NI 43-101 Technical Report Prefeasibility Study Toroparu Gold Project Upper Puruni River Area,...

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NI 43-101 Technical Report

Prefeasibility StudyToroparu Gold ProjectUpper Puruni River Area, GuyanaEffective Date: May 8, 2013Report Date: May 24, 2013

Report Prepared for

Sandspring Resources Ltd.8000 South Chester Street, Suite 375

Centennial CO 80112

Report Prepared by

SRK Consulting (U.S.), Inc.

7175 West Jefferson Avenue, Suite 3000

Lakewood, CO 80235

SRK Project Number: 349800.020

Signed by QP(s):

Alex Fisher, B.Sc. Geological Engineering, P.E.

Allan Moran, B.Sc., Geol. Eng., RG, CPG

D. Erik Spiller, MMSA

Daniel Lloyd Evans, CFM, PE

Daniel Y. Yang, MEng., PEng.

Dawn H. Garcia, PG, CPG

Fernando Rodrigues, BS Mining, MBA, MAusIMM, MMSAQP

Frank Daviess, MAusIMM, R.M. SME

José Enrique Sánchez Marrou, M.Sc, P.EngKeith Mountjoy, MASc, PGeo

Peter Clarke, BSc Mining, MBA, PE

Thomas A Chapel, CPG, P.E.

Reviewed by:

Bret Swanson, BEng Mining, MAusIMM, MMSAQP

Grant Malensek, P.Eng./P.Geo



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

PC/MLM Toroparu_NI43-101_TechnicalReport_349800.020_044_MLM.docx May 24, 2013

Table of Contents

1 Summary ....................................................................................................................... 1

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

1.2 Geology and Mineralization ................................................................................................................ 2

1.3 Mineral Processing and Metallurgical Testing .................................................................................... 4

1.4 Mineral Resource Estimate ................................................................................................................. 5

1.5 Mineral Reserve Estimate ................................................................................................................... 6

1.6 Mining .................................................................................................................................................. 7

1.7 Processing Recovery Methods ........................................................................................................... 9

1.8 Tailings Management Area ............................................................................................................... 10

1.9 Project On-Site Infrastructure ........................................................................................................... 10

1.10 Project Off-site Infrastructure ............................................................................................................ 11

1.11 Project Implementation ..................................................................................................................... 11

1.12 Environmental Studies and Permitting .............................................................................................. 11

1.13 Capital and Operating Costs ............................................................................................................. 14

1.14 Economic Analysis ............................................................................................................................ 16

1.15 Conclusions and Recommendations ................................................................................................ 18

2 Introduction ................................................................................................................ 29

2.1 Terms of Reference and Purpose of the Report ............................................................................... 29

2.2 Qualifications of Consultants ............................................................................................................ 29

2.3

Sources of Information ...................................................................................................................... 31

2.4 Units of Measure ............................................................................................................................... 31

2.5 Glossary and Abbreviated Terms ..................................................................................................... 31

2.5.1 Mineral Resources ................................................................................................................ 31

2.5.2 Mineral Reserves .................................................................................................................. 32

2.5.3 Definition of Terms ................................................................................................................ 32

2.5.4 Abbreviations ......................................................................................................................... 34

3 Reliance on Other Experts ........................................................................................ 38

4 Property Descript ion and Locat ion .......................................................................... 40

4.1 Property Description and Tenure ...................................................................................................... 40

4.2 Location ............................................................................................................................................. 46

5 Accessibi li ty, Climate, Local Resources, Infrast ructure and Physiography ........ 49

5.1 Topography, Elevation and Vegetation ............................................................................................. 49

5.2 Accessibility and Transportation to the Property .............................................................................. 49

5.3 Climate and Length of Operating Season ......................................................................................... 49

5.4 Infrastructure Availability and Sources.............................................................................................. 50



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5.4.1 Proximity to Population Center .............................................................................................. 50

5.4.2 Power .................................................................................................................................... 50

5.4.3 Water ..................................................................................................................................... 50

5.4.4 Mining Personnel ................................................................................................................... 50

5.4.5 Potential Tailings Storage Areas ........................................................................................... 50

5.4.6 Potential Waste Disposal Areas ............................................................................................ 50

5.4.7 Potential Processing Plant Sites ........................................................................................... 50

5.5 Physiography .................................................................................................................................... 50

6 History ......................................................................................................................... 51

6.1 Alluvial and Saprolite Exploration and Mining - 1997 to 2006 .......................................................... 51

6.2 Bedrock Exploration - 2006 to 2009.................................................................................................. 52

6.3 Sandspring Resources Ltd; Exploration from 2010 to Present ......................................................... 54

6.3.1 2010 Programs ...................................................................................................................... 54

6.3.2 2011 Programs ...................................................................................................................... 56

6.3.3 2012 Programs ...................................................................................................................... 58

6.4 Historic Production ............................................................................................................................ 60

6.5 Previous Metallurgical Testing .......................................................................................................... 60

7 Geological Setting and Mineralization ..................................................................... 62

7.1 Regional Geology – Guiana Shield ................................................................................................... 62

7.2 Regional Geology – Western Guyana .............................................................................................. 63

7.3 Property Geology .............................................................................................................................. 64

7.3.1 Weathering ............................................................................................................................ 65

7.3.2 Lithology ................................................................................................................................ 65

7.3.3 Structure ................................................................................................................................ 66

7.3.4 Alteration ............................................................................................................................... 66

7.4 Mineralization .................................................................................................................................... 67

8 Deposit Type .............................................................................................................. 74

8.1 Geological Model .............................................................................................................................. 76

9 Exploration ................................................................................................................. 77

9.1 Exploration – 2011 and 2012 ............................................................................................................ 77

9.2 Relevant Exploration Work – Post-PEA Drilling ................................................................................ 79

9.2.1 Surveys and Investigations ................................................................................................... 80

9.2.2 Procedures and Parameters ................................................................................................. 80

9.2.3 Sampling Methods and Sample Quality ................................................................................ 80

9.2.4 Significant Results and Interpretation ................................................................................... 81

10 Dri ll ing ......................................................................................................................... 92

10.1 Collar Surveys ................................................................................................................................... 92



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10.2 Downhole Surveys ............................................................................................................................ 93

10.3 Drill Core Logging ............................................................................................................................. 93

10.4 Interpretation ..................................................................................................................................... 93

10.5 Results .............................................................................................................................................. 94

11 Sample Preparation, Analysis and Security ............................................................ 98

11.1 Sampling Methods ............................................................................................................................ 98

11.2 Security Measures ............................................................................................................................ 98

11.3 Sample Preparation .......................................................................................................................... 98

11.4 Sample Analysis ................................................................................................................................ 98

11.5 QA/QC Procedures ........................................................................................................................... 99

11.5.1 QA/QC Results ...................................................................................................................... 99

11.5.2 QA/QC Actions ...................................................................................................................... 99

11.6 Opinion on Adequacy ...................................................................................................................... 100

12 Data Verification ....................................................................................................... 105

12.1 Procedures ...................................................................................................................................... 105

12.2 Limitations ....................................................................................................................................... 106

12.3 Opinion on Data Adequacy ............................................................................................................. 106

13 Mineral Processing and Metallurgical Testing ...................................................... 107

13.1 Summary ......................................................................................................................................... 107

13.2 Metallurgical Testing ....................................................................................................................... 107

13.3 Mineralogy ....................................................................................................................................... 108

13.3.1 Sulfide and Gangue Minerals .............................................................................................. 108

13.3.2 Gold Deportment ................................................................................................................. 108

13.4 Comminution Tests (ACO/LCO) ..................................................................................................... 109

13.4.1 Grindability .......................................................................................................................... 109

13.4.2 Abrasion Index .................................................................................................................... 109

13.4.3 HPGR Testing ..................................................................................................................... 109

13.5 Gravity Separation (ACO/LCO) ....................................................................................................... 110

13.5.1 Gravity Separation (ACO) ................................................................................................... 111

13.5.2 Gravity Separation (LCO) .................................................................................................... 111

13.6 Flotation Testwork (ACO/LCO) ....................................................................................................... 111

13.6.1 Phase 1 ............................................................................................................................... 111

13.6.2 Phase 2 ............................................................................................................................... 111

13.6.3 Phase 2 Extension .............................................................................................................. 112

13.6.4 Gold Ore with Average Copper (ACO) ................................................................................ 113

13.6.5 Gold Ore with Low Copper (LCO) ....................................................................................... 114

13.7 Cyanide Leaching (ACO/LCO) ........................................................................................................ 115



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13.7.1 Phase 1 ............................................................................................................................... 115

13.7.2 Phase 2 ............................................................................................................................... 116

13.7.3 Phase 2 Extension .............................................................................................................. 116

13.7.4 Locked Cycle Test Cleaner Tailing Cyanidation (ACO) ...................................................... 117

13.7.5 Gravity Concentrate Intensive Cyanidation (ACO) ............................................................. 117

13.7.6 Gravity Concentrate Intensive Cyanidation (LCO) .............................................................. 118

13.7.7 Gravity Tailing Bulk Cyanidation ......................................................................................... 118

13.7.8 Rougher Concentrate Cyanidation (LCO) ........................................................................... 118

13.7.9 Comparison of Cyanide Leaching of Gravity Tailing and Flotation Products for ACO andLCO ..................................................................................................................................... 119

13.8 Saprolite Test Work Program 2012-2013 ....................................................................................... 122

13.8.1 Gravity Separation Testwork ............................................................................................... 122

13.8.2 Flotation Testwork ............................................................................................................... 122

13.8.3 Cyanidation Testwork .......................................................................................................... 123

14 Mineral Resource Estimate ..................................................................................... 127

14.1 Drillhole Database ........................................................................................................................... 127

14.2 Targeted In-fill Drilling ..................................................................................................................... 127

14.3 Geologic Model ............................................................................................................................... 129

14.3.1 External Domain Envelope .................................................................................................. 130

14.3.2 Anisotropy Model ................................................................................................................. 131

14.3.3 Internal Domain Envelopes ................................................................................................. 131

14.4 Domain Analysis & Grade Capping ................................................................................................ 131

14.5 Compositing .................................................................................................................................... 133

14.6 Bulk Density .................................................................................................................................... 133

14.7 Block Model ..................................................................................................................................... 135

14.8 Search Orientation/ Anisotropy Model ............................................................................................ 135

14.9 Grade Assignment .......................................................................................................................... 136

14.10 Mineral Resource Classification/Confidence Assignment .............................................................. 137

14.11 Resource Statement ....................................................................................................................... 138

14.12 Model Validation & Mineral Resource Sensitivity ........................................................................... 140

14.12.1

Visual Comparison .......................................................................................................... 140

14.12.2 Comparative Statistics .................................................................................................... 141

14.12.3 Swath Analysis ................................................................................................................ 142

15 Mineral Reserve Estimate ........................................................................................ 167

15.1 Conversion Assumptions, Parameters and Methods ...................................................................... 167

15.2 Reserve Estimate ............................................................................................................................ 168

16 Mining Methods ........................................................................................................ 171



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16.1 Proposed Mining Method ................................................................................................................ 172

16.2 Geotechnical Mine Design Parameters .......................................................................................... 172

16.3 Pit Optimization ............................................................................................................................... 173

16.3.1 Mineral Resource Models.................................................................................................... 173

16.3.2 Topographic Data ................................................................................................................ 173

16.3.3 Optimization Constraints ..................................................................................................... 173

16.3.4 Optimization Parameters ..................................................................................................... 174

16.3.5 Optimization Process .......................................................................................................... 175

16.3.6 Optimization Results ........................................................................................................... 176

16.3.7 South-East Deposit ............................................................................................................. 177

16.4 Design Criteria ................................................................................................................................ 177

16.5 Pit Phase and Ultimate Pit Designs ................................................................................................ 179

16.6 Mine Production Schedule .............................................................................................................. 179

16.6.1 Mine Production .................................................................................................................. 179

16.7 Waste and Low Grade Stockpile Design ........................................................................................ 182

16.8 Mining Equipment Requirements .................................................................................................... 186

16.8.1 Summary ............................................................................................................................. 186

16.8.2 General Parameters and Fleet Selection ............................................................................ 186

16.8.3 Drilling .................................................................................................................................. 188

16.8.4 Blasting ................................................................................................................................ 189

16.8.5 Loading ................................................................................................................................ 190

16.8.6 Hauling ................................................................................................................................ 192

16.8.7 Auxiliary Equipment ............................................................................................................. 194

16.9 Mine Dewatering ............................................................................................................................. 195

16.9.1 Water Data Sources ............................................................................................................ 195

16.9.2 Water from Precipitation ...................................................................................................... 195

16.9.3 Groundwater Inflow ............................................................................................................. 195

16.9.4 Dewatering System ............................................................................................................. 195

17 Recovery Methods ................................................................................................... 208

17.1 Summary ......................................................................................................................................... 208

17.2 Overview ......................................................................................................................................... 208

17.3 Design Basis ................................................................................................................................... 208

17.3.1 Preproduction-Phase ........................................................................................................... 208

17.3.2 Phase 1: 15,000 t/d ACO and Saprolite .............................................................................. 209

17.3.3 Phase 2: 15,000 t/d LCO with 7,500 t/d ACO ..................................................................... 209

17.4 Mass and Water Balance ................................................................................................................ 210

17.5 Process Design Criteria .................................................................................................................. 210



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17.5.1 Primary/Secondary Crushing and Stockpile ........................................................................ 210

