied to as appropriate to each area to account for exclusions in the estimate. All costs are in 1Q 2014 US dollars.

LoM capital costs are shown in Table 21.1.1. The initial capital estimate is US$72.7million. An additional US$18.2 million in sustaining capital will be required. Total LoM capital will therefore be about US$90.8 million.

Table 21.1.1: LoM Capital Cost Summary (US$000’s)

Description Initial Sustaining LoM Mining $919 $467 $1,386 Process & Infrastructure $59,538 $9,797 $69,335 Tailings Storage Facility $9,170 $4,900 $14,070 Owners Costs $3,056 $3,000 $6,056 Total Capital $72,682 $18,165 $90,847

Source: SRK

21.1.1 Mining Mining will be performed by a contractor and therefore the purchase of production equipment will not be required. Owner-related mine capital associated with mining are for minor support functions, which will not be performed by the mining contractor. A provision for mine equipment is included in the process cost estimate. Capital costs estimated in this section include provisions for site development and mine shop equipment. In addition, there is provision for a flatbed truck, two pickups, two light plants, and computer equipment and software.

Owner-related mine capital costs, as shown in Table 21.1.1.1 will initially be US$919,000 and will total US$1.4 million over the LoM. Mine-related support equipment is shown in Table 21.1.1.2.

Sustaining capital, totaling US$467,000 over the 25-year LoM provisions scheduled equipment replacements and software license renewals.



Table 21.1.1.1: Mine Owner Cost Summary (US$000’s)

Description Initial Sustaining LoM Mine Support Equipment 159 159 318 Waste Dump Construction 100 0 100 Buildings-Trailer and Break Room 15 0 15 Fuel Tank 9 0 9 Generator 12 0 12 Computer Equipment & Software 150 248 398 Misc. Equipment 0 0 0 Subtotal $799 $407 $1,205 Contingency 120 61 181 Mine Capital $919 $467 $1,386


Table 21.1.1.2: Support Equipment (US$000’s)

Description Units Unit Cost Total Cost Flat Bed 1 55 55 Pickups 2 30 60 Light Plants 2 22 44 Total

$159


21.1.2 Process & Infrastructure The estimated LoM capital cost for the process facility is about US$69.3 million as shown in Table 21.1.2.1.

Table 21.1.2.1: LoM Process & Infrastructure Capital (US$000’s)

Description Initial Sustaining LoM Equipment 49,600 0 49,600 Indirect Costs 9,938 0 9,938 Sustaining Capital 0 9,797 9,797 Process Capital $59,538 $9,797 $69,335


The basis for this cost is the capital cost estimate provided by R&S/KBR and shown in Table 21.1.2.2. Sustaining capital was estimated to be 2% per annum of the Purchased Equipment Cost from the R&S/KBR estimate.



Table 21.1.2.2: Process Capital Cost (US$000’s) Description Cost Purchased Equipment: Purchased Equipment 19,540 PEC Installation Labor 5,862 Instr. & Controls - Materials 977 Instr. & Controls - Labor 49 Piping - Materials 586 Piping - Labor 586 Electrical - Materials 3,908 Electrical - Labor 3,126 Structural - Materials 6,839 Structural - Labor 2,736 Civil & Concrete - Materials 1,563 Civil & Concrete - Labor 3,533 Yard Improvements 293 Mobile Equipment 19,540 Subtotal $49,600 Indirect Costs: Engineering Cost 1,954 Project Management 376 Construction Management 751 Construction Equipment Rental 150 Sales & Use Tax 1,172 Freight 586 Contractors Fee 819 Start-up Costs 195 Contingency 2,729 Mobile Equipment 1,136 Sales Tax 68 Subtotal $9,938 Sustaining Capital 9,797 Total $69,335


21.1.3 Tailings Storage Facility (TSF) The construction estimate was developed based on unit costs and construction rates provided by Mr. Lonnie Simpson of Debco Construction in 2012 and updated in 2014. The cost schedule was broken down on an annual basis for pre-production and the entire mine operation into closure.

Per Mr. Gary Nelson of I-Minerals, the operating cost estimate was based on 7,812 working hours per year with approximately 87,000 t/y of tailings. The current estimate indicated that an average rate of 78,000 t/y of tailings will be used.

Total LoM Capital for the TSF is listed in Table 21.1.3.1.

Table 21.1.3.1: TSF LoM Summary Capital Cost (US$000’s)

Description Initial Sustaining LoM Tailings Storage Facility 9,170 4,900 14,070 Total Capital $9,170 $4,900 $14,070

Source: Tetra Tech, 2014



Table 21.1.3.2 presents an estimate of the material quantities and unit construction costs anticipated for the major components of the Boyle TSF.

Table 21.1.3.2: TSF Capital Cost (US$000’s)

Description Units Quantity Unit Cost Cost Site Preparation Survey Lump 1 35,875 36 Stockpile Areas yd3 5,727 3.33 19 TSF Footprint yd3 170,064 3.33 567 Processing Ponds yd3 4,630 3.33 15 Access Road yd3 6,321 3.33 21 Topsoil Stockpiling yd3 176,385 4.10 723

subtotal - $1,381 Underdrain Installation Trench Excavation yd3 3,763 5.13 19 Geotextile Placement yd2 9,559 1.90 18 Gravel Placement yd3 3,708 30.75 114 Supply/Install Perf. Pipe lf 3,286 10.25 34

subtotal - $185 Liner Installation Impoundment (3’) yd3 147,229 4.61 679 Embankment (1’) yd3 34,975 5.89 206

subtotal - $885 Overdrain Installation Geotextile Placement yd2 9,559 1.90 18 Gravel Placement yd3 3,708 30.75 114 Supply/Install Perf. Pipe lf 3,286 10.25 34

subtotal - $166 Drain Transition Trench Excavation yd3 1,487 5.13 8 Geotextile Placement yd2 6,492 1.90 12 Gravel Placement yd3 1,454 30.75 45 Supply/Install Perf. Pipe lf 4,460 10.25 46

subtotal - $110 Embankment Channel Drains yd3 95,109 28.70 2,730 Structural Fill yd3 948,930 2.05 1,945 Clay Liner yd3 44,694 5.13 229 Rip Rap yd3 19,859 17.94 356 6” HDPE Toe Drain lf 2,234 10.25 23

subtotal - $5,283 Water Management Excavate Ponds yd3 15,137 3.59 54 Clay Liner yd3 2,315 5.13 12 LLDPE Liner ft2 187,504 2.05 384 Excavate Ditches yd3 398 3.08 1 Supply/Install Rip Rap yd3 221 17.94 4 Excavate Sump yd3 15 2.05 1 Supply/Install Sump Lump 3 5,638 17

subtotal - $473 Reclaim System TSF Pump (1) Lump 1 51,250 51 TSF Pipeline lf 3,700 37.41 138 TSF Pipeline Insulation Kit ea 93, 205 19 TSF Supply/Install Butterfly Valves ea 4 1,230 5 Pond Pumps (3) Lump 1, 83,080 83 Pond Pipeline lf 7,100 37.41 266 Pipeline Insulation Kit ea 178, 205 36 Pond Supply/Install Butterfly Valves ea 11 1,230 14