17.5.2 Grinding ............................................................................................................................... 211

17.5.3 Gravity Circuit and Intense Cyanide Leaching .................................................................... 212

17.5.4 Rougher Flotation Circuit..................................................................................................... 212

17.5.5 Regrind Circuit ..................................................................................................................... 212

17.5.6 Cleaner Flotation Circuit ...................................................................................................... 212

17.5.7 Concentrate Dewatering Circuit .......................................................................................... 213

17.5.8 CIP Circuit ........................................................................................................................... 213

17.5.9 Desorption ........................................................................................................................... 214

17.5.10 Electrowinning and Gold Room ...................................................................................... 215

17.5.11 Carbon Regeneration ...................................................................................................... 215

17.5.12 CIP Tails Detoxification and Tails Dewatering ................................................................ 215

17.5.13 Reagents ......................................................................................................................... 216

18 Project Infrastructure............................................................................................... 218

18.1 On-Site Infrastructure ...................................................................................................................... 218

18.1.1 Geotechnical ....................................................................................................................... 218

18.1.2 Site Water Management ...................................................................................................... 219

18.1.3 Service Roads and Bridges ................................................................................................. 220

18.1.4 Mine Operations Support Facilities ..................................................................................... 220

18.1.5 Process Support Facilities ................................................................................................... 221

18.1.6 Man Camp ........................................................................................................................... 222

18.1.7 Additional Support Facilities ................................................................................................ 222

18.1.8 Power Supply and Distribution ............................................................................................ 223

18.1.9 Water Supply ....................................................................................................................... 223

18.1.10 Waste Water Treatment and Solid Waste Disposal ........................................................ 224

18.2 Tailings Management Area ............................................................................................................. 224

18.3 Off-Site Infrastructure and Logistic Requirements .......................................................................... 225

18.3.1 Off-Site Infrastructure .......................................................................................................... 225

19 Market Studies and Contracts ................................................................................ 228

19.1 Summary of Information .................................................................................................................. 228

19.2 Market Studies ................................................................................................................................ 228

19.2.1 Gold in Dore’ ....................................................................................................................... 228

19.2.2 Gold & Copper in Concentrate ............................................................................................ 228

19.3 Commodity Price Projections .......................................................................................................... 228

19.4 Contracts and Status....................................................................................................................... 229

19.4.1 Metal Treatment, Refining, and Transportation .................................................................. 229

19.4.2 Supplier & Service Contracts .............................................................................................. 229



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19.5 Indicative Terms .............................................................................................................................. 229

19.5.1 Doré Net Smelter Return ..................................................................................................... 229

19.5.2 Copper Concentrate Net Smelter Return ............................................................................ 229

19.5.3 Diesel and Fuel Oil Prices ................................................................................................... 230

19.6 Royalties & Taxes ........................................................................................................................... 230

20 Environmental Studies, Permit ting and Social or Community Impact ................ 231

20.1 Environmental Study Results .......................................................................................................... 232

20.1.1 Results of Baseline Studies................................................................................................. 233

20.1.2 Results of Geochemical Studies of Tailings, Waste Rock and Low Grade Ore ................. 237

20.2 Environmental Issues ...................................................................................................................... 242

20.3 Operating and Post Closure Requirements and Plans ................................................................... 242

20.4 Post-performance or Reclamations Bonds ..................................................................................... 243

20.5 Social and Community .................................................................................................................... 243

20.6 Mine Closure ................................................................................................................................... 245

20.7 Reclamation Measures during Operations and Project Closure ..................................................... 246

20.7.1 Tailings Management Area ................................................................................................. 246

20.7.2 Open Pits ............................................................................................................................. 247

20.7.3 Waste Rock Storage Areas ................................................................................................. 247

20.7.4 Plant Site and Facilities ....................................................................................................... 247

20.8 Closure Monitoring .......................................................................................................................... 247

20.9 Reclamation and Closure Cost Estimate ........................................................................................ 248

21 Capital and Operating Costs [All ] ........................................................................... 251

21.1 Summary ......................................................................................................................................... 251

21.2 Capital Cost Estimate...................................................................................................................... 252

21.2.1 Mining Capital Cost ............................................................................................................. 252

21.2.2 Process and On-Site Infrastructure Capital Cost ................................................................ 254

21.2.3 Off-Site Infrastructure Capital Cost ..................................................................................... 256

21.2.4 Owner’s Cost ....................................................................................................................... 258

21.2.5 Sustaining Capital Costs ..................................................................................................... 259

21.3 Operating Cost Estimates ............................................................................................................... 263

21.3.1 Summary ............................................................................................................................. 263

21.3.2 Mining Operating Costs ....................................................................................................... 264

21.3.3 Process Operating Costs .................................................................................................... 269

21.3.4 Off-Site Infrastructure Operating Cost ................................................................................. 276

21.3.5 General & Administrative Costs .......................................................................................... 276

22 Economic Analysis .................................................................................................. 278

22.1 Method of Evaluation ...................................................................................................................... 278



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22.2 Input Parameters............................................................................................................................. 279

22.3 Cashflow Forecasts and Annual Production Forecasts .................................................................. 283

22.4 Sensitivity Analysis.......................................................................................................................... 285

23 Adjacent Proper ties ................................................................................................. 290

24 Other Relevant Data and Information ..................................................................... 291

25 Interpretation and Conclusions .............................................................................. 292

25.1 Geology and Resources ................................................................................................................. 292

25.2 Mining and Reserves ...................................................................................................................... 292

25.3 Metallurgy, Processing and Recoveries ......................................................................................... 293

25.4 Infrastructure ................................................................................................................................... 294

25.5 Project Implementation ................................................................................................................... 294

25.6 Environmental Studies and Permitting ............................................................................................ 295

25.7

Economic Analysis .......................................................................................................................... 296

25.8 Risks and Uncertainties .................................................................................................................. 296

26 Recommendations ................................................................................................... 298

26.1 Recommended Work Programs and Costs .................................................................................... 298

26.1.1 Phase I ................................................................................................................................ 298

26.1.2 Phase II ............................................................................................................................... 303

26.1.3 Summary of Recommended Work Program Costs ............................................................. 305

27 References ................................................................................................................ 306

List of TablesTable 1.4.1 Resource Statement @ 0.30 g/t Au cut-off as of March 31, 2013 .................................................. 6

Table 1.5.1: March 31, 2013 Mineral Reserve Estimate ................................................................................... 7

Table 1.6.1: Planned Mine Production Schedule .............................................................................................. 9

Table 1.13.1: Summary of Capital Costs by Area ........................................................................................... 15

Table 1.13.2: Operating Cost Life of Mine, US$000s ...................................................................................... 16

Table 1.14.1: Project Evaluation Economic Results ......................................................................................... 18

Table 2.5.3.1: Definition of Terms ................................................................................................................... 32

Table 2.5.4.1: Abbreviations ............................................................................................................................ 34

Table 4.1.1: Land Tenure – Medium Scale Prospecting Permits .................................................................... 41

Table 4.1.2: Land Tenure – Mining Permits .................................................................................................... 44

Table 4.1.3: Land Tenure – Prospecting Licenses .......................................................................................... 44

Table 6.2.1: Toroparu 2008 Mineral Resources ............................................................................................... 53

Table 6.3.1.1: Toroparu 2010 Mineral Resources ........................................................................................... 55

Table 6.3.1.2: Toroparu 2010 Updated Mineral Resources ............................................................................ 55



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Table 6.3.2.1: Toroparu 2011 Mineral Resources in a PEA ............................................................................ 58

Table 6.3.3.1: PEA Mineral Resources, Effective date of January 2012 ......................................................... 60

Table 8.1: Primary Geological and Mineralization Features of Several Gold and Gold –Copper Deposits ofthe Guiana Craton ............................................................................................................................... 75

Table 9.1.1: Saprolite Sampling in 2011 and 2012 .......................................................................................... 77

Table 9.1.2: 2012 RC Reconnaissance Drilling Program ................................................................................. 78

Table 10.1: Summary of the Core Drill Programs on the Toroparu Resource Zone ....................................... 92

Table 10.5.1: Example of Drillhole assays – Hole TPD-022 ............................................................................ 94

Table 12.1.1 SRK Spot Sample Verification Assays ...................................................................................... 106

Table 13.2.1: Test Work Programs and Reports ........................................................................................... 108

Table 13.4.1.1: Summary of JK Tech/SMC Data (2011) ............................................................................... 109

Table 13.4.1.2: Grindability Data ................................................................................................................... 109

Table 13.4.3.1: Summary of HPGR Test Findings ........................................................................................ 110

Table 13.5.1: Gravity Separation Results Summary for Phase 1, Phase 2, Phase 2 Extension .................. 110

Table 13.6.4.1: ACO Combined Gravity and Flotation and Recovery of Au and Ag ..................................... 113

Table 13.6.4.2: Locked Cycle Test Results ................................................................................................... 114

Table 13.6.4.3: ACO Combined Results from Gravity Separation and LCT Tests ....................................... 114

Table 13.6.5.1: Combined Gravity and Flotation gold Recovery for the LCO Composites ........................... 115

Table 13.7.4.1: ACO Gravity, Cleaner Flotation, and Cleaner Tail Leach Summary .................................... 117

Table 13.7.7.1: LCO Combined Results from Gravity Separation and Gravity Tailing Cyanidation Tests ... 118

Table 13.7.8.1: LCO Combined Results from Gravity Separation, Rougher Flotation and RougherConcentrate Leaching........................................................................................................................ 119

Table 13.7.9.1: Overall Gold Recovery for ACO Composite, Gravity and Gravity Tailing Cyanide Leaching119

Table 13.7.9.2: Overall Gold Recovery for ACO Composite, Gravity and Gravity Tailing Rougher Flotation120

Table 13.7.9.3: Overall Gold Recovery for ACO Composite, Gravity, Gravity Tailing Cleaner Flotation andCleaner Tail cyanide Leaching .......................................................................................................... 120

Table 13.7.9.4: Overall Gold Recovery for LCO Composites, Gravity and Gravity Tailing Cyanide Leaching120

Table 13.7.9.5: Overall Gold Recovery for LCO Composites, Gravity and Gravity Tailing Rougher Flotationand Rougher Concentrate Cyanide Leaching ................................................................................... 120

Table 13.7.9.6: Copper Extraction and Cyanide Consumption for Gravity Tails Leaching of ACO .............. 121

Table 13.7.9.7: Copper Extraction and Cyanide Consumption for Cleaner Tails Leaching of ACO ............. 121

Table 13.7.9.8: Copper Extraction and Cyanide Consumption for Gravity Tails Leaching of LCO ............... 121

Table 13.7.9.9: Copper Extraction and Cyanide Consumption for Rougher Concentrate Leaching of LCO 122

Table 14.2.1: Final Ranking of Proposed In-fill Drillholes Based on Expected Oz Au and Pit Shell ............. 129

Table 14.4.1 Assay Summary Statistics by Domain ....................................................................................... 132

Table 14.4.2: Assay Capping Thresholds ...................................................................................................... 132

Table 14.4.3 Capped Assay Summary Statistics by Domain ......................................................................... 133

Table 14.5.1: Composite Summary Statistics by Domain ............................................................................. 133

Table 14.6.1: Bulk Density by Major Rock types ............................................................................................ 134



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Table 14.6.2: Bulk Density Data by Rock Code ............................................................................................ 134

Table 14.6.3: Summary of Bulk Density Data ............................................................................................... 135

Table 14.7.1: Toroparu Model Limits .............................................................................................................. 135

Table 14.9.1: Kriging Parameters .................................................................................................................. 136

Table 14.9.2 Search Neighborhood Strategy ................................................................................................. 137

Table 14.9.3 Confidence Classification Scheme ............................................................................................ 137

Table 14.11.1: Resource Reporting Cut-offs ................................................................................................. 138

Table 14.11.2: Resource Statement @ 0.30 g/t Au cut-off as of March 31, 2013 ........................................ 139

Table 14.11.3: Mineral Resource Estimate M&I Sensitivity Analysis – All Zones ......................................... 140

Table 14.11.4: Mineral Resource Estimate Inferred Sensitivity Analysis – All Zones ................................... 140

Table 14.12.2.1: Fresh Rock Composite/Model Statistics ............................................................................. 141

Table 14.12.2.2: Saprolite Composite/Model Statistics ................................................................................. 141

Table 14.12.2.3: Fresh Rock Inventory – Alternative Estimators Au ............................................................. 142

Table 14.12.2.4: Fresh Rock Inventory – Alternative Estimators Cu ............................................................ 142

Table 15.1.1: In-pit Cut-off Grade Calculation Results .................................................................................. 168

Table 15.2.1: Mineral Reserve Estimate as of March 31, 2013 .................................................................... 169

Table 16.2.1: Recommended Pit Slope Configurations ................................................................................ 173

Table 16.3.1.1: Block Model Block Sizes ...................................................................................................... 173

Table 16.3.4.1: Processing Parameters ......................................................................................................... 175

Table 16.3.4.2: Optimization Parameters (Base Case) ................................................................................. 175

Table 16.3.6.1: Whittle™ Results for Toroparu .............................................................................................. 176

Table 16.3.7.1: Whittle™ Results for South East Deposit .............................................................................. 177

Table 16.4.1 shows the final pit design parameters. ...................................................................................... 177

Table 16.4.1: Final Design Parameters ......................................................................................................... 177

Table 16.4.2: Toroparu Pit Final Geotech Pit Design Parameters Used in Updated PEA ............................ 178

Table 16.4.3: Toroparu and South-East Pits Geotech Pit Design Parameters .............................................. 178

Table 16.6.1.1: Planned Mine Production Schedule ..................................................................................... 180

Table 16.6.1.2: Production Mill Schedule (Mill Feed) .................................................................................... 181

Table 16.7.1: Toroparu Waste Dump Parameters ........................................................................................ 183

Table 16.7.2: Ore Stockpile Inventories ........................................................................................................ 185

Table 16.8.6.1: Toroparu Haulage Truck Speeds (km/hr) ............................................................................. 193

Table 17.5.1.1: Primary Jaw Crusher ............................................................................................................ 210

Table 17.5.1.2: Secondary Crusher ............................................................................................................... 211

Table 17.5.2.1: HPGR Phase 1 ..................................................................................................................... 211

Table 17.5.2.2 Ball Mill, Phase 1 .................................................................................................................... 211

Table 17.5.2.3: HPGR Phase 2 ..................................................................................................................... 212

Table 17.5.2.4 Ball Mill, Phase 2 .................................................................................................................... 212



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Table 17.5.4.1: Rougher Flotation Circuit ...................................................................................................... 212

Table 17.5.6.1: 1st Cleaner Flotation Circuit ................................................................................................. 213

Table 17.5.6.2: 1st Cleaner Scavenger Flotation Circuit ............................................................................... 213

Table 17.5.6.3: 2nd Cleaner Flotation Circuit ................................................................................................ 213

Table 17.5.8.1: CIP - Phase 1 ....................................................................................................................... 214

Table 17.5.8.2: CIP - Phase 2 ....................................................................................................................... 214

Table 17.5.9.1: Desorption Operation ........................................................................................................... 214

Table 17.5.9.2: Acid Wash ............................................................................................................................ 214

Table 17.5.9.3: Elution ................................................................................................................................... 214

Table 17.5.10.1: Electrowinning .................................................................................................................... 215

Table 17.5.10.2: Smelting .............................................................................................................................. 215

Table 17.5.11.1: Kiln Carbon Dewatering Screen ......................................................................................... 215

Table 17.5.11.2: Barren Carbon Dewatering Screen .................................................................................... 215

Table 20.1: Environmental Permits ............................................................................................................... 232

Table 20.1.1.1: Monitor Well Locations and Groundwater Elevation Data .................................................... 235

Table 20.1.1.2: International Status of Species ............................................................................................. 237

Table 20.5.1: Summary of Socio-cultural Impacts and Mitigation Strategies ................................................ 244

Table 21.1.1: Summary of Capital Costs by Area ......................................................................................... 251

Table 21.2.1.1: Initial Mining Equipment Capital Cost Estimate (US$000s) ................................................. 253

Table 21.2.2.1: Initial Capital Cost Estimate – Process and On-site Summary ............................................ 255

Table 21.2.3.1: Off-site Infrastructure Capital Cost Estimate (US$ millions) ................................................ 258

Table 21.2.4.1: Major Components of Owner’s Costs ($US million) ............................................................. 258

Table 21.2.5.1: LoM Mining Equipment Sustaining Capital Cost Estimate (US$000s) ................................. 260

Table 21.2.5.2: TMA Facilities Sustaining Capital Cost Estimate ................................................................. 262

Table 21.2.5.3: Onsite Infrastructure Sustaining Capital Cost Estimate ....................................................... 262

Table 21.3.1.1: Operating Cost Life –of Mine, US$ x 1,000 .......................................................................... 263

Table 21.3.1.2: Annual Operating Cost, US$ x 1,000 .................................................................................... 263

Table 21.3.2.2: Mine Hourly Labour Requirements ....................................................................................... 269

Table 21.3.3.1: Plant Operating Costs (LoM) ................................................................................................ 269

Table 21.3.3.2: Plant Power Requirements ................................................................................................... 270

Table 21.3.3.3: Reagent Consumption Rates and Cost – Preproduction ..................................................... 271

Table 21.3.3.4: Reagent Consumption Rates and Cost – Phase 1 .............................................................. 272

Table 21.3.3.5: Reagent Consumption Rates and Cost – Phase 2 .............................................................. 273

Table 21.3.3.6: Consumables Cost Estimate – Pre-production .................................................................... 274

Table 21.3.3.7: Consumables Cost Estimate – Phase 1 ............................................................................... 275

Table 21.3.3.8: Consumables Cost Estimate – Phase 2 ............................................................................... 275

Table 22.2.1: Key Criteria, Principal Assumptions and Input Parameters Used in the Base Case .............. 280



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project Page xii


Table 22.2.2: Project Stages ......................................................................................................................... 281

Table 22.2.3 Production Parameters .............................................................................................................. 282

Table 22.3.1: Project Evaluation Economic Results ....................................................................................... 284

Table 22.3.2: Summary of LoM Production and Cashflow ............................................................................. 285

Table 22.3.3 Project LoM Annual Production and Revenues......................................................................... 285

Table 22.4.1: Sensitivity to Capital Costs ....................................................................................................... 286

Table 22.4.2: Sensitivity to Operating Costs .................................................................................................. 286

Table 22.4.3: Sensitivity to Metal Prices ......................................................................................................... 286

Table 26.1.1.1: Condemnation Drilling Program ........................................................................................... 299

Table 26.1.3.1: Cost Summary for Recommended Work for FS Completion ............................................... 305

List of FiguresFigure 1.1.1: Upper Puruni Property and Toroparu Deposit Location Map ..................................................... 23

Figure 1.6.1: Toroparu Mine Site ..................................................................................................................... 24

Figure 1.6.2: Mine Plan Progress Maps - Year 14 End, Ultimate Pit .............................................................. 25

Figure 1.7.1: Overall Simplified Process Flow Diagram .................................................................................. 26

Figure 1.9.1: Toroparu Mine Overall Site Including Tailings Management Area ............................................. 27

Figure 1.14.1: NPV Sensitivity ......................................................................................................................... 28

Figure 4.2.1: Upper Puruni Property Claim Map ............................................................................................. 47

Figure 4.2.2: Detail of Toroparu Deposit ......................................................................................................... 48

Figure 7.1.1: Geological Sketch of the Guyana ............................................................................................... 69

Figure 7.2.1: Upper Puruni District: Regional Geological Sketch ................................................................... 70

Figure 7.3.2.1: Geological Sketch Map of the Toroparu Deposit Area ............................................................ 71

Figure 7.3.2.2: Geological Sketch of Drill Section 2+50 W Center Part of the Main East Zone (Toroparudeposit) ................................................................................................................................................ 72

Figure 7.4.1: Drillhole Plan Map with Gold Grade Distribution in Resource Areas .......................................... 73

Figure 9.1.1: Saprolite Auger Sampling - Toroparu .......................................................................................... 82

Figure 9.2.4.1: Toroparu – 2013 Drillhole Collar Locations and traces – All Holes (486), Current MineralizedShape .................................................................................................................................................. 83

Figure 9.2.4.2: Toroparu – Drillhole Collar Locations – PEA Drilling through Aug. 2011 (342 holes).............. 84

Figure 9.2.4.3: Toroparu – Drillhole Collar Locations – Post PEA and Pre Targeted In-Fill (166 holes) ......... 85

Figure 9.2.4.4: Toroparu – Drillhole Collar Locations – Targeted In-Fill (Aug-Dec 2012) ................................ 86

Figure 9.2.4.5: Cross-Section showing Block Classification and Areas of Targeted Inferred Mineralization . 87

Figure 9.2.4.6: Cross-Section showing Block Au Grade and Areas of Targeted Inferred Mineralization ....... 88

Figure 9.2.4.7: Cross-Section showing Targeted Inferred Blocks and In-Fill Drill Data .................................. 89

Figure 9.2.4.8: Cross-Section Showing Current Block Classification and Areas of Targeted InferredMineralization that have been Converted to Measured and Indicated ................................................ 90



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project Page xiii


Figure 9.2.4.9: Cross-Section showing Block Grade for Current Resource model ......................................... 91

Figure 10.1: Toroparu – Plan Map of Drillholes and Traces, Showing Mineralized Envelope ........................ 96

Figure 10.3.1: Example of Un-cut Core Photograph for Drillhole TPD-103 .................................................... 97

Figure 11.5.1.1: Check Assay Scatter Plot of Actlab’s versus Acme Lab’s Au Assays ................................. 101

Figure 11.5.1.2: Check Assay Scatter Plot of Actlab’s versus Acme Lab’s Cu Assays ................................. 102

Figure 11.5.1.3: Check Assay Scatter Plot of Actlab’s versus Acme Lab’s Au Pulp Duplicate Assays ......... 103

Figure 11.5.1.4: Scatter Plot of Actlab’s versus Acme Lab’s Au Coarse Duplicate Assays ........................... 104

Figure 13.6.2.1: Effect of Sample P80 on Rougher Tailings Grade of Au and Cu ......................................... 124

Figure 13.8.3.1: Au Cyanide Leach Kinetics for Saprolite Fines ................................................................... 125

Figure 13.8.3.2: Au Cyanide Leach Kinetics for Coarse Saprolite ................................................................ 126

Figure 14.3.1: Toroparu Plan ......................................................................................................................... 143

Figure 14.3.1.1: Toroparu Plan “Fresh Rock” (grey) & “Saprolite” (orange) Domains .................................. 144

Figure 14.3.1.2: Toroparu Plan “Fresh Rock” (grey) & “Saprolite” (orange) Domains .................................. 145

Figure 14.3.2.1: Toroparu Perspectives “Fresh Rock” (grey) & “Anisotropy “ (red) ...................................... 146

Figure 14.3.3.1: Toroparu Domains ............................................................................................................... 147

Figure 14.3.3.2: Toroparu Model, Elevation -90 & Elevation -240, Mineralized/non Mineralized (green/grey)148

Figure 14.3.3.3: Toroparu Model Cross Sections, Mineralized/non Mineralized (green/grey) .................... 149

Figure 14.4.1: Lognormal Probability Plot, Au (g/t) Assays ........................................................................... 150

Figure 14.4.2: Lognormal Probability Plot, Cu (%) Assays ........................................................................... 151

Figure 14.6.1: Histogram of Bulk Density Data - Toroparu ........................................................................... 152

Figure 14.8.1: Variogram, Au (g/t) Modeled Anisotropic & Isotropic Variogram ........................................... 153

Figure 14.8.2: Variogram, Au (g/t) Modeled Anisotropic Variogram ............................................................. 154

Figure 14.8.3: Variogram, Cu (%) Modeled Anisotropic & Isotropic Variogram ............................................ 155

Figure 14.8.4: Variogram, Cu (%) Modeled Anisotropic Variogram .............................................................. 156

Figure 14.8.5: Anisotropy Points ................................................................................................................... 157

Figure 14.9.1: Resource Model Plans -160 Elevation -250 Elevation ........................................................... 158

Figure 14.9.2: Resource Model Representative Cross Sections................................................................... 159

Figure 14.9.3: Resource Model Perspective ................................................................................................. 160

Figure 14.10.1: Resource Model Confidence Classification -225, -200, -100 Plan Views ............................ 161

Figure 14.11.1: Optimized Resource Pits ...................................................................................................... 162

Figure 14.11.2: Cut-off/Price ......................................................................................................................... 163

Figure 14.12.3.1: Vertical Swath Diagram ...................................................................................................... 164

Figure 14.12.3.2: North-South Swath Diagram ............................................................................................. 165

Figure 14.12.3.3: East-West Swath Diagram ................................................................................................ 166

Figure 15.1.1: Cut-off Grade Calculation Graph (by different Gold selling price).......................................... 170

Figure 16.3.6.1: Toroparu Pit Optimization Results (Cashflow) .................................................................... 197

Figure 16.3.7.1: South-East Pit Optimization Results (Cashflow) ................................................................. 198



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project Page xiv


Figure 16.4.1: Toroparu Pit Geotechnical Sectors ........................................................................................ 199

Figure 16.5.1: Toroparu Pit Phase Design .................................................................................................... 200

Figure 16.5.2: South-East Pit Phase Design ................................................................................................. 201

Figure 16.5.3: Toroparu Final Pit Design – Measured and Indicated Blocks ................................................ 202

Figure 16.5.4: South-East Final Pit Design - Measured and Indicated Blocks .............................................. 203

Figure 16.5.5: Mine Plan Progress Maps - Year 5 End ................................................................................. 204

Figure 16.5.6: Mine Plan Progress Maps - Year 10 End ............................................................................... 205

Figure 16.6.11: Planned Mine Production Schedule (Material Movement) ................................................... 206

Figure 16.7.1: Pit and Waste Dump Locations .............................................................................................. 207

Figure 18.2.1: Tailings Management Area ..................................................................................................... 227

Figure 20.1.1.1: Location of Surface Water Samples .................................................................................... 249

Figure 20.1.1.2: Location of Well Sites .......................................................................................................... 250

Figure 22.1.1: Cumulative Cash Flow ............................................................................................................ 287

Figure 22.4.1: IRR Sensitivity ........................................................................................................................ 288

Figure 22.4.2: NPV Sensitivity to Discount Rate ............................................................................................ 289

Appendices Appendix A: Certificates of Authors



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


1 SummaryThis report is as a National Instrument 43-101 (NI 43-101) Technical Report on the Prefeasibility

Study of the Toroparu Gold Project (Toroparu Project or Project) prepared for Sandspring Resources

Ltd. (Sandspring) by SRK Consulting (U.S.), Inc. (SRK), Tetra Tech (Tt), Klohn Crippen Berger(KCB), Knight Piésold (KP), and FMG Engineering (FMG), (collectively the Consultants). Sandspring

is a Canadian based company continued under the laws of Ontario and trades on the TSX Venture

Exchange (TSX-V) under the symbol “SSP”. The quality of information, conclusions, and estimates

contained herein is consistent with the level of effort involved in the Consultant’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

Sandspring subject to the terms and conditions of its contract with SRK and relevant securities

legislation. The contract permits Sandspring 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 withSandspring. 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.