Description Units Quantity Unit Cost Cost Pipe Trenching yd3 820 5.13 4

subtotal - $616 Tailings Line Access Road yd3 8,119 5.13 42 Knife Gate Valves ea 12 10,250 123 Distribution Line lf 5,900 103 605 Insulation Kit ea 148 205 30 Drop Down Bars ea 12 15,375 185 Pig Launching/Receiving Piping (2) Lump 1 51,250 51

subtotal - $1,035 Closure Costs Clay Liner yd3 81,933 7.18 588 Topsoil yd3 245,799 6.15 1,512

subtotal - $2,100 Direct TSF $12,235

Contingency 15% $1,835 Total TSF $14,070

Source: Tetra Tech 2014

21.1.4 Owner’s Costs Most owner-related costs are included in the indirect cost estimate for process and infrastructure as well as closure cost for the TSF. However, provision for owner’s costs not included elsewhere is shown in Table 21.1.4.1. Included is US$200,000 for permitting and US$2 million for closure costs, excluding the TSF. Contingency is also included in the estimate.

Table 21.1.4.1: Owner Costs (US$000’s)

Description Initial Sustaining LoM Feasibility Study and Permitting 3,056 0 3,056 Mine Closure 0 3,000 3,000 Total $3,056 $3,000 $6,056


21.2 Operating Cost Estimates Operating costs are estimated on preliminary mine and process design criteria, engineering, as well as budgetary quotes. All costs are in 3Q 2012 US dollars. Operating costs are to a PFS-level of accuracy and are expected to be within ±25%.

LoM operating costs are shown in Table 21.2.1. Over the LoM, operating costs will be about US$65.35/t of product produced.

Table 21.2.1: LoM Operating Cost Summary Description $/t-Product LoM (US$000’s) Contract Mining 8.89 44,340 Mine G&A (Owner) 1.00 5,000 Mine Support (Owner) 0.21 1,028 Processing 59.18 295,090 G&A 1.43 7,152 Total $70.72 $352,610

Source: I-Minerals, R&S/KBR, Tetra Tech, SRK


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 152 21.2.1 Mining

Mine operating costs will average US$7.30/t-RoM (US$10.10/t-product) as summarized in Table 21.2.1.1. The majority of the cost is attributable to an average contract mining rate of US$4.33/t- mined.

Table 21.2.1.1: Mine Operating Cost Summary

Description US$/t-RoM LoM (US$000’s) Contract Mining 6.45 44,340 Mine G&A (Owner) 0.73 5,000 Mine Support (Owner) 0.13 861 Total $7.30 $50,201

Source: I-Minerals, 2014; Debco, 2014

The mine contractor cost estimate as developed by SRK is summarized in Table 21.2.1.2. Contract mining will be performed at an average cost of US$4.33/t-material moved. This equates to a value of US$7.52/t-RoM over the LoM.

Table 21.2.1.2: Mine Contractor Cost Estimate

Description $/t-mined Waste Load, Haul & Place @ Disposal Site 2.50 Product Load Haul & Place @ Plant Site 4.75 Average Contractor Cost $4.33

Source: Debco, 2014 Note: Above prices include haul road maintenance for the road to the plant site from the mined area.

Mine G&A costs, shown in Table 21.2.1.3 include leasing space from the mine contractor’s maintenance facility (US$100,000/year), dewatering the pit (US$40,000/year) and exploration (US$250,000/3-years). Mine-related supervisory, engineering, geology, and surveying personnel are included in the process operating cost estimate in Section 21.2.2.

Mine G&A costs will be US$5.0 million over the LoM or US$0.73/t-RoM over the LoM.

Table 21.2.1.3: Mine G&A (Owner)

Description LoM Cost (000’s) Mine G&A Exploration QA/QC 1,500 Debco Maintenance Facility 2,500 Dewatering 1,000 Total $5,000

Source: I-Minerals, 2014; Debco, 2014

Support equipment costs are estimated from first principles using the productivity table presented below. Table 21.2.1.4 estimates operating costs based upon equipment operating parameters for the project. Support equipment costs will be US$861,000 over the LoM, or US$0.13/t RoM.

Support equipment annual hours are based upon 60% utilization operating 260 day/year, and 1-8 hour shift per day.



Table 21.2.1.4: Mine Support Equipment (Owner)

Description Fuel

gal/hr Lube

gal/hr Fuel

US$/hr Lube

US$/hr Tires

US$/hr

O/H Parts

US$/hr

Maint. Parts

US$/hr

Wear Parts

US$/hr Total

US$/hr Flat Bed 3.39 0.10 11.86 0.76 0.25 0.36 0.67 0.00 $13.90 Pickups 1.56 0.06 5.46 0.47 0.10 0.20 0.38 0.00 $6.61 Light Plants 0.32 0.15 1.12 1.13 0.01 0.15 0.00 0.00 $2.41 Source: SRK

Fuel and lubricant costs are based upon budget estimates of US$3.50/gal and US$7.50/gal, respectively.

21.2.2 Processing Process operating costs for the Project were estimated by R&S/KBR. LoM average processing operating cost is US$59.18/t-Product as shown in Table 21.2.2.1.

Table 21.2.2.1: LoM Process Operating Cost

Description US$/t-Product LoM (US$000s) Process Labor 22.35 111,450 Reagents 5.39 26,870 Electricity 10.38 51,750 Natural Gas 14.98 74,675 Maintenance 5.88 29,300 Maintenance 0.21 1,045 Total $59.18 $295,090 Source: R&S/KBR, 2014

Process operating costs were developed using first principles and based upon the flowsheet described in Section 17.2, shown in Figures 17.2.1 and 17.2.2, and detailed in Table 21.2.2.2.

Labor will consist of five supervisory personnel and 50 operators. It should be noted that mine supervisory and engineering tasks are included in this estimate, and therefore are not in the mine G&A (owner) operating costs as presented in Section 21.2.1. Salary and hourly rates are consistent with those in the Northwest US.

Flotation reagent consumption rates were determined from process testwork. Reagent unit costs are based upon budgetary quotes provided to I-Minerals.

Operating/Maintenance cost for the tailings storage facility was developed by Tetra Tech and include the pumping and placement cost for the thickened tailings in the Boyle TSF. The cost is estimated to be US$1.0 million LoM or US$0.15/t of RoM.



Table 21.2.2.2: Annual Process Cost Detail Cost Component Unit Units/Year US$/t US$/Year Salaried Personnel

Manager of Operations No. 1

150,000 Mine Engr./Asst. Manager No. 1

112,500

Plant Metallurgist/Proc. Engr. No. 1

106,250 Manager, Safety, Health & Env. No. 1

93,750

Chemist/Laboratory Manager No. 1

65,000 Subtotal

5

$527,500

Hourly Employees Mine Surveyor No. 1

64,896

Assistant Surveyor No. 1

48,672 Plant Shift Supervisor No. 4

295,277

Office Personnel No. 2

97,344 Maintenance/Electrician No. 8

588,224

Lab Technicians No. 6

324,480 Operators & Helpers No. 42

2,510,934

Subtotal

64

$3,929,827 Total Labor

69

$4,457,327

Flotation Reagents (NaPO3)6 t/y 277 1,500 415,500

H2SO4 t/y 118 200 23,600 HF t/y 161 450 72,450 Amine t/y 72 3,000 216,000 Fuel Oil (Sp.Gr.0.9) t/y 46 930 42,780 Frother t/y 21 3,000 63,000 Lime (CaO) t/y 63 150 9,450 Flocculent t/y 58 4,000 232,000 Chlorine t/y 0.05 500 25 Subtotal

$1,074,805

Electricity $/y

2,069,917 Natural Gas $/y

2,986,767

Annual Maintenance Cost

Maintenance $/y

1,172,405 Subtotal

$6,229,088

Total

$11,761,220 Source: R&S/KBR, 2014

21.2.3 G&A G&A costs will average US$1.04/t-product, or an average of US$280,000/y over the LoM. The estimate includes general maintenance, outside services, permitting and environmental, and a provision for the operation of the laboratory.

Table 21.2.3.1: G&A Operating Cost Summary

Description $/t-Product LoM (US$000’s) G&A 0.36 2,500 Permitting & Environmental 0.20 1,376 Laboratory 0.48 3,276 Total $1.04 $7,152

Source: SRK



22 Economic Analysis The financial results of this report are based upon work performed by SRK, R&S/KBR, Tetra Tech, and I-Minerals and have been prepared on an annual basis. All costs are in Q1 2014 US dollars.

22.1 Principal Assumptions A financial model was prepared on an unleveraged, post-tax basis; the results are presented in this Section. Key criteria used in the analysis are discussed in detail throughout this report. Financial assumptions used are shown summarized in Table 22.1.1.

Table 22.1.1: Model Parameters

Model Parameter Technical Input Pre-Production Period 2 years Mine Life 25 years Product Price (Consolidated LoM Average of all products) US$254.79 Operating Days per Year 350 Discount Rate 6.0%

Source: SRK

A 2-year preproduction period is assumed to allow for permitting, detailed engineering, and due diligence/financing. The mine will have an estimated mine life of 25 years given the Mineral Reserve described in this report.

The analysis calculates a LoM weighted average market price of US$254.79/t for all products produced and sold. All products will be sold FOB at the Bovill Processing Plant.

Financial Assumptions used in the analysis are shown in Table 22.1.2.

Table 22.1.2: Financial Assumptions

Description Parameter Project Equity 100% (no gearing) Working Capital Requirement 20% of cash costs Depreciation 10-yr MACRS Discount Rate 6.0% Federal Tax 35% State Tax 6.7%

Source: SRK

22.2 Cashflow Forecasts and Annual Production Forecasts The financial analysis results, shown in Table 22.2.1, indicate an NPV 6% of US$330.8 million and an IRR of 38.2% on a pre-tax basis. After-tax basis produces an NPV 6% of US$212.7 million with an IRR of 30.5%.. Payback will be 3 years from the start of production. The following assumptions and estimates provide the basis of the SRK LoM plan and economics:

• A mine life of 25 years; • Total RoM processed is estimated to be 6,878 kt of which 4,986 kdt of product is produced; • Product yields include, 3.8% halloysite, 6.9% kaolin, 7.3% metakaolin, 16.6% K-Spar, and

37.9% quartz, over the LoM;



• An average operating cost of US$70.68/t-product; • Capital costs of US$90.8 million, consisting of initial capital costs of US$72.7 million, and

sustaining capital over the LoM of US$18.2 million; • Mine closure cost, included in the above estimates is US$5.1 million; and • The analysis does not include any allowance for salvage value.

Table 22.2.1: Financial Model Results

Description Units Value (000’s) Unit Cost (US$/t-Prod) Production

Waste Material kt 4,743 RoM Ore Processed kt 6,878 - Recovered Products kdt 4,986 - Estimate of Cash Flow Gross Revenue US$000’s $1,270,410 $254.79 Freight & Marketing US$000’s ($4,986) ($1.00) Net Revenue US$000’s $1,265,424 $253.79 Royalties US$000’s ($63,521) ($12.74) Gross Income US$000’s $1,201,904 $241.05 Operating Costs Contract Mining US$000’s $44,340 $8.89 Mine G&A US$000’s $5,000 $1.00 Mine Support US$000’s $861 $0.17 Process Opex US$000’s $295,090 $59.18 G&A US$000’s $7,152 $1.43 Operating Costs US$000’s $352,443 $70.68 Operating Margin US$000’s $849,460 $170.36 Capital Costs

Mining US$000’s $1,394 - Process Capital US$000’s $69,335 - Tailings Capital US$000’s $14,070 - Owners Costs US$000’s $6,056 - Total Capital US$000’s $90,854 - Income Tax US$000’s ($262,600) - Mining License Tax US$000’s $0 - Cash Flow US$000’s $496,006 - Present Value 6% $212,767 - IRR 30.5% -

Source: SRK

Table 22.2.2 shows the annual production tonnage, free cash flow for debt servicing, and present values at a 6% discount rate for the project on a yearly after-tax basis.



Table 22.2.2: Annual Production and Cashflow US$(000) Year Waste (kt) RoM (kt) Total Product kt Cashflow US$(000) PV @ 6% US$(000) 2015 0 0 0 (20,917) (20,917) 2016 75 0 0 (51,923) (48,984) 2017 188 217 151 20,611 18,344 2018 192 288 202 26,666 22,389 2019 219 278 204 25,895 20,511 2020 197 300 207 25,677 19,187 2021 196 301 212 25,872 18,239 2022 184 296 212 23,655 15,732 2023 188 292 212 25,013 15,693 2024 190 290 210 24,819 14,690 2025 198 282 206 24,440 13,647 2026 200 280 201 24,234 12,766 2027 200 280 199 23,235 11,547 2028 234 246 199 21,907 10,271 2029 219 261 198 21,414 9,471 2030 198 282 200 22,357 9,329 2031 192 282 201 22,248 8,758 2032 197 283 201 22,420 8,326 2033 152 287 205 22,669 7,942 2034 139 293 209 22,646 7,485 2035 136 296 211 23,022 7,178 2036 131 300 209 22,268 6,550 2037 184 296 208 22,111 6,136 2038 210 270 206 22,039 5,770 2039 219 261 204 21,942 5,419 2040 203 277 202 21,780 5,075 2041 103 140 118 12,666 2,784 2042 0 0 0 (2,760) (572) Total 4,743 6,878 4,986 496,006 212,767 Source: SRK Some minor variation of the total tons due to time exists; however, it is not material to the economic results.