1.1 Property Descr ipt ion and Ownership

The Toroparu Deposit is located within Sandspring’s 98,214 hectare mineral exploration concession

area in the Upper Puruni River Area, Region 7 of northwestern Guyana, South America (referred to

as the “Upper Puruni Property” or the “Property”).

The airstrip at the Property is located at 06° 27’ North Latitude and 60° 03’ West Longitude, a

position approximately 220 km by air west southwest of Georgetown, the capital city of Guyana. The

2,500 ft airstrip at the Property camp is accessible by charter aircraft from Ogle airfield in

Georgetown. The 220 km trip takes approximately one hour.

Access to the Upper Puruni Property and the Toroparu Project by road includes 128 km via paved

highway from Georgetown to Bartica, a ferry crossing of the Essequibo River at Bartica to Itaballi,

200 km of gravel road to the Upper Puruni Property south gate, and 30 km within the property to the

Toroparu Project. Overland travel time is approximately 12 to 16 hours in the dry season from

August to May.

Heavy equipment and cargo is transportable by small, ocean going vessels and barges on the

Essequibo River to Itaballi. There it is loaded on to trucks for the 7-10 hour, 225 km overland journey

to the Toroparu Project crossing the Puruni River by ferry at the town of Puruni Landing, 60 km from

Itaballi (Figure 1.1.1).

The Property is comprised of seven Small Scale claims, 167 contiguous Medium Scale Prospecting

Permits and 13 Mining Permits that together cover an area of 184,693.8 acres or 74,742.9 hectares,

and five contiguous Prospecting Licenses covering an area of 57,997 acres or 23,471 hectares. ETK

Inc. (“ETK”), a private company in Guyana, and a wholly owned subsidiary of Sandspring, owns the

rights to the Upper Puruni Property. Sandspring acquired its interest in ETK, and thus its interest in

the Upper Puruni Property, on November 24, 2009.





1.2 Geology and Mineralization

The Toroparu Gold Project is located in northwestern Guyana within the Guiana Shield. The Guiana

Shield encompasses Venezuela, Guyana, Surinam, French Guyana and parts of northern Brazil.

This portion of the Guiana Shield is composed of alternating volcano-sedimentary belts and large

granitoid batholiths of Paleo-Proterozoic age. These supracrustal rocks form the northern part of theGuiana Shield, which is the northern segment of the Amazonian Craton of South America, and a

dismembered portion of the West African Craton. The West African Craton and parts of the Guiana

Shield are well known for hosting gold deposits.

The concession package of Sandspring Resources (1000 km²) is located in the Upper Puruni area,

in between the Cuyuni and Mazaruni rivers, in the north-west part of Guyana. The regional geology

is not well documented due to dense tropical vegetation and thick lateritic/saprolitic weathering

profiles, causing a general lack of bedrock exposure. Sandspring has generated a litho-structural

sketch of the Upper Puruni area, using all available regional data (Project and public data): airborne

magnetics and radiometrics, topographic data, satellite imagery, existing geological maps, and

regional geochemical data.The northeastern half of the Upper Puruni concession is underlain by thick volcano-sedimentary

sequences consisting of alternating mafic, intermediate and to a lesser extent, felsic volcanic flows

and pyroclastics, with intercalated sedimentary successions, generally metapelites and greywackes.

These formations form the Puruni volcano-sedimentary (VS) belt which extends in a northwesterly

direction in between two large plutonic areas, the Aurora batholith located to the northeast of the

concession, and the Putareng batholith underlying most of the southwestern part of the property.

Regional metamorphic grade is greenschist facies and can reach the amphibolite facies in the

vicinity of the granitoid intrusions.

The Putareng batholith corresponds to a calc-alkaline composite intrusive complex, ranging in

composition from granite and tonalite to diorite. Exploration revealed the existence of small, more orless elongated, intra-belt plutons, generally of tonalitic to quartz-dioritic composition. The Toroparu

gold-copper deposit developed along the contact zone of one of these small intrusive bodies.

The Upper Puruni area is marked by sets of NW to WNW and NNW to N-S lineaments. The NW

oriented features seem to constitute typical belt parallel shearing structures, following lithological

contact zones and dominating the regional trend of the belt.

The knowledge of the local geology is mainly based on limited geological mapping of Saprolite in

road and river cuts (alluvial workings), the current Toroparu open pit (entirely in Saprolite) and the

core drilling in the Toroparu deposit area. The dominant lithologies of the Toroparu deposit are

metamorphosed (greenschist facies), often fine grained, acid to mafic volcanics (pyroclastics) and

sediments. The Toroparu gold-copper deposit occurs along the northwestern boundary of a tonaliticto quartz dioritic intrusion, close to the southeastern edge of the pluton.

On a deposit scale, the western part of the Toroparu mineralization system, including the West Zone,

and Main West Zones and the SE satellite deposit are predominantly hosted by intrusive: ne. In the

eastern part of the deposit area (Main East Zone), the mineralization forms an elongated cloud along

a contact zone of a greenschist metamorphic volcanic sequence, draped over a deeper seated

tonalitic intrusive. The different zones of mineralization are interpreted to be separated by WNW and

NNW oriented fault sets.





Core logging defines irregular zones of silicification and sericitization/chloritization, with associated

epidote. Carbonate is ubiquitous in most lithologies as small disseminated grains in the

groundmass, and is abundant and associated to quartz in veinlets. Quartz-carbonate veinlets are

common in the gold-copper and gold-only mineralized zones. The veinlets (mm to cm widths)

represent a fine fracturing network that defines the mineralization system

The deposit forms a west-northwestern oriented mineralized corridor, where the gold and copper

mineralization appears to be controlled by a moderately developed, probably dilational type of brittle

fracture/veinlet stockwork.

The mineralization system corresponds to a 2.7 km long and 200 to 400 m wide, WNW oriented

body, consisting of a low grade gold mineralized envelope surrounding several more or less east-

west oriented lenses of higher grade. Mineralization extends to depths of over 400 m.

Exploration and definition core drilling revealed that the larger part of the deposit is comprised of

several more or less east-west oriented lenses:

The Main Eastern lens (Main Zone), containing the larger part of the resource and displaying

in its core zone the highest average Au and Cu grades;

The Main Western lens, marked by lower average gold grades and very low grades of Cu;

and

The SE lens, carrying mainly gold mineralization, forms a near-by satellite body, 1.2 km SE

of the Main Zone.

In the center of the Main East Zone there is clearly a relation between the intensity of the fracturing

and the grade of gold and copper mineralization. The same comment can be made for the SE Zone

mineralization, but involves mainly higher grade gold as copper is nearly absent in this satellite

deposit. Field observations in the historic mining pit, logging of core holes drilled parallel to higher

grade zones (>1.5 g/t), and results of the borehole scanning survey reveal a predominant east-west

fracture set. Drilling is at angle holes with azimuth bearing orthogonal to the east-west fracture/vein

sets.

A significant resource definition drilling program was completed by Sandspring from the end of 2006

to the end of 2012 on the Toroparu deposit, which provides the basis for understanding the Project

geology. Drill core logging allowed for detail understanding of the lithology, structure, and the gold–

copper mineralization in bedrock. Up to the end of 2010 all exploration was focused on the Toroparu

deposit and immediate surrounding areas. Several resource estimates for Toroparu were completed

and presented in NI 43-101 technical reports (P&E, 2009 and 2010).

At the end of the 1st quarter of 2011 an exploration campaign was launched with the main objective

to test the gold potential of the company’s 1,000 km² Upper Puruni Concession. This program

consisted of systematic regional and semi-regional geochemistry sampling, geological mapping,

ground geophysics and reconnaissance exploration drilling. The results of that work are a total of 10

gold anomalies detected, which form a cluster around the Toroparu deposit. The Toroparu gold-

copper deposit was advanced to a scoping study stage with and updated Preliminary Economic

Assessment (PEA) NI 43-101 Technical Report (P&E, 2012) in early March of 2012.

After March of 2012, Sandspring concentrated all efforts on advancing the Toroparu deposit to

prefeasibility study, and the primary focus was deposit definition and in-fill drilling. Sandspring

conducted additional drilling from September 2011 through December 2012, which is included in the





current updated resource estimate as presented in Section 14 of this report. That additional drilling

was composed of the following:

Post PEA drilling from August 2011 to September 2012, a combination of in-fill and step-out

holes to further define the extents of mineralization: 166 holes for 44,096 m; through hole

TPD-426; and A program of targeted infill drilling to convert Inferred classification mineral resources to

Measured or indicated classification conducted from September through December 2012,

resulted in 48 holes for 12,163 m in both the Main Zone and the Southeast Zone.

Post PEA drilling accounts for an additional 214 drillholes for 56,259 m, or an increase of 38% in

total meters of drilling that are used to update the resources in this report.

The Toroparu gold-copper deposit is sufficiently drill-defined to support feasibility level study.

1.3 Mineral Processing and Metallurgical Testing

Ores tested from the Toroparu deposit are separated into three distinct categories, saprolitic Au ore,

hard rock ores containing Au and recoverable Cu, and hard rock ores containing Au with minor

amounts of Cu. Au recoveries expected from these ores are 98%, 88%, and 95%, respectively. Cu

recovery of the Au-Cu ores is expected to be 91% in a marketable concentrate.

Sandspring initiated several metallurgical testwork programs beginning in 2009 to obtain information

regarding the physical properties of the various ore grade mineralization in the deposit and their

response to comminution, gravity concentration, rougher and cleaner flotation, and cyanide leaching.

Metallurgical testwork and financial analysis tradeoffs were performed to show that processing the

deposit with both flotation and cyanide leaching, depending on Cu content, would provide economic

benefit due to the recovery of a marketable Cu concentrate.

Testwork has shown that both generalized ore designations, Gold Ore with Average Copper (ACO)also described elsewhere in the report as “Au/Cu Ore”, and Gold Ore with Low Copper (LCO), also

described elsewhere in the report as ”Au Ore”, benefit from gravity concentration prior to further

processing. Gravity gold recoveries of 38% were demonstrated for both ACO and LCO ores.

Flotation recoveries achieved from ACO ore were 91% Cu and 42% Au, in addition to gravity gold

recoveries. Testwork shows that both Cu and Au recoveries from LCO ore were acceptable, but the

relative loss in Au recovery versus a cyanide leach was not offset by Cu flotation recovery.

Cyanide leach testwork was conducted to determine the amenability of the ACO and LCO ores. It

was determined that ACO flotation cleaner tailings and LCO gravity tailings leach recoveries were

8% and 58%, respectively, in addition to gravity and flotation recoveries.

Cyanidation testwork was conducted to determine the amenability of the ACO and LCO ores to

cyanide leaching following gravity concentration and flotation. It was determined that ACO flotation

cleaner tailings and LCO gravity tailings leach recoveries were 8% and 58%, respectively, in addition

to gravity and flotation recoveries.

Overall Au recoveries from ACO and LCO ores were determined to be 88% and 95%, respectively.

These recoveries include gravity concentration, flotation, and cyanide leaching.





In addition to the primary hard rock ACO/LCO ores, saprolitic cover ore was also tested for

amenability to gravity concentration, flotation, and cyanide leaching. Testwork showed that gravity

concentration and flotation do not achieve desired recoveries compared to whole ore cyanidation.

Recovery achieved for 72 hour whole ore cyanide leaching was approximately 98% for both RoM

saprolite fines, and coarse saprolite ground to P80 129 µm.

The testwork on ACO Composite samples has shown that the overall gold recovery from gravity

separation and gravity tailings leaching is higher than the case of gravity separation, cleaner flotation

and cleaner tailing leaching. However, due to the lower proportion of copper reporting to the cleaner

tailing leaching stage, the cyanide consumption is lower for this process than in the direct gravity

tailing leaching. This indicates that flotation of gravity tailings plus leaching of the flotation cleaner

tailings would be a better option than leaching of gravity tailings for this type of material.

Similar to the ACO Composite sample testwork, overall gold recovery from direct cyanide leaching of

LCO composites gravity tailings is higher than recovery from cyanide leaching of the rougher

concentrate. Moreover, the cyanide consumption is also lower for the direct cyanide leaching of the

gravity tailings, indicating this method would be more favorable for LCO ore types.

1.4 Mineral Resource Estimate

SRK estimated the Mineral Resources for the Toroparu deposit during January 2013; the estimation

was carried out in compliance with NI 43-101 regulations and CIM standards. The estimate utilized

all drilling available through December 27, 2012. The estimate was prepared by Frank Daviess,

Associate Principal Resource Geologist, SRK Consulting (U.S.), Inc., of Lakewood, Colorado in

accordance with National Instrument 43-101 (NI 43-101). The resources are “in-pit resources”; the

resource model was investigated with a Whittle™ pit optimization to ensure a reasonable stripping

ratio was applied and a reasonable assumption of potential economic extraction could be made.