22.3 Taxes, Royalties and Other Interests As described in the following sections, I-Minerals will be subject to the following taxes as they relate to the Project:

• Federal income tax; • State income tax; and • Mineral Royalty.

22.3.1 Federal Income Tax Corporate Federal income tax is determined by computing and paying the higher of a regular tax or a Tentative Minimum Tax (TMT). If the TMT exceeds the regular tax, the difference is called the Alternative Minimum Tax (AMT). Regular tax is computed by subtracting all allowable operating expenses, overhead, depreciation, amortization and depletion from current year revenues to arrive at taxable income. The tax rate is then determined from the published progressive tax schedule. An operating loss may be used to offset taxable income, thereby reducing taxes owed, in the previous three and following 15 years. The highest effective corporate income tax is 35%.



The AMT is determined in three steps. First, regular taxable income is adjusted by recalculating certain regular tax deductions, based on AMT laws, to arrive at AMT Income (AMTI). Second, AMTI is multiplied by 20% to determine TMT. Third, if TMT exceeds regular tax, the excess is the AMT amount payable in addition to the regular tax liability.

22.3.2 State Taxes The State of Idaho corporate income tax rate is 6.7%. A deduction is allowed for depletion.

22.3.3 Royalty I-Minerals is required to pay 5.0% of the gross revenues from the sale of all products, as discussed in Section 4.4.

22.4 Sensitivity Analysis Sensitivity analysis for key economic parameters on an after-tax basis is shown in Table 22.4.1. The project is most sensitive to product prices (revenues). The Project’s sensitivities to operating and capital costs are quite similar.

Table 22.4.1: Project NPV Sensitivities (US$ million)

Unit -15% -10% -5% Base 5% 10% 15% Revenues $159,313 $177,131 $194,949 $212,767 $230,585 $248,403 $266,221 Operating Costs $227,958 $222,895 $217,831 $212,767 $207,703 $202,639 $197,576 Capital Costs $220,749 $218,088 $215,428 $212,767 $210,106 $207,446 $204,785

Source: SRK

Source: SRK

Figure 22.4.1: Project NPV Sensitivities (US$ million)

$125,000

$150,000

$175,000

$200,000

$225,000

$250,000

$275,000

$300,000

-15% -10% -5% Base 5% 10% 15%

NPV

@ 6

% U

S$(0

00)

Sensitivity: NPV @ 6% US$(000)

Revenues Operating Costs Capital Costs



23 Adjacent Properties Hammond Engineering currently operates a small raw clay operation on the old A.P. Green Refractories pit north of Helmer. About 10 years ago, the owner revealed that the operation produced roughly 1,300 t/y RoM product of Maple member clays, which were sold at a price of US$20/t (fob). The clay was used by Wendt Pottery in Lewiston, Idaho to produce a buff-firing porcelain ceramic body and by Clayburn Industries as a clay binder for refractories. The owner of Wendt Pottery states that he still uses this clay, about the same amount each year. In 1997 reserves for this property, which are considered historic and were not prepared in accordance with NI 43-101, were 1.65 Mt, based upon 50 ft drill centers. This was extrapolated from an Information Memorandum offered by A.P. Green containing tonnages and property information in an effort to sell the operation.

24 Other Relevant Data and Information There are no other relevant data and information.

25 Interpretation and Conclusions This section summarizes the results of this PFS and the risks and uncertainties associated with the Project.

25.1 Results The results of this PFS evaluation indicate that I-Minerals can produce 800 t/d of clay (halloysite, kaolin, metakaolin), K-feldspar, and quartz mineral products. Products such as halloysite clay and kaolinite clay are used in a number of applications such as ceramic wares, household and personal care products, cosmetics, pharmaceuticals, paint, glass, roofing granules, and bricks. Metakaolin adds extra durability to pozzolan cement and is used in ready mix concrete and paving and bridge repair products. K-feldspar is used in ceramic products and glazes, and quartz is used in a variety of glass products depending on purity. Inquiries have come for these mineral products from potential markets in western U.S. and Canada. The materials would be sold FOB at the Bovill Processing Plant and/or shipped to customers depending on the product.

The Project is on state-leased land and as such, no federal mine permit approval is required. The Idaho Department of Land is the lead agency. I-Minerals will work with IDL to secure approval for its operating and reclamation plan. The mine permitting process requires permits from a number of other state agencies such as IDFG, IDWR, and IDEQ. Work completed to date includes PFS-level engineering design of the pit and plant facilities and the TSF. The geochemical characterization work completed to date indicates that the potential to generate acid or alkaline drainage from the ore and waste materials is extremely low to non-existent under normal environmental conditions. Some delineations and environmental studies have been initiated or completed. Some remaining tasks include:

• Completing the analysis of potential impacts to groundwater and surface water and developing plant to ensure compliance with all water quality standards;

• Preparing an air quality emissions inventory, model and permit application; and



• Identifying potential storm water discharges, preparation of a NOI, and preparation of a SWPPP including mitigation and monitoring.

Based on the updated mineral reserves and mine plan, approximately 5.0 Mt of mineral materials would be produced from small pits with an anticipated Project mine life of 25 years. The economic analysis results indicate an after-tax NPV of US$212.7 million at a 6% discount rate with an IRR of 30.5%. The economics are based on a consolidated LoM average price for all products of US$254.79/t-produced. The gross income, after a 5% royalty on gross proceeds is applied, is estimated to be US$1,202 million or US$241.05/t-produced. Direct operating costs are estimated at US$70.68/t-produced. Total operating costs are US$352.4 million. Total capital costs are estimated at US$90.8 million, consisting of initial capital costs of US$72.7 million, and ongoing sustaining capital over the life of operations of US$18.2 million. Mine closure costs are estimated to be US$5.1 million with no allowance for salvage value.

25.2 Significant Risks and Uncertainties

25.2.1 Exploration There are no significant risks or uncertainties associated with the exploration of the Project area.

25.2.2 Mineral Resource Estimate There are no significant risks or uncertainties associated with the mineral resources. The mineral resource estimate is adequate to support the current study.

25.2.3 Mineral Reserve Estimate There are no significant risks or uncertainties associated with the mineral reserves. The reserve estimate is adequate to support the current study.

25.2.4 Metallurgy and Processing Some of the risks associated with processing including lower than anticipated recoveries for specific mineral products or lower purity than anticipated, which would potentially increase the annual production costs and decrease revenues. The metallurgical test work and pilot plant studies are adequate to support the current study.

25.2.5 Tailings Disposal Some of the risks associated with the Boyle TSF design may include the following:

• Should a geosynthetic liner be required, this will affect the quantities and cost of the TSF; • Should the wetlands boundaries vary and be within the footprint of construction for the Boyle

TSF, the facility will need to be modified or moved; • Wetlands may be indirectly impacted by the TSF at the Boyle site location; • It is estimated that the average in situ dry density of the tailings is 90 pcf; should the actual

density vary significantly, the capacity of the TSF will change; and • The schedules are based on 7,812 hour work year; should this vary over time, the scheduled

costs will change.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 161 25.2.6 Environmental

A number of environmental permits are required for construction and operation of the facilities. Environmental permitting has an estimated completion time line of six months to one year. A moderate risk exists that that permitting may take longer than estimated thus delaying the start of the project by several months. If I-Minerals proceed as they have indicated and avoid wetlands, such that no CWA 404 permit is required, then the only Federal permitting is for stormwater (Section 402 NPDES) and the remaining permits (Idaho Mine Operation and Reclamation Authorization and Air Quality Permit) are with the State of Idaho. Furthermore, by not having a CWA 404 permit, there is no federal nexus triggering the NEPA process, and fewer potential delays related to environmental permitting would therefore be encountered.

25.2.7 Projected Economic Outcomes The economic analysis shows positive results which would indicate that the Project could move to the next phase of advancement.



26 Recommendations Recommendations for future work programs, the approximate timeline, and costs are provided in the sections below. The studies, data, and engineering designs derived in the Feasibility Study phase will be sufficient to prepare an Operations and Reclamation Plan, initiate the environmental permitting process, and justify a full production decision.

26.1 Exploration SRK recommends that I-Minerals conduct additional infill drilling within the areas supporting the first 10 years of mine life to increase the confidence of the reserve to a proven classification. This could be achieved with approximately 75 additional drillholes located on 100 ft centers. . Additional density testing should also be conducted to better characterize the relation between clay content and density. At least 300 samples from a range of total clay contents and different resource areas need to be analyzed to provide better confidence in the current density model. Upon completion of the drilling, analytical testing, and density measurements, a revised resource and reserve estimation should be completed. This will reduce project risk during the critical payback period.

26.2 Mining Study As part of the mine planning process, the following recommendations are made for the Feasibility Study:

• Review the key drivers of the project (e.g. Halloysite, Kaolin versus Sand); • Consider the use of mining pits that contain only Sand material (i.e. no clay); • Review waste dump locations (consider sterilization drilling); • Consider tendering the earthworks contract to ensure a competitive bid is being used for

costing purposes; and • Complete a review of geotechnical conditions for the Kelly’s Hump and Kelly South deposits;

26.3 Environmental Permitting and Supporting Studies It is premature in the PFS phase to develop a complete list of permits required until the project design is completed, including the ultimate mine plan and TSF design. The subsequent feasibility level design will confirm the extent of surface disturbance expected and potential impacts to surface water, groundwater, and the expected emissions related to fugitive dust from plant operations and tailings storage. A list of studies, inventories, plans, and applications to be completed in the future are outlined in Table 20.2.1.1. I-Minerals has secured qualified environmental consultants to advance the next stage of permitting, and the process with the IDL and the state and federal environmental agencies is well understood.

The timeline for securing the major permits is estimated to be 12 months once the application documents have been prepared. The major permits and expected timeline includes:

• Idaho Mine Operation and Reclamation Authorization – 12 months from time of submittal of the Plan of Operation, Reclamation Plan, and application;

• Air Quality Permit – 12 months from time of sufficient facility design (would occur after feasibility stage is completed).



• Section 402 NPDES – these are general permits and require submittal of a NOI and development of a SWPPPs. There is a 30-day waiting period after submittal of NOI. The SWPPPs and NOI application are expected to be completed two months prior to project construction startup.

26.4 Tailings Characterization and Design The Prefeasibility level study was completed to provide a design and scheduled cost estimate for the Bovill Kaolin Boyle TSF. The following tasks should be completed or investigated further in future studies:

• A combined stress-strain and seepage model should be completed to further understand the consolidation, seepage, evapotranspiration, infiltration rate, and the cover design of the Boyle TSF;

• A liquefaction analysis is required due to the gradation of the tailings and their non-plastic nature;

• Bench level tailings testing is required, as the tailings rheology and materials have changed since the last test;

• Additional investigations and laboratory testing of the native soils should be completed to further understand how much native material is suitable for use as the clay liner and structural fill;

• An investigation of springs and seeps in the impoundment area should be completed to assess the flow that may occur within the foundation of the TSF;

• A tradeoff study is recommended to optimize the type of liner system; and • A tradeoff study is recommended to evaluate the potential for a dry stack versus a

conventional tailings facility to reduce earthwork quantities.

26.5 Metallurgy and Processing Recommendations for future work include a more comprehensive material testing program with specific equipment throughout the clay and feldspathic process. Major process equipment will need further analysis on sizing to ensure proper capacity for the needs of the market. Major equipment testing is precursory to detailed engineering and final plant design.

26.6 Recommended Work Programs and Costs The recommended work programs for the Feasibility Study phase of the project will be performed during the next 12 to 18 months and will refine PFS level engineering studies completed in 2011-2012.


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Page 164 26.6.1 Costs

The estimated costs for the recommended work programs to be completed in the next 12 to 18 months are presented in Table 26.6.1.1. SRK has reviewed this program and believes that the costs and timeline are reasonable.

Table 26.6.1.1: Summary of Costs for Recommended Work (US$)

Study Items Project Costs Project Management 320,000 Marketing 450,000 Production Development 130,000 Environmental and Permits 325,000 Drilling and Resources 500,000 Reserves and Mine Planning 80,000 Metallurgical Testwork 100,000 Process and Infrastructure 451,000 Tailings and Waste Storage 515,000 Hydrogeologic Modeling 130,000 Technical Model and Final Report 55,000 Total Feasibility Cost $3,056,000 Source: SRK



27 References Browne, James L., 2006, The report on the Helmer-Bovill feldspar, quartz, and kaolin mineral leases,

Latah County, Idaho: unpublished report prepared for I-Minerals Vancouver, B.C., March 13, 2006.

Bush, J.H., Odenborg, L.J. and Odenborg, N.D., 1999, Bedrock map of the Deary Quadrangle, Latah County, Idaho: Idaho Geological Survey Tech. Report 99-3.

Clark, J.G., 2003a, Geological evaluation of two feldspar-clay prospects associated with intrusive rocks of the Idaho batholith in Latah County, east of Moscow, Idaho: Confidential Report prepared for Alchemy Ventures Ltd., 122 p.