Table 1.4.1 summarizes the resource for the Main and SE Zones at a 0.30 g/t Au cut-off within the

global optimal pit shells.





Table 1.4.1 Resource Statement @ 0.30 g/t Au cut-off as of March 31, 2013

Resource Classi ficat ion Tonnes Au Au oz Cu Cu(All rock types) (000’s) (g/t) (000’s) % (Mlb)

Main ZoneMeasured 41,542 0.98 1,307 0.109 100

Indicated 185,957 0.87 5,203 0.082 334Measured & Indicated 227,500 0.89 6,510 0.087 434Inferred 127,756 0.74 3,045 0.042 118South East ZoneMeasured 2,905 0.97 91 0.037 2Indicated 9,836 0.93 294 0.035 8Measured & Indicated 12,741 0.94 384 0.036 10Inferred 1,768 0.78 45 0.041 2 Al l ZonesMeasured 44,447 0.98 1,398 0.104 102Indicated 195,793 0.87 5,497 0.079 342Measured & Indicated 240,240 0.89 6,894 0.084 444Inferred 129,525 0.74 3,090 0.042 120

Source: SRK, 2013

1. Mineral resources are inclusive of mineral reserves;2. All resources in the revised mineral resource statement are In-Pit resources reported within an optimized pit shellabove an economic cut-off grade of 0.30 g/t Au. The economic cut-off grade was determined using a gold price ofUS$1,350/oz Au, an average metallurgical recovery of 95.9% for gold, Processing + G&A costs of US$11.49/t, andincludes US$112/oz Au for freight, smelting, refining and royalties. Copper metallurgical recovery used was 91%.Pit slopes used in the pit optimization were 45 degrees, and the mining costs used were US$2.06/t for fresh rock.

3. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is nocertainty that all or any part of the Mineral Resources estimated will be converted into Mineral Reserves;

4. Mineral Resources are reported in accordance with Canadian Securities Administrators (CSA) National Instrument43-101 (NI 43-101) and have been estimated in conformity with generally accepted Canadian Institute of Mining,Metallurgy and Petroleum (CIM) "Estimation of Mineral Resource and Mineral Reserves Best Practices" guidelines;

5. The grades for Au and Cu were estimated separately, and presented as associated average metal grades at the Aucut-off;

6. Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, andnumbers may not add due to rounding;

7. The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there has beeninsufficient exploration to define these Inferred resources as an Indicated or Measured mineral resource and it is

uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category;and

8. The mineral resource estimate for the Project was calculated by Frank Daviess, MAusIMM, R.M. SME, AssociateResource Geologist of SRK Consulting, Inc. in accordance with the Canadian Securities Administrators NationalInstrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and generally accepted CanadianInstitute of Mining, Metallurgical and Petroleum “Estimation of Mineral Resource and Mineral Reserves BestPractices” guidelines (“CIM Guidelines”).

1.5 Mineral Reserve Estimate

The estimates of mineral reserves are effective as of March 31, 2013 and are presented in Table

1.5.1. The prefeasibility (PFS) models an open pit mine with a Proven and Probable mineral reserve

containing 4.1 Moz of gold and 211 Mlb of copper, which in contained gold terms represents 60% ofthe 6.9 Moz (in resource-pit shell) Measured and Indicated mineral resource estimate, as disclosed

herein.

Measured and Indicated resources were used for conversion to Proven and Probable reserves within

the optimized PFS pit designs. The mineral reserve (in-pit) cut-off grades (CoGs) used were 0.35 g/t-

Au for saprolite and 0.38 g/t-Au for fresh rock, which correspond to a gold price of US$970/oz Au for

saprolite, and US$1,070/oz Au for fresh rock, respectively.





The reserves are contained within the Toroparu pit (Toroparu Pit) and South-East pit (South-East Pit)

and are associated with 468.9 Mt of waste and a life of mine stripping ratio of 3.69:1.

Reserves are valid at the time of estimation and include CoG assumptions made before the final

economic model is published. SRK confirmed the overall Project economics are favorable at the

approximate four-year moving average gold price of $1400/oz Au.

Table 1.5.1: March 31, 2013 Mineral Reserve Estimate

MaterialReserveClassification

Tonnes(000's)

Gold(g/t)

Gold(koz)

Copper(%)

Copper(Mlb)

AuEq(g/t)

AuEq**(koz)

Saprolite Au OreProven 1,621 0.95 50 0.09 *** 0.95 50Probable 3,400 0.90 98 0.10 *** 0.90 98Proven + Probable 5,022 0.91 148 0.10 *** 0.91 148

Fresh Au OreProven 13,976 0.93 419 0.05 *** 0.93 419Probable 56,333 0.88 1,587 0.05 *** 0.88 1,587Proven + Probable 70,309 0.89 2,006 0.05 *** 0.89 2,006

Fresh Au/Cu Ore

Proven 14,183 1.27 581 0.20 64 1.62 740

Probable 37,597 1.14 1,373 0.18 147 1.44 1,740

Proven + Probable 51,780 1.17 1,953 0.18 211 1.49 2,480

All Ore TypesProven 29,780 1.10 1,049 0.13 64 1.26 1,209Probable 97,331 0.98 3,058 0.10 147 1.09 3,425Proven + Probable 127,111 1.00 4,107 0.11 211 1.13 4,634

Source: SRK, 20131. Mineral reserves are based on a gold cut-off-grade (CoG) price of US$1,070/oz. for fresh rock and US$970/oz. for

saprolite. Cash flow Base Case used a gold price of US$1,400/oz. and copper price of US$3.25/lb.;2. Open pit reserves assume complete mine recovery;3. Open pit reserves are diluted (further to dilution inherent in the resource model and assumes selective mining unit of 5 m

x 5 m x 5 m);a. Contained In-situ gold ounces do not include metallurgical recoveries of 96% for gold in saprolite (Oxide), 85% for

gold in Au/Cu fresh rock, 91% for copper in Au/Cu fresh rock, and 96% for gold in Au fresh rock;b. ** AuEq= Gold Equivalent ounce calculated using US$1,403/oz. Au (US$1,394/oz. after refining), US$3.47/lb. Cu

(US$3.17/lb. after NSR deductions), 85.46% gold recovery, 91% copper recovery, Formula 1% Cu = 1.714 g/t-Au);c. *** No copper will be recovered from this ore type (and thus the Gold Equivalent Grade = Gold Grade);

4. Waste tonnes within pit is 468.9 Mt at a strip ratio of 3.69:1 (waste to ore);

5. An open pit CoG of 0.35 g/t-Au saprolite and 0.38 g/t-Au fresh rock was applied to open pit resources constrained by thefinal pit design;

6. Mineral reserve tonnage and contained metal have been rounded to reflect the accuracy of the estimate, and numbersmay not add due to rounding;

7. “(000)” = thousands, “g/t” = gram per metric tonne, “koz” = thousand troy ounces. Ore tonnes are rounded to the nearestone thousand tonnes, gold to nearest 1000 oz Au, gold grade to nearest 0.01 g/t Au, copper rounded to nearest millionpounds.

8. The mineral reserve estimate for the Project was calculated by Fernando P. Rodrigues, BSc, MBA MMSAQP #01405QPof SRK Consulting, Inc. in accordance with the Canadian Securities Administrators National Instrument 43-101 –Standards of Disclosure for Mineral Projects (“NI 43-101”) and generally accepted Canadian Institute of Mining,Metallurgical and Petroleum “Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines (“CIMGuidelines”); and

9. Reserves Effective Date: March 31, 2013.

1.6 MiningMore accurate resource and geologic models produced over the course of 2011/2012 during the

prefeasibility definition drilling campaigns identified two geographically distinct populations of gold

bearing saprolite and fresh rock ores, distinguishable by their copper sulfide contents, ore with

recoverable copper being defined as “Au/Cu Ore”, also described elsewhere in the report as ACO

(Average Copper Ore) and without recoverable copper content as “Au Ore”, also described

elsewhere in the report as LCO (Low Copper Ore). The mine plan and production schedule defined

in the PFS were optimized for higher metallurgical recovery by processing these ores separately in





Mining will consist of a conventional open pit operation including drilling and blasting, loading and

hauling. It is planned for a normal hydraulic excavator and haul truck mining fleet to be utilized along

with supporting auxiliary mining equipment (motor graders, water trucks, etc.). For mine planning

purposes and expected fleet numbers, 5 to 22 m3 (bucket size) excavators and 50 to 133 t off-

highway diesel haul trucks have been used.

Drilling and blasting are planned to be performed on 10 m benches in both pits. This matches a

multiple of the block size in the geological block model. Due to the expected selective mining that will

be required for ore mining, loading and hauling are planned to be performed using a half-bench

height for ore, and full bench heights for waste handling.

The Toroparu pit is planned to be developed first, with the process facility to be constructed adjacent

to this pit. This will minimize the ore haulage requirements during the early years of the Project. The

final pit and dump designs are presented in Figure 1.6.2.

The Project plans to use proven technology, with no requirement for untried or untested technology.

Table 1.6.1 summarizes the planned mining production schedule.

Table 1.6.1: Planned Mine Product ion Schedule

Year Saprolite ACOFreshRock

LCOFreshRock

Waste Saprolite ACOFreshRock

LCOFreshRock

ACOFreshRock

LCOFreshRock

Unit kt kt kt kt Au Grade

(g/t) Au Grade

(g/t) Au Grade

(g/t)Cu Grade

(%)Cu Grade

(%)

-3 0 0 0 0 0 0 0 0 0

-2 1,545 20 0 698 1.09 1.15 1.47 0.20% 0.07%

-1 1,466 4,044 353 7,137 0.85 1.09 0.75 0.24% 0.07%

1 402 10,324 2,898 26,375 0.82 1.12 0.81 0.20% 0.06%

2 335 4,981 2,684 31,999 0.69 1.20 0.79 0.21% 0.06%

3 237 4,872 3,056 32,132 0.83 1.26 0.78 0.20% 0.06%

4 478 1,935 5,568 31,371 1.01 1.05 0.89 0.14% 0.05%

5 203 5,628 3,131 31,907 0.80 1.12 0.85 0.18% 0.06%

6 80 2,258 3,560 41,855 0.58 0.95 0.92 0.16% 0.05%

7 7 4,576 3,611 44,385 0.54 1.23 0.92 0.18% 0.05%

8 234 4,985 14,476 37,200 0.81 1.26 0.84 0.14% 0.05%

9 6 324 1,269 45,490 0.65 0.94 0.94 0.17% 0.04%

10 0 1,754 9,957 33,335 0.00 1.19 0.84 0.15% 0.04%

11 29 1,986 10,710 37,428 0.63 1.59 0.99 0.15% 0.05%

12 0 61 2,194 48,125 0.00 1.38 1.01 0.17% 0.02%

13 0 2,319 2,180 17,301 0.00 1.07 0.92 0.19% 0.04%

14 0 1,713 4,661 2,139 0.00 1.28 0.97 0.15% 0.04%

Total 5,022 51,780 70,309 468,875

Source: SRK ACO terminology: Material where the copper content is above 0.09%. -- Material subject to Flotation – a.k.a Au/Cu OreLCO terminology: Material where the copper content is below 0.09%.- Material subject to Cyanide Leaching – a.k.a Au Ore

1.7 Processing Recovery Methods

The Toroparu processing facility will be developed in three phases over the life of mine (LoM). This is

to accommodate for variation in ore types over the production schedule.

The first period is characterized as the preproduction phase. This phase is estimated to consist of

two years, during which the facility will process 3,250 t/d of saprolite through a carbon-in-pulp (CIP)





leach circuit using a refurbished ball mill already in the possession of Sandspring. The CIP circuit

and downstream process equipment will be designed to expand to the full production rate for

processing a combination of flotation tailings and saprolite during Phase 1.

Phase 1 consists of processing 15,000 t/d of Gold Ore with Average Copper (ACO) via flotation of

gravity tailings with cyanide leaching of the cleaner scavenger flotation tailings via a CIP circuitalongside saprolite. It is estimated that this will occur for the first five years of the mine life. Based on

performed metallurgical testwork recovery by flotation is expected to result in a marketable Cu

concentrate with grade of approximately 21% Cu.

Phase 2 consists of processing 15,000 t/d of Gold Ore with Low Copper (LCO) via CIP leaching and

7,500 t/d of ACO via flotation with CIP leaching of the cleaner scavenger tailing. This phase

continues over the remaining LoM.

A simplified process flow diagram is provided in Figure 1.7.1.

1.8 Tailings Management Area

The Tailings Management Area (TMA) will be located on the northeast side of the Toroparu property,

approximately 8 km from the mine area. This facility will be staged and operated in three

independent cells that will operate separately at different stages of the service life of the facility as

follows:

Cell 1 will operate up to the first quarter of Year 5 of full production;

Cell 2 will operate from the second quarter of Year 7 to second quarter of Year 9; and

Cell 3 will operate from the third quarter of Year 11 to end of the mine life at Year 16.

Detoxed tailings will be discharged into the TMA cells from the crests of the tailings dams, and

supernatant water volumes reclaimed to the plant by a floating pump barge positioned on the decant

ponds of the cells. Excess water volumes will be discharged through spillways into three collectionsponds located adjacent to the cells for monitoring and control before release to the environment.

The design is based on the near surface ground conditions being sufficiently impermeable to prevent

ground contamination without a geosynthetic liner. Fill and compaction of in-situ saprolite was

included in cost estimates. The TMA has a storage capacity of up to 143 Mt.