Clark, J.G., 2003b, Petrographic and mineralogical characterization of feldspar-quartz ores and products from the Helmer-Bovill project and comparison with selected commercial and prospective commercial feldspar-quartz ores and products: Confidential Report prepared for Alchemy Ventures Ltd., 170 p.

Clark, J.G., 2003c, Petrographic characterization of feldspar and quartz products and residuals from the Blue Lagoons bulk tailings sample and from drill core intervals MC-3, 14-40’ and MC-22, 45-80’: Confidential Report prepared for Alchemy Ventures Ltd., 89 p.

Clark, J.G., 2004a, Report on the Helmer-Bovill feldspar, quartz, and kaolin mineral leases, Latah County, Idaho: Confidential NI 43-101 Report prepared for I-Minerals Inc., 73 p.

Clark, J.G., 2004b, Petrographic/cathodoluminescence characterization of drill core samples from the Alchemy Kaolin 2003 WBL drilling program: Confidential Report prepared for Alchemy Ventures Ltd., 175 p.

Clark, J.G., 2004c, Petrographic/cathodoluminescence characterization of granitoid samples from drillhole MC-3, Moose Creek Basin, Latah County, Idaho: Confidential Report prepared for I-Minerals Inc., 26 p.

Clark, J.G., 2004d, Petrographic/cathodoluminescence study of bulk sample feldspar and quartz products from the 2003 exploration drilling program, Helmer-Bovill project, Latah County, Idaho: Confidential Report prepared for I-Minerals Inc., 89 p.

Clark, J.G., 2005, Petrographic/Cathodoluminescence characterization of drill core samples from the Kelly's Basin 2005 diamond drilling program: Confidential Report prepared for I-Minerals Inc., 149 p.

Erikson, E.H., 1986, Northwest kaolin resources – A review of the Bovill deposits, northern Idaho: paper presented at Northwest Mining Association’s Annual Meeting, 1986, 18 p.

FLSmidth Krebs, 2008, Laboratory Report For I-Minerals – File N: 1716, Series No: 3044.

Ginn Mineral Technology, Inc., 2005, Kaolinite/Halloysite clay evaluation: Unpublished confidential report prepared for I-Minerals Inc., 20 p.

Ginn Mineral Technology, Inc., 2006, Kaolinite/Halloysite clay evaluation, Phase II: Confidential Report prepared for I-Minerals Inc., 26 p.



Ginn Mineral Technology, Inc., 2008, Hydrocyclone Evaluation: Confidential Report prepared for I-Minerals Inc., 30 p.

Ginn Mineral Technology, Inc., 2010, Hydrocyclone Evaluation: Confidential Report prepared for I-Minerals Inc., 33 p.

Hodgson, C.J., 2000, Interim report on Bovill kaolin project, Latah County, Idaho: Unpublished confidential NI 43-101 report prepared for Alchemy Ventures Ltd., 22 p.

Hosterman, J.W., and Prater, C.S., 1964, Clays, in Mineral and Water Resources of Idaho, Idaho Bureau of Mines and Geology Special Report #1, p. 51 – 57.

Hosterman, J.W., Scheid, V.E., Allen, V.T., and Sohn, I.G., 1960, Investigations of some clay deposits in Washington and Idaho: U.S. Geol. Survey Bull. 1091, 147 p.

Idaho Department of Labor. 2013. http://labor.idaho.gov/publications/lmi/pubs/latahProfile.pdf.

Kelly, H.J., Carter, G.J., and Todd, G.H., 1963, Bovill clay and sand deposit, Latah County, Idaho: U.S. Bureau of Mines R.I. 6330, 56 p.

Kirkham, V.R.D., and Johnson, M.M., 1929, The Latah formation in Idaho: Jour. Geology, v. 37, p. 483 – 504.

Montgomery, J.H., 2002, Helmer-Bovill feldspar, clay and silica mineral leases, Latah County, Idaho: Unpublished confidential NI 43-101 report prepared for Alchemy Ventures Ltd., 76 p.

Murray, H.H., Partridge, P., and Post, J.L., 1978, Alteration of a granite to kaolin – Mineralogy and geochemistry: Schriftenr. Geol. Wiss. 11, Berlin, p. 197-208.

Priebe, K.L., and Bush, J.H., 1999, Bedrock geologic map of the Stanford quadrangle, Latah County, Idaho: Idaho Geological Survey Technical Report 99-8.

Schlanz, J.W., 2011, Recovery of high grade quartz from I-Minerals WBL ore body: unpublished report prepared by Minerals Research Laboratory, North Carolina State University, September 14, 2011..

SRK Consulting, 2010, NI 43-101 Technical Report Helmer-Bovill Project, Kelly’s Basin Mine, Latah County, Idaho: technical report published on SEDAR filing system, November 5, 2010.

SRK Consulting, 2012a, NI 43-101 Preliminary Economic Assessment, Bovill Kaolin Project, Latah County, Idaho: technical report published on SEDAR filing system, January 18, 2012.

SRK Consulting, 2012b, NI 43-101 Technical Report on Resources, WBL Tailings Project, Latah County, Idaho: technical report published on SEDAR filing system, January 18, 2012.

STRATA, 2012, Geotechnical Exploration-Prefeasibility, High Wall Stability, Plant Site, Access Road study on the Bovill Kaolin Project, File No. IMINER M012040A, June 27, 2012.

Streckeisen, E.C., 1976, To each plutonic rock its proper name: Earth Science Reviews, International Magazine for Geo-scientists, Amsterdam, vol. 12, p. 1-33.

Swanson, D.A., Wright, T.L., Hooper, P.R., and Bentley, R.D., 1979, Revisions in stratigraphic nomenclature of the Columbia River Basalt group: U.S. Geol. Survey Bull. 1457-G, 59 p.


http://labor.idaho.gov/publications/lmi/pubs/latahProfile.pdf


Tetra Tech, 2012, Volume 5: Tailings Management Technical Report, Bovill Kaolin Project, Bovill, Latah Country, State of Idaho. Tetra Tech Project No. 114-311277: Unpublished report prepared for I-minerals Inc., October 5, 2012.

Tullis, E.L., 1944, Contributions to the geology of Latah County, Idaho: Bulletin of the Geol. Soc. of America, v. 55, p. 131-164.

U.S. Fish and Wildlife Service (USFWS) June 10, 2014. Idaho Fish and Wildlife Office Endangered, Threatened, Proposed, and Candidate Species With Associated Proposed and Critical Habitats in Idaho

United States Census Bureau (USCB) 2013. Latah County, Idaho, http://quickfacts.census.gov/qfd/states/16/16057.html

Yuan, J., 1994, Clay mineralogy and its influence on industrial uses of some kaolin clay deposits from south China and eastern Washington – Idaho, U.S.A.: [Ph.D. Thesis] Indiana University, 153 p.


http://quickfacts.census.gov/qfd/states/16/16057.html


28 Glossary 28.1 Mineral Resources

The mineral resources and mineral reserves have been classified according to the “CIM Standards on Mineral Resources and Reserves: Definitions and Guidelines” (November 27, 2010). Accordingly, the Resources have been classified as Measured, Indicated or Inferred, the Reserves have been classified as Proven, and Probable based on the Measured and Indicated Resources as defined below.

A Mineral Resource is a concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.

An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.

An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough to confirm both geological and grade continuity.

28.2 Mineral Reserves A Mineral Reserve is the economically mineable part of a Measured or Indicated Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate information on mining, processing, metallurgical, economic and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified. A Mineral Reserve includes diluting materials and allowances for losses that may occur when the material is mined.

A ‘Probable Mineral Reserve’ is the economically mineable part of an Indicated, and in some circumstances a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility



Study. This Study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified.

A ‘Proven Mineral Reserve’ is the economically mineable part of a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction is justified.

28.3 Definition of Terms The following general mining terms may be used in this report.

Table 28.3.1: Definition of Terms Term Definition Assay The chemical analysis of mineral samples to determine the metal content. Capital Expenditure All other expenditures not classified as operating costs. Composite Combining more than one sample result to give an average result over a larger

distance. Concentrate A metal-rich product resulting from a mineral enrichment process such as gravity

concentration or flotation, in which most of the desired mineral has been separated from the waste material in the ore.

Crushing Initial process of reducing ore particle size to render it more amenable for further processing.

Cut-off Grade (CoG) The grade of mineralized rock, which determines as to whether or not it is economic to recover its gold content by further concentration.

Dilution Waste, which is unavoidably mined with ore. Dip Angle of inclination of a geological feature/rock from the horizontal. Fault The surface of a fracture along which movement has occurred. Footwall The underlying side of an orebody or stope. Gangue Non-valuable components of the ore. Grade The measure of concentration of gold within mineralized rock. Hangingwall The overlying side of an orebody or slope. Haulage A horizontal underground excavation which is used to transport mined ore. Hydrocyclone A process whereby material is graded according to size by exploiting centrifugal

forces of particulate materials. Igneous Primary crystalline rock formed by the solidification of magma. Kriging An interpolation method of assigning values from samples to blocks that minimizes

the estimation error. Level Horizontal tunnel the primary purpose is the transportation of personnel and

materials. Lithological Geological description pertaining to different rock types. LoM Plans Life-of-Mine plans. LRP Long Range Plan. Material Properties Mine properties. Milling A general term used to describe the process in which the ore is crushed and ground

and subjected to physical or chemical treatment to extract the valuable metals to a concentrate or finished product.

Mineral/Mining Lease A lease area for which mineral rights are held. Mining Assets The Material Properties and Significant Exploration Properties. Ongoing Capital Capital estimates of a routine nature, which is necessary for sustaining operations. Ore Reserve See Mineral Reserve. Pillar Rock left behind to help support the excavations in an underground mine. RoM Run-of-Mine. Sedimentary Pertaining to rocks formed by the accumulation of sediments, formed by the erosion

of other rocks. Shaft An opening cut downwards from the surface for transporting personnel, equipment,

supplies, ore and waste.



Term Definition Sill A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by the

injection of magma into planar zones of weakness. Smelting A high temperature pyrometallurgical operation conducted in a furnace, in which the

valuable metal is collected to a molten matte or doré phase and separated from the gangue components that accumulate in a less dense molten slag phase.

Stope Underground void created by mining. Stratigraphy The study of stratified rocks in terms of time and space. Strike Direction of line formed by the intersection of strata surfaces with the horizontal

plane, always perpendicular to the dip direction. Sulfide A sulfur bearing mineral. Tailings Finely ground waste rock from which valuable minerals or metals have been

extracted. Thickening The process of concentrating solid particles in suspension. Total Expenditure All expenditures including those of an operating and capital nature. Variogram A statistical representation of the characteristics (usually grade).

28.4 Abbreviations The following abbreviations may be used in this report.

Table 28.4.1: Abbreviations Abbreviation Unit or Term ° degree (degrees) °C degrees Centigrade % percent µm micron or microns ac acre (acres) Ag silver Au gold AuEq gold equivalent grade BMP Best Management Practice CoG cut-off grade cm centimeter cm2 square centimeter cm3 cubic centimeter CWA Clean Water Act EIS Environmental Impact Statement EMP Environmental Management Plan ft foot (feet) ft2 square foot (feet) ft3 cubic foot (feet) g gram gal gallon g/L gram per liter g-mol gram-mole gpm gallons per minute g/t grams per ton ha hectares HDPE Height Density Polyethylene hp horsepower ID2 inverse-distance squared ID3 inverse-distance cubed IDAPA Idaho Administrative Procedure Act IDEQ Idaho Department of Environmental Quality IDWR Idaho Department of Water Resources IDL Idaho Department of Lands



Abbreviation Unit or Term IFC International Finance Corporation kg kilograms km kilometer km2 square kilometer koz thousand troy ounce kt thousand tons kt/d thousand tons per day kt/y thousand tons per year kV kilovolt kW kilowatt kWh kilowatt-hour kWh/t kilowatt-hour per short ton L liter L/sec liters per second L/sec/m liters per second per meter lb pound LoM Life-of-Mine m meter m2 square meter m3 cubic meter masl meters above sea level mg/L milligrams/liter mm millimeter mm2 square millimeter mm3 cubic millimeter Moz million troy ounces MSGP Multi-Sector General Permit Mt million tons MTW measured true width MW million watts m.y. million years NEPA National Environmental Policy Act NI 43-101 Canadian National Instrument 43-101 NOI Notice of Intent NPDES National Pollutant Discharge Elimination System oz troy ounce ppb parts per billion ppm parts per million QA/QC Quality Assurance/Quality Control RC rotary circulation drilling RoM Run-of-Mine RQD Rock Quality Description SEC U.S. Securities & Exchange Commission sec second SHPO State Historic Preservation Office st short ton (2,000 pounds) SWPPP Stormwater Pollution Prevention Plan t short ton (2,000 pounds) t/h tons per hour t/d tons per day t/y tons per year TSF tailings storage facility USACE U.S. Army Corps of Engineers USFWS U.S. Fish & Wildlife Service V volts W watt XRD x-ray diffraction y year


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Appendices

Appendices


SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Bovill Kaolin Project Appendices

Appendix A: Certificates of Qualified Persons


SRK Denver Suite 3000 7175 West Jefferson Avenue Lakewood, CO 80235 T: 303.985.1333 F: 303.985.9947 [email protected] www.srk.com

CERTIFICATE OF QUALIFIED PERSON

I, Bart A. Stryhas PhD, CPG # 11034, do hereby certify that:

1. I am an Associate Resource Geologist of SRK Consulting (U.S.), Inc., 7175 W. Jefferson Ave, Suite 3000, Denver, CO, USA, 80235.