1.9 Project On-Site Infrastructure

The overall site location, orientation and layout of the on-site infrastructure was based on the criteria

of providing sufficient space for the process and mine facilities, while placing them in close proximity

to the mine pit locations and TMA facility (Figure 1.9.1). Key site infrastructure is located at a safe

elevation above the 100-year storm event boundary limit.

The on-site infrastructure includes the on-site service roads and river crossings; water supply and

treatment; power supply and distribution; mine support facilities; process support facilities; entry

station; and the man camp facilities. Site access roads, which interconnect the various site service

areas are segregated to the maximum extent possible from the mine haul roads.





vegetation is primarily secondary growth, mixed forest that shows indications of human disturbance.

The mining concession is in a disturbed area consisting mainly of swamp, Morabukea forest, and

mixed forests.

The area is drained by the Puruni River and several tributaries, including the Wynamu River.

Relatively little historical water quality data is available. Surface water quality samples were collectedat two locations in the Puruni River and one location in the Wynamu River during the baseline

sampling in 2007, 2008 and 2010. The majority of the surface water quality results were below the

IFC effluent requirements. Five groundwater monitoring wells have been installed at the site and

samples were collected in 2007, 2008 and 2010. The groundwater levels were lower in the dry

season, which indicates that the water table is recharged by precipitation. Groundwater quality

results indicated slight exceedances of some constituents (iron, pH) and relatively high exceedances

of total suspended solids.

No archaeological resources of interest have been identified at the site.

The fauna documented during the biological survey were fairly common species in Guyana. One

mammal (Jaguar) is listed as globally threatened species by CITES (Convention on InternationalTrade in Endangered Species of Wild Fauna and Flora), which is an international trade agreement

regarding wild animals and plants. Up to 160 bird species were noted during any one survey, and

migratory birds have been recorded. There were 52 individual fish species recorded. There were 32

herpetofauna species recorded and 24 Arthropod orders recorded. No locally rare, threatened or

endangered species were recorded, however a number of species were identified as having status

under CITES.

Mining Wastes

Tailings and waste rock will be produced by the mining operations, plus there will be a Low Grade

Ore (LGO) stockpile. Metallurgical testing was done on three main ore types (saprolite, ACO and

LCO). The metallurgical tailings generated from each of the ore types, the dominant bedrock

lithologies in the waste rock and the LGO were tested for their Acid Rock Drainage/Metal Leaching

(ARD/ML) potential during geochemical characterization studies conducted by KCB (2013). The

geochemical testing was based on Canadian and industry guidance documents for the prediction of

ARD/ML. In the absence of specific Guyanese water quality guidelines, the water quality guidelines

applicable to British Columbia and Canada were applied as a screening tool to assess laboratory

leachate results.

Lithologies identified as potential waste rock included acid intrusive; fragmental mafic volcanic;

granodiorite; mixed facies; saprolite; and undifferentiated intermediate volcanics. The waste rock,

metallurgical tailings and LGO samples were analyzed for mineralogical analysis, solid-phase

elemental analysis and Acid-base Accounting. In addition, the waste rock samples were analyzed forShake Flask Extraction (SFE) and Net Acid Generation (NAG) tests. The metallurgical tailings

samples were also subjected to supernatant aging tests.

The majority of waste rock lithologies and LGO samples contained very low sulfide-sulfur

concentrations, which indicates negligible ARD risk. The waste rock lithologies and LGO were

classified as not-Potentially Acid Generating (N-PAG), with the exception of the saprolite samples.

The saprolite samples were predominantly Acid Generating (AG) and PAG, with a low amount of

available Neutralization Potential (NP). The saprolite waste is approximately 15% of the total waste.





A standard operating procedure will be developed to encapsulate the saprolite waste within the fresh

rock waste rock facility.

All metallurgical tailings samples contained low to negligible sulfide-sulfur concentrations and were

classified as N-PAG. It was also noted that the saprolite metallurgical tailings contained little to no

reactive carbonate minerals, which indicates that the NP present in the saprolite metallurgical tailingsis related to the lime added during the metallurgical processing.

The leachate extraction tests results indicated that waste rock and LGO may develop alkaline

drainage with elevated concentrations of aluminum, selenium, chromium and, to a lesser extent,

elevated concentrations of copper and phosphorus based on freshwater aquatic life guidance. The

supernatant aging test results for the metallurgical tailings indicated that alkaline drainage with

elevated concentrations of aluminum, selenium, chromium, arsenic, cobalt, copper, iron,

molybdenum, WAD cyanide and sulfate may develop. Iron and WAD cyanide concentrations in the

LCO tailings supernatant were also elevated. Although the metallurgical tailings leachate extraction

test results indicated elevated concentrations that may be soluble and mobile under laboratory test

conditions, the results do not imply that they will be elevated to levels above these guidelines undersite-specific field conditions, rather they identify elements that are prone to leaching. The tailings

management area (TMA), will receive a combination of precipitation, water treatment plant brine, and

supernatant from the tailings slurry. The TMA water quality will be influenced by contributions from all

these sources. Natural attenuation of these elements will occur primarily through dilution from high

precipitation, which is reported to average about 2.6 m annually at the Toroparu Project site.

Additional geochemical testing would be required to assess future site-specific field conditions;

however the current status of the geochemistry program is in line with a PFS project.

Environmental Impact Analysis

Potential environmental impacts were evaluated as part of the ESIA preparation. The most

significant associated impacts and associated planned mitigation strategies are as follows:

• Soil erosion and sedimentation from disturbed areas will be minimized or avoided by

implementation of BMPs to be described in an Erosion and Sediment Control Plan;

• Emissions of fugitive dust and combustion products will be mitigated by scheduling land-

clearing activities to less windy days, limiting vehicle speeds and employing dust

suppression techniques;

• Accidental spills of fuels, oils and grease will be minimized as described in a spill Prevention

and Contingency Plan; and

• Impacts to wildlife will be mitigated by minimizing the disturbance areas and by maintaining

wildlife corridors through cleared areas.

An Environmental Management Plan and a monitoring program have been developed to mitigate

and monitor changes to the both the physical and socio-cultural environments.

Social or Community Impact

There are no formal or established communities or settlements in the immediate vicinity of the mine

site. Since there are no communities in the area of direct influence of the Project, the social baseline

was developed based on a broader area of indirect influence, such as the communities which could

serve as entry points to the mine site and vulnerable communities that could supply employees.





The total estimated initial cost to design and build the Toroparu Project identified in this report is

US$464 million. A summary overview of the estimate by area is presented in Table 1.13.1.

Table 1.13.1: Summary of Capital Costs by Area

PFS Capital Cost Estimates

(US$ Millions) Scope

Initial

Capital(Pre-Prod)

Expansion and

SustainingCapital

LoM

Capital

Fresh Rock Pre-Stripping SRK $24 $0 $24Mine Site Preparation / Roads SRK $2 $0 $2Mining Equipment SRK $69 $168 $237Milling Circuit Tt $75 $0 $75Leaching Circuit Tt $36 $0 $36Flotation Circuit Tt $24 $0 $24Process Plant Infrastructure Tt $6 $0 $6Plant Expansion Tt $0 $50 $50Tailings Storage Facility Tt/KCB $16 $63 $80On-Site Infrastructure Tt $11 $11 $22Power Generation Tt $27 $0 $27Water Management Tt $9 $0 $9

Camp and Ancillary Buildings Tt $25 $0 $25Port and Logistics FMG $9 $0 $9 Access Road Upgrades FMG $33 $0 $33Construction Indirects (incl. EPCM) Tt $79 $0 $79Owner's Costs (Incl. Closure) Sandspring $20 $15 $35

Sub-Total Project Capital Costs $464 $307 $771Mining Contingencies (Site Prep + Equip) SRK $4 $8 $12Process and Infrastructure Contingencies Tt/KCB/FMG $32 $0 $32Owner’s Costs Contingencies Sandspring $2 $4 $6

Total Contingencies All $37 $13 $50

Total Capital Requirement All $501 $319 $821

Contribution from Saprolite Au OreMargin

All ($37) $0 ($37)

Total Project Costs w/ Contingencies All $464 $319 $784

The aggregate capital estimate is considered to be within a +30% / -25% weighted average accuracy

of actual costs. Base pricing will be in Q1 2013 United States dollars, with no allowances for inflation

or escalation beyond that time.

The contingency cost is based on the total direct and/or indirect costs and are included to account for

unanticipated costs within the scope of the estimate. The contingency percentage allowances vary

and are individually assessed based on the accuracy of the quantity measurement, type and scope

of work, and price information for the capital cost estimate.

Operating Costs

The PFS estimate is based on independent reputable vendor quotations and where not available,

first principal calculations completed by the Consultants. Each Consultant used their standard

estimating system to calculate its respective operating costs for the Project.

Operating costs have been prepared in Q4 2012 US dollars and exclude:

Contingency;

Escalation;

Taxes (VAT); and





Import Duties.

Imported equipment, materials, and operating supplies are not subject to Taxes (VAT), import or

other duties as per the Mineral Development Agreement.

The operating cost estimates have been assembled by area and component, based upon estimated

staffing levels, consumables and expenditures according to the mine and process design. Life-of

mine (LoM) operating costs are shown in Table 1.13.2 (rounded to nearest US$1,000).

Table 1.13.2: Operating Cost Li fe of Mine, US$000s

Area Labor (US$000s) Expenses (US$000s) US$/t-Mined US$/t-Mi ll

Mine 159,656 946,003 1.86 8.70

Processing 22,238 1,313,509 n/a 10.51

G&A 92,515 81,306 n/a 1.37

Total Operating 274,408 2,340,818 n/a 20.57

1.14 Economic Analysis A discounted cash flow model was created to evaluate the Toroparu Project assuming the Project is

100% equity financed. All revenues and costs are expressed in US dollars.

Mining cost estimates were provided by SRK and process costs were provided by Tt. Offsite

infrastructure costs were provided by FMG, and Owner’s cost by Sandspring. Additional costs such

as refining, royalties and administrative costs provided by SSP were subtracted from the revenue to

calculate an estimated cash operating margin.

The Prefeasibility makes use of Proven and Probable reserves only.

An income tax rate of 30% is used based on the rate defined in the Mineral Agreement that governs

the Project. The resulting cash flow in each year of the Project life is discounted back to January ofYear -3 to determine the estimated discounted cash flow at a 5%, 8% and 10% discount rate. Using

this same data, the estimated internal rate of return and the undiscounted cash flow were also

determined.

The major input parameters to the model include gold and copper prices, sustaining capital,

operating costs, mining rates, and estimated taxes and royalties. Additionally, several minor

assumptions throughout the model such as working capital, environmental accruals and depreciation

rates affect the estimated Project economics to a lesser degree.

SRK and Sandspring prepared a detailed financial model (the Financial Model) estimating cash flows

by year for the forecast mine life.

The model was also based on the following Project basic schedule:

Construction period: 3 years, two of which are also the saprolite preproduction years; and

Production period: 16 years.

Start-up – For the purpose of the model, the plant is estimated to commence the processing of

saprolite ore on January 1st of Year -2 and will continue for two years (Years -2 and -1). Fresh rock

starts to be processed on January 1st of Year 1.





Working capital – Working capital was included in the model. This estimate was considered as 20%

of all operating costs for each period.

Escalation – The components of the economic model were based on the following:

Base capital pricing for the Project is in Q1 2013 United States dollars, with no allowances

for inflation or escalation beyond that time;

Equipment quotes from vendors were obtained in Q4 2012; and

Operating costs were prepared in Q4 2012 terms.

All financial results are based in Q4 2012 and Q1 2013 dollars and no escalation has been assumed

for the metal prices or cost inputs.

Closure Costs – Estimated closure costs for the dismantling of the plant and infrastructure, and for

long-term water treatment, were estimated at US$16 million, including contingencies. For the

purposes of the financial model, these costs were incurred over a period of one year, following the

processing of the last ore through the mill. No credit was provided in the model for the potential

salvage value of equipment.Table 1.14.1 presents further details of the economic results. Payback from plant start of operations

(January of Year 1) is 2.6 years. Figure 1.14.1 shows the after-tax sensitivities to capital costs,

operating costs and metal prices at the 5% discount rate.





Table 1.14.1: Project Evaluation Economic Results

Descript ion Value (US$) Units

Market PricesGold $1,400 /Au-ozCopper $3.25 /Cu-lb

Estimate of Cash Flow (all values in $000s)Gross Income $/Au-ozPayable Gold (Doré+Concentrate) $5,190,263 $1,400.00Payable Copper $593,609 $160.00Gross Income $5,783,872

$/Au-ozTreatment Charges ($39,353) ($10.61)Refining Charges ($18,937) ($5.11)Predicted Penalties ($2,253) ($0.61)Freight Insurance Cost ($71,641) ($19.32)Gross Revenue $5,651,687

$/Au-ozGuyana Au Royalty ($413,937) ($111.65)Guyana Cu Royalty ($7,162) ($1.93)

One Time Royalty to Surface Owner ($20,000) ($5.39)Net Revenue $5,210,588Operating Costs $/Au-ozMining Cost ($1,105,659) ($298.24)Processing Cost ($1,335,747) ($360.30)Site G&A Cost ($173,821) ($46.89)Total Operating ($2,615,227)$/t-ore ($20.57)Cash Cost ($/Au-oz) ($700)Operating Margin (EBITDA) $2,595,362Initial Capital ($501,192)Total Capital ($820,651)Income Tax ($506,310)Cash Flow Available for Debt Service $1,268,400

Pre-Tax IRR 27.19%Pre-Tax Present Value 0% $1,774,710Pre-Tax Present Value 5% $991,516Pre-Tax Present Value 8% $702,064

After -Tax IRR 23.14% After -Tax Present Value 5% $690,869 After -Tax Present Value 8% $476,171 After -Tax Present Value 10% $367,345

1.15 Conclusions and Recommendations

Mineral Resources

The mineral resource within the material encompassed by the PFS design pit has been sufficiently

sampled and delineated to achieve a confidence classification (Measured and Indicated) which is

adequate for mine planning purposes at the feasibility study level.