2. This certificate applies to the technical report titled “NI 43-101 Updated Prefeasibility, Technical Report, Bovill Kaolin Project, Latah County, Idaho” with an Effective Date of April 20, 2014 (the “Technical Report”).

3. I graduated with a Doctorate degree in Structural Geology from Washington State University in 1988. In addition, I have obtained a Master of Science degree in Structural Geology from the University of Idaho in 1985 and a Bachelor of Arts degree in Geology from the University of Vermont in 1983. I am a current member of the American Institute of Professional Geologists. I have worked as a Geologist for a total of 26 years since my graduation from university. My relevant experience includes minerals exploration, mine geology, project development and resource estimation. I have conducted resource estimations since 1988 and have been involved in technical reports since 2004.

4. I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

5. I visited the Bovill Kaolin property for 1 day on May 12, 2010 and again on September 5, 2013. 6. I am responsible for preparation of the Geology and Resource Sections 1.1 through 1.3, 1.5, 2 through

12 except for 4.4 and 5.5, 14, 23, 25.1, 25.2.1, 25.2.2, 26.1, 27 and 28 of the Technical Report. 7. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101. 8. I have had prior involvement with the property that is the subject of the Technical Report. I was the QP of

Mineral Resources supporting an NI 43-101 compliant PEA of the Bovill Kaolin Project in 2012. I was also an authoring QP of the report titled “NI 43-101 Prefeasibility Technical Report, Bovill Kaolin Project, Latah County, Idaho”, with an Effective date, November 23, 2012.

9. I have read NI 43-101 and Form 43-101F1 and the sections of the Technical Report I am responsible for have been prepared in compliance with that instrument and form.

10. As of the aforementioned effective date, to the best of my knowledge, information and belief, the sections of the Technical Report I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 26th Day of June, 2014. “Signed” “Sealed” ________________________________ Bart A. Stryhas PhD, CPG

U.S. Offices: Anchorage 907.677.3520 Denver 303.985.1333 Elko 775.753.4151 Fort Collins 970.407.8302 Reno 775.828.6800 Tucson 520.544.3688

Mexico Offices: Querétaro 52.442.218.1030

Canadian Offices: Saskatoon 306.955.4778 Sudbury 705.682.3270 Toronto 416.601.1445 Vancouver 604.681.4196 Yellowknife 867.873.8670

Group Offices: Africa Asia Australia Europe North America South America

CERTIFICATE OF QUALIFIED PERSON

I, David S. Hallman, PE, do hereby certify that:

1. I am a Geological Engineer with Tetra Tech with a business address at 350 Indiana Street, Suite 500. Golden, Colorado, U.S.A. 80401.


3. I am a graduate of The Colorado School of Mines at Golden, Colorado (Bachelor of Science, 1983). I am a member in good standing of the Board of Professional Engineers of the State of Idaho (License No. 8350). My relevant experience includes over twenty five years of professional experience in the design and construction of tailings storage facilities in the United States, Mexico, Panama, Guatemala, El Salvador, Venezuela, Guyana, Bolivia, Peru, Chile, Argentina, Dominican Republic, Turkey, and Australia. I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).


5. I am responsible for preparation of the Tailings Sections 18.5, 21.1.3, 25.2.5 and 26.4 of the Technical Report.

6. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101.

7. I have had limited prior involvement in the Pre-Feasibility Study for the subject of the Technical Report. I was also an authoring QP of the report titled “NI- 43-101 Prefeasibility Technical Report, Bovill Kaolin Project, Latah County, Idaho,” with an Effective Date of November 23, 2012.

8. I have read NI 43-101 and Form 43-101F1 and the sections of the Technical Report I am responsible for have been prepared in compliance with that instrument and form.

9. As of the aforementioned effective date, to the best of my knowledge, information and belief, the sections of the Technical Report I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 26th Day of June, 2014. [SIGNED, SEALED] David S. Hallman, PE

SRK Denver 7175 West Jefferson Avenue. Suite 3000 Lakewood, CO 80235 T: 303.985.1333 F: 303.985.9947 [email protected] www.srk.com

CERTIFICATE OF AUTHOR

I, Valerie Obie, BS Mining, RM-SME, do hereby certify that:

1. I am Principal Consultant (Mineral Economics) of SRK Consulting (U.S.), Inc., 7175 W. Jefferson Ave, Suite 3000, Denver, CO, USA, 80235.


3. I graduated with a degree in B.S. Mining Engineering from University of Arizona in1985. In addition, I have obtained a M.A. Organizational Management from the University of Phoenix in 1995. I am a Registered Member of the Society of Mining Engineers. I have worked as a Mining Engineer for a total of 29 years since my graduation from university. My relevant experience includes development of capital and production estimation and costing, depreciation, and development of cash flow and costing models. My experience also include open pit mine operations and planning for short to long term strategic and business plan since 1987 and have been involved in technical reports since 1995.


5. I have not visited the property. 6. I am responsible for the preparation of Mining, Reserves and Economic Sections 1.6, 1.7, 1.10, 1.11,

1.12, 15, 16, 19, 21.1.1, 21.1.4, 21.2.1, 21.2.3, 22, 24, 25.2.3, 25.2.7, 26.2, and 26.6 of the Technical Report.

7. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101. 8. I have had prior involvement with the property that is the subject of the Technical Report. The nature of

my prior involvement was in preparation of the report entitled, “NI 43-101 Prefeasibility Technical Report, Bovill Kaolin Project, Latah County, Idaho” with an Effective Date of November 23, 2012.

9. I have read NI 43-101 and Form 43-101-F1 and the sections of the Technical Report I am responsible for have been prepared in compliance with that instrument and form.

10. As of the aforementioned Effective Date, to the best of my knowledge, information and belief, the sections of the Technical Report I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 26th Day of June, 2014. “Signed” “Sealed” ________________________________

Valerie Obie, BS Mining, RM-SME

U.S. Offices: Anchorage 907.677.3520 Denver 303.985.1333 Elko 775.753.4151 Fort Collins 970.407.8302 Reno 775.828.6800 Tucson 520.544.3688

Mexico Offices: Querétaro 52.442.218.1030

Canadian Offices: Saskatoon 306.955.4778 Sudbury 705.682.3270 Toronto 416.601.1445 Vancouver 604.681.4196 Yellowknife 867.873.8670

Group Offices: Africa Asia Australia Europe North America South America

ni 43-101 updated prefeasibility technical report bovill

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