Mineral Reserves and Mining Methods

Through the process of pit optimization, fleet estimation, mine design, production scheduling and

economic modeling, the Toroparu open pit operations have been sized and estimated appropriately

at prefeasibility level. SRK has estimated a mining cost of US$1.86/t mined for in-situ and re-handled

material over the LoM. Combined with a ramp-up to 52 Mt/y, the benefit of high production rates,





grade bin schedule and support from two pit operations, SRK is of the opinion the costs estimated

are reasonable at the present time (2013).

Open pit initial mine capital is estimated at US$74.7 million (site preparation and equipment including

contingency) with an addition US$176.5 million in sustaining capital (including contingency), which

includes purchase of new equipment.

Additional mining studies required at a feasibility level of design for the Project include:

Incorporate additional geotechnical design data into the SE Pit design;

Feasibility level pit designs including dewatering structures;

Improved estimates of groundwater in-flow from local structures into the pit;

Assessment of a condemnation drilling program to confirm the locations of the low grade ore

stockpile, primary crusher and waste dumps;

Development of a feasibility level mine production schedule including monthly periods to

start, and completing the LoM schedule in quarterly periods to determine continuous ore

exposure; and

Assessment of an expanded articulated dump truck (ADT) fleet to mine part of the saprolite

waste and ore throughout most of the LoM mine production schedule.

Metallurgy Process and Recovery

The testwork of ACO Composite samples has shown that the overall gold recovery from gravity

separation and gravity tailings leaching is higher than the case of gravity separation, cleaner flotation

and cleaner tailing leaching. However, due to the lower proportion of cyanide soluble copper

reporting to the cleaner tailing leaching stage, the cyanide consumption is lower in the flotation

process than direct gravity tailing leaching. The lower reagent costs and the benefit of the copper as

a by-product credit indicates that flotation plus leaching of the flotation cleaner tailings would be a

better option for this type of material.

Similar to the ACO Composite, overall gold recovery from direct cyanide leaching of LCO composites

gravity tailings is higher than the cyanide leaching of the rougher concentrate. Moreover, the cyanide

consumption is also lower for the direct cyanide leaching of the gravity tailings than leaching of

concentrate, indicating leaching of gravity tailings is more favorable for LCO ore types when

compared to flotation of LCO ore.

Testwork on the ACO and LCO composites has also shown that a significant amount of copper is

loaded onto the carbon following cyanide leaching. To limit the amount of copper loading on the

carbon, the initial cyanide concentration may need to be increased. The higher cyanide

concentration will likely increase the extraction of copper and we recommend further optimization

testing to further define the design parameters. It is also recommended that a Gravity RecoverableGold (GRG) test be conducted to further define the effect of grind size on the recovery of gold.

Flotation testwork conducted during this program focused on evaluating the effect of grind size on

recovery, and additional study will be required to finalize a flotation circuit flowsheet. The results from

the locked cycle test of the ACO Master Composite indicated that there is a build-up of non-sulfide

gangue in the cleaning circuit. It is recommended that further flotation testwork be conducted to

better understand the effect of cleaner circuit flowsheet and use of gangue depressant, CMC, and

other reagents and flotation flowsheets to potentially improve the process.





We recommend additional variability testing with regard to gravity separation, copper flotation and

cyanide leaching on the LCO composite. The testing should identify the gold response in variation to

head grade.

A pilot scale HPGR test, in which 2 tons of material is tested, is recommended to properly size the

HPGR unit, as the relevant testwork to date only provides preliminary sizing information on a scopinglevel. This should be performed in conjunction with variability comminution testing to provide better

selection of ball mills.

Due to the number of thickeners present in the proposed flowsheet, additional thickening tests are

recommended to properly size the thickeners. These tests should evaluate the following:

ACO rougher flotation tailings;

A mixture of the CIP tailings comprised of saprolite and ACO cleaner flotation tailings;

LCO whole ore material; and

A mixture of the CIP tailings comprised of LCO material and ACO cleaner flotation tailings.

In the event that variability gravity/float/and leach testing of the ACO or LCO material indicates analternative process flowsheet, then thickening tests should be expanded to reflect such a change.

This is especially relevant if an LCO cleaner flotation process is considered to better define the

impact of finer particles due to the regrind circuit which will impact settling behavior.

If sufficient sample is produced, thickening and filtration testing of the copper concentrate should be

performed as selection and sizing of concentrate dewatering equipment is currently assumed due to

the absence of any related testwork.

Lastly, additional cyanide destruction (CND) testing should be performed to confirm earlier findings.

In addition, should any of the variability testing of the ACO or LCO material yield a flowsheet

significantly different than what is currently proposed then CND testing should be expanded to

included such changes.

On-Site Infrastructure

Considering that the Project is at prefeasibility level, the following work to progress the Project to the

next level is recommended.

Perform condemnation drilling at the proposed plant site and man camp site;

Perform a geotechnical site investigation that includes test pits, drillholes and seismic

refraction surveys to support foundation design and engineering;

Perform a feasibility study design on the on-site infrastructure;

Confirm water supply source, quality and quantity requirements; and

Confirm wastewater treatment, discharge point, quality and quantity requirements.Tailings Management Area

The geotechnical conditions of the tailings site have been assumed from the geology information

previously described and from the results of the investigation program carried out by KCB (2012) at

the location of the tailings site defined in the PEA. A specific site investigation program for the

tailings site should therefore be carried out for the new location defined in this study during feasibility

engineering of the Project.





Off-Site Infrastructure

Considering the location of the Toroparu Mine, off-site infrastructure will play a key role in the

operation and success of the Project. The presence of an existing roadway from the Mazaruni River

to the mine is advantageous, and upgrades to the roadway will ensure safe, reliable transport of

cargo to and from the mine. Overall operation and maintenance of the roadway is currentlyadministered by the GoG via a contract with ETK and local contractors. After upgrade of the roadway

by Sandspring-ETK as contemplated in this technical report, the GoG will bear the full operating

costs associated with the Toroparu Mine Road.

Construction of the port facility at Pine Tree Landing will create a secure shipping point for mine

equipment, bulk mine supplies and concentrate, and will serve as a base camp for mine supply chain

transport vehicles.

Detailed engineering design will be required to prepare construction documents and to finalize

construction cost estimates for the port facility and roadway upgrades.

Environment, Permits, Licenses and Authorizations

The Project area has been historically impacted by mining activities, logging, and hunting. With only

a few exceptions, species classified as rare, threatened or endangered have not been observed in

the Project area. No indigenous hunting activity or cultural resources were identified within the

proposed mining area. The water quality baseline sampling has not included specific sampling

events to establish a baseline characterization trend with seasonal variability. SRK recommends that

quarterly sampling be re-established to coincide with the variation in the wet and dry seasons. SRK

recommends that the sampling methodology and water construction and development procedures be

further reviewed due to some abnormalities in the water sampling results. It is recommended that

monitoring at the weather station be continued on a monthly basis.

Results of the geochemical testing of the waste rock by KCB showed that the waste rock lithologiesand LGO samples contained very low sulfide-sulfur concentrations, indicating low risk of PAG,

except for the saprolite. The saprolite and transition zone samples contained very low NP, whereas

the waste rock and LGO had NP related to reactive carbonate minerals. The saprolite samples were

classified primarily as acid-generating and PAG, whereas the other waste rock and LGO samples

were classified as non-PAG.

The tailings samples contained low to negligible sulfide-sulfur concentrations and were classified as

non-PAG. The majority of the NP of the tailings was associated with the reactive carbonate minerals

and/or lime added during the metallurgical testing. The saprolite tailings contained little to no reactive

carbonate minerals, and thus the NP present in the saprolite tailings was related to the lime added

during the metallurgical process.

Leachate testing indicated that the waste rock may develop alkaline drainage with the possibility of

elevated concentrations of aluminum, selenium, chromium and, to a lesser extent, copper and

phosphorus. The tailings could develop alkaline drainage with the possibility of elevated

concentrations of aluminum, selenium, chromium, arsenic, cobalt, copper, iron, molybdenum, WAD

cyanide and sulfate. The geochemistry program should be advanced to a more detailed program that

will include predictions of water quality associated with the mining wastes run-off and discharges.

Water quality management strategies are needed for the tailings pond. Further static and kinetic

testing is recommended.





Additional geochemical studies that are recommended by KCB include humidity cell testing and/or

field lysimeters or leach barrels to evaluate alkaline metal leaching rates, and follow-up predictive

water quality modeling for waste rock run-off and seepage, and TMA pond discharge water quality.

For the metallurgical tailings, KCB recommended that laboratory kinetic tests be initiated to assess

the TMA pond and porewater quality. An ESIA was prepared and submitted to the Guyana EPA,

which subsequently issued an environmental permit for mining and processing. A variety of

compliance items are required as part of the environmental permit. The application for and receipt of

the Mining License will be required prior to commencing full-scale operations

Although additional studies are recommended to further develop mining waste management

strategies, there are no known environmental issues that could materially impact Sandspring’s ability

to extract the mineral resources or reserves at the site. Preliminary mitigation strategies have been

developed to reduce environmental impacts to meet regulatory requirements and the specifications

of the environmental permit.

Community Relations and Social Responsib ility

There are no formal or established communities in the immediate vicinity of the site. The Project isnot expected to generate many direct socio-economic impacts. A Social Management Plan has been

proposed to mitigate the socio-cultural impacts identified in the ESIA. It is recommended that

consultation with the community be continued. The Social Management Plan proposed in the ESIA

should be prepared and implemented as required under the environmental permit.

Economic Analysis

A discounted cash flow model was created to evaluate the Toroparu Project assuming the Project is


An income tax rate of 30% has assumed. The resulting cash flow in each year of the Project life was

discounted back to January of Year -3 to determine the estimated discounted cash flow. Using thissame data, the estimated internal rate of return and the undiscounted cash flow were also

determined.

Using a gold price of US$1,400/oz and a copper price of US$3.25/lb, results of the base case

analysis indicate that the Toroparu Project has a potential after-tax internal rate of return of 23% and

a present value of approximately US$690 million, based on a 5% discount rate.

The base case payback period is estimated at 2.6 years, including sustaining capital, from the start

of the production period (from the start of Year 1).

These positive results indicate that the Project should be advanced to a Feasibility Study.



TOROPARU GOLD PROJECT

TOROPARU MINE SITE

04/26/13 ISSUED FOR PRE-FEASIBILITY DEN PC PC

114-311317-DWG-C-101

04/2013

DEN

A

ASSHOWN

LEGEND

EXISTING MAJOR CONTOUR

EXISTING MINOR CONTOUR

PROPOSED MAJOR CONTOU

PROPOSED MINOR CONTOU

EXISTING ROAD

EXISTING RIVER

EXISTING FLOODPLAIN BOU

PROPOSED FLOODPLAIN BO

PROPOSED ROADS

PROPOSED CHANNEL

PROPOSED LEVEE

HAUL ROAD

PROPOSED PAD

PROPOSED POND

METER

125 125 250

1:1

0

DRAWING DESCRIPTION:

PROJECT NAME:

DRAWING NO:

AUTHORIZED BY:

DESIGNED BY: REVIEWED BY:PREPARED BY:

SCALE:

SEALPERMIT STAMPSUB CONSULTANT(S) : CLIENT:

NO. DESCRIPTIONDATE

REVISIONS/ISSUE DRAFTING

PREPAREDREVIEW DESIGN AUTHORIZE

ENGINEERING

NO. DESCRIPTION

REFERENCEDRAWINGS

DATE:

3D MODEL REF No:

NOTE:

MINE PLAN 18 ULTIMATE PIT BOUNDARY, SRK AP



Toroparu Gold Project,

Guyana

Figure 1.6.2

Mine Plan Progress MapsYear 14 End, Ultimate PitSource: SRK, 2013



JAW

CRUSHER

SECONDARY

CONE

CRUSHER

DUMP

TRUCK

P

TCIP TANK

NO.1

TAILING MANAGEMENT AREA

ACID WASH

TANK

ELECTROWINING CELLSKILN

TO CARBON FINES

RECOVERY

PROCESS

WATER TO

DISTRIBUTION

DETOXIFICATION

FRONT END

LOADER

SAPROLITE

SCREEN

FLOTATION

STOCKPILE

LEACH

STOCKPILE

HPGR

SCREEN

ELUTION

COLUMNS

SCREEN

GRAVITY

CONCENTR

BALL MILL

HPGROVER SIZE

BALL MILL

AUTHORIZED BY:

DESIGNED BY:SEALPERMIT STAMPSUB CONSULTANT(S) :

NO. DESCRIPTIONDATE


PREPARE D RE VIE W DE SIGN AU TH ORIZ E

ENGINEERING

NO. DESCRIPTION

REFERENCE DRAWINGS

A 03/26/13 ISSUED FOR INTERNAL REVIEW R.P. R.P.

B 04/02/13 ISSUED FOR CLIENT REVIEW R.P. R.G.

0 05/07/13 ISSUED FOR PRE-FEASIBILITY STUDY R.G. P.C.

SCREENSCREEN

SCREEN

PUMP BOX

TANK

PUMP BOX

CONVEYOR

SCREEN

CONVEYOR

CONVEYOR

CIP TANK

NO.2

CIP TANK

NO.3

CIP TANK

NO.4

CIP TANK

NO.5

TRAIN A & B

CONVEYOR CONVEYOR

HYDROCYCLONESHYDROCYCLONES

CONVEYOR

CONVEYOR

HOPPER

CONVEYOR

REBA

REFINERY

TAILINGS

THICKENERS

1 & 2



TOROPARU GOLD PROJECT

TOROPARU MINE OVERAL

04/26/13 ISSUED FOR PRE-FEASIBILITY DEN PC PC

114-311317-DWG-C-100

04/2013

DEN

A

ASSHOWN

LEGEND

EXISTING MAJOR CONTOUR

EXISTING MINOR CONTOUR

PROPOSED MAJOR CONTOU

PROPOSED MINOR CONTOU

EXISTING ROAD

EXISTING RIVER

EXISTING FLOODPLAIN BOU

PROPOSED FLOODPLAIN BO

PROPOSED ROADS

PROPOSED CHANNEL

PROPOSED LEVEE

HAUL ROAD

PROPOSED PAD

PROPOSED POND

METER

300 300 600

1:3

0

DRAWING DESCRIPTION:

PROJECT NAME:

DRAWING NO:

AUTHORIZED BY:

DESIGNED BY: REVIEWED BY:PREPARED BY:

SCALE:

SEALPERMIT STAMPSUB CONSULTANT(S) : CLIENT:

NO. DESCRIPTIONDATE


PREPAREDREVIEW DESIGN AUTHORIZE

ENGINEERING

NO. DESCRIPTION

REFERENCEDRAWINGS

DATE:

3D MODEL REF No:

NOTE:

1. MINE PLAN 18 ULTIMATE PIT BOUNDARY, SRK APRIL 15, 201

2. TMA FACILITY BOUNDARY, KCB MARCH 18, 2013.



0

200,000

400,000

600,000

800,000

1,000,000

1,200,000

1,400,000

80% 90% 100% 110% 120%

NPV 5% Sensivity ($000s)

Capital Costs

Operating Costs

Revenue


Guyana

Figure 1.14.1

NPV Sensit ivitySource: SRK, 2013





2 Introduction

2.1 Terms of Reference and Purpose of the Report

This report is as a National Instrument 43-101 (NI 43-101) Technical Report on the PrefeasibilityStudy of the Toroparu Gold Project (Toroparu or Project) prepared for Sandspring Resources Ltd.

(Sandspring) by SRK Consulting (U.S.), Inc. (SRK), Tetra Tech (Tt), Klohn Crippen Berger (KCB),

Knight Piésold (KP), and FMG, Inc. (FMG), (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 Sandspring subject to the terms and conditions of its contract with SRK and

relevant securities legislation. The contract permits Sandspring 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 thisdisclosure remains with Sandspring. 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).

2.2 Qualifications of Consultants

The Consultants preparing this technical report are specialists in the fields of geology, exploration,

mineral resource and mineral reserve estimation and classification, underground 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 Sandspring. The Consultants are not insiders, associates, or affiliates of

Sandspring. 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 Sandspring 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:

Alex Fisher, B.Sc. Geological Engineering, P.E. is the QP responsible for the off-site

infrastructure (road and port) Sections 18.3, 21.2.3, 21.3.4, and portions of Sections 1, 25

and 26 summarized therefrom, of this Technical Report. Mr. Fisher conducted an

investigation of the road and port sites on December 2 through 6, 2011;





2.3 Sources of Information

This report is based, in part, on internal company technical reports and maps, published government

reports, company letters and memoranda, and public information as listed in the "References"

section of this report. Several sections from reports authored by other consultants have been directly

quoted in this report, and are so indicated in the appropriate sections. SRK has not conducteddetailed land status evaluations, and has relied upon previous qualified reports, public documents

and statements by Sandspring regarding property status and legal title to the Project.

2.4 Units of Measure

Unless otherwise stated all units used in this report are metric. Gold and silver assay values are

reported in grams per tonne (g/t) unless ounces per ton (oz/ton) are specifically stated. Base metal

assay values are given in percent (%) or in parts per million (ppm). The US$ is used throughout this

report unless otherwise noted.

2.5 Glossary and Abbreviated Terms

2.5.1 Mineral Resources

The mineral resources and mineral reserves have been classified according to the “CIM Standards

on Mineral Resources and Reserves: Definitions and Guidelines” dated November 27, 2010 (CIM).

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, geologicalcharacteristics 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 mineplanning 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.

2.5.2 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.

2.5.3 Definit ion of Terms

The following general mining terms may be used in this report.

Table 2.5.3.1: Definiti on 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 largerdistance.

Concentrate A metal-rich product resulting from a mineral enrichment process such as gravityconcentration or flotation, in which most of the desired mineral has been separatedfrom the waste material in the ore.

Crushing Initial process of reducing ore particle size to render it more amenable for furtherprocessing.

Cut-off Grade The grade of mineralized rock, which determines as to whether or not it is economicto 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 centrifugalforces 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 minimizesthe estimation error.

Level Horizontal tunnel the primary purpose is the transportation of personnel and





Term Definitionmaterials.

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 groundand subjected to physical or chemical treatment to extract the valuable metals to aconcentrate 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.Pillar Rock left behind to help support the excavations in an underground mine.

QualifyingTransaction

A transaction where a CPC acquires significant assets, other than cash, by way ofpurchase, amalgamation, merger or arrangement with another company or by othermeans.

RoM Run-of-Mine.

Saprolite

A chemically weathered rock, mostly soft or friable and commonly retaining thestructure of the parent rock since it is not transported. Saprolites containpredominantly quartz and a high percentage of kaolinite with other clay minerals

which are formed by chemical decomposition of primary minerals, mainly feldspars.Sedimentary Pertaining to rocks formed by the accumulation of sediments, formed by the erosionof other rocks.

Shaft An opening cut downwards from the surface for transporting personnel, equipment,supplies, ore and waste.

Sill A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by theinjection of magma into planar zones of weakness.

Smelting A high temperature pyrometallurgical operation conducted in a furnace, in which thevaluable metal is collected to a molten matte or doré phase and separated from thegangue 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.

Stripping Ratio

The ratio of tonnes of waste rock divided by the tonnes of mineralization destined for

the processing plant.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).





Abbreviation Defini ti on

ft square foot (feet)

ft cubic foot (feet)

g gram

g Au/t grams of gold per tonne

g/L gram per liter

g/t grams per tonne

Ga billion years

gal gallon

GGMC Guyana Geology and Mines Commission.

g-mol gram-mole

GoldHeart GoldHeart Investment Holdings Ltd.

gpm gallons per minute

gpt grams per tonne

ha hectare

HDPE Height Density Polyethylene

HFO Heavy Fuel Oil

hp horsepower

HTW horizontal true width

ICP induced couple plasma

ID2 inverse-distance squared

ID3 inverse-distance cubed

IFC International Finance Corporation

ILS Intermediate Leach Solution

Inorg-CaNP Inorganic Carbon Neutralization Potential

kA kiloampere

kg kilogram

km kilometer

km square kilometer

koz thousand troy ounce

kt thousand tonnes

kt/d thousand tonnes per daykt/y thousand tonnes per year

kV kilovolt

kW kilowatt

kWh kilowatt-hour

kWh/t kilowatt-hour per metric tonne

L liter

L/sec liters per second

L/sec/m liters per second per meter

lb pound

LCO Low Copper Ore

LGO Low Grade Ore

LHD Long-Haul Dump truck

LLDDP Linear Low Density Polyethylene PlasticLOC Low Copper Ore

LOI Loss On Ignition

LoM Life-of-Mine

m meter

M million

m.y. million years

m square meter

m cubic meter

Ma million years

masl meters above sea level





Abbreviation Defini ti on

mg/L milligrams per liter

mL meter level

mm millimeter /millimeters

mm square millimeter

mm cubic millimeter

MME Mine & Mill Engineering

Moz million troy ounces

MPs Mining Permits

Mt million tonnes

MTW measured true width

MW million watts

N north

NAG Net Acid Generation

NE northeast

NGO non-governmental organization

NI National Instrument

NI 43-101 Canadian National Instrument 43-101

NP Neutralization Potential

N-PAG Not Potentially Acid Generating

NPR Net Potential Radio

NSR net smelter return

NTS National Topographic System

NW northwest

OSC Ontario Securities Commission

oz troy ounce

oz/t ounces per short tonne

P&E P&E Mining Consultants Inc.

PAG Potentially Acid Generating

PEA Preliminary Economic Assessment

PFS Prefeasibility Study

PL Prospecting LicensePLC Programmable Logic Controller

PLS Pregnant Leach Solution

PLs Prospecting Licences

PMF probable maximum flood

ppb parts per billion

ppm parts per million

PPMSs Medium Scale Prospecting Permits

QA/QC Quality Assurance/Quality Control

RC rotary circulation drilling

RoM Run-of-Mine

RQD Rock Quality Description

S south

Sandspring Sandspring Resources Ltd.SE southeast

sec second

SEC U.S. Securities & Exchange Commission

SEDAR System for Electronic Document Analysis and Retrieval

SG specific gravity

SPT standard penetration testing

st short ton (2,000 pounds)

SW southwest

t metric tonne (2,204.6 pounds)

t/d tonnes per day





3 Reliance on Other ExpertsThe Consultant’s opinion contained herein is based on information provided to the Consultants by

Sandspring throughout the course of the investigations. SRK has relied upon the work of other

consultants in the Project areas in support of this Technical Report. The sources of informationinclude data and reports supplied by Sandspring 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. This

report includes technical information, which required subsequent calculations to derive subtotals,

totals and weighted averages. Such calculations inherently involve a degree of rounding and

consequently introduce a margin of error. Where these occur, the Consultants do not consider them

to be material.

The authors wish to emphasize that they are Qualified Persons in respect of the areas in this report

identified in their certificates of Qualified Persons submitted with this technical report.

Although copies of the licenses, permits and work contracts were reviewed, an independent

verification of land title and tenure was not performed. The Consultants have not verified the legality

of any underlying agreement(s) that may exist concerning the licenses or other agreement(s)

between third parties. The authors have relied, and believe that they have a reasonable basis to rely,

upon Sandspring who has contributed portions of the tenure, legal, environmental, marketing and

taxation information stated in this report.

A draft copy of the report has been reviewed for factual errors by Sandspring. Any changes made as

a result of these reviews did not involve any alteration to the conclusions made. Hence, the

statements and opinions expressed in this document are given in good faith and in the belief that

such statements and opinions are not false and misleading at the date of this report.

All geological data were collected and compiled, and geological interpretations were performed by

the Sandspring exploration team under the management of L.W. Claessens, P. Geo, VP-Exploration,

and Pascal van Osta, P.Geo, Exploration Manager. In the appropriate sections covering the

geological data SRK has expressed its opinions on the methods of collection and data quality. SRK

considers the data to be adequate to support the resource estimation.

SRK is reliant upon Mr. Scott Issel, Chief Financial Officer of Sandspring as an expert in regard to

the projected gold and copper royalty rates, depreciation method and the Guyana corporate tax rate

used in the cashflow model presented in Section 22 of this Technical Report. SRK’s assessment of

the data provided for these aspects was that it is reasonable for use in the present PFS, on which

this report is based.

Mr. Greg Barnes, Executive Vice President, provided SRK and Tt with unit costs including projected

labor rates including payroll burdens by position for Guyana and expatriate personnel; travel costs,

mine camp unit operating costs, diesel fuel and heavy fuel oil costs, and domestic freight rates. SRK

utilized this unit cost information to develop the mining operating costs, Tt to develop the processing

operating costs, and Sandspring to develop the general and administration costs presented in

Section 21, and used in the cashflow model presented in Section 22 of this report. FMG provided





SRK with the estimated capital costs for port development, access roads upgrading, and bridge

construction referenced in Sections 21 and 22 of this report.

Mr. Barnes was also relied upon as an expert for the preliminary commercial terms, such as

estimated concentrate transport and insurance costs, and smelter payables, deductions, treatment

and refining costs. SRK has reviewed the projected commercial terms provided by Sandspring andthe gross gold revenue and copper credit estimates and have determined them to be reasonable and

acceptable for use in the present PFS, on which this report is based.

Sandspring has been present in Guyana for over 12 years and has obtained valuable knowledge of

the local and regional markets and thus has contributed to certain inputs.





4 Property Description and Location

4.1 Property Descript ion and Tenure

The Property descriptions of land held by ETK have been updated based on information provided byMr. Rich A. Munson, CEO of Sandspring. The descriptions contained herein therefore supersede

those in any previous reports on the Property.

ETK, a private company in Guyana, has rights to small scale claims, Medium Scale Prospecting

Permits (PPMSs), Mining Permits (MPs), and Prospecting Licenses (PLs) within the Property. All of

ETK’s issued and outstanding stock is owned by GoldHeart. All of the issued and outstanding stock

of GoldHeart is owned by Sandspring.

Sandspring acquired its interest in GoldHeart and ETK pursuant to the terms of a share purchase

agreement dated May 11, 2009, as amended.

The Property is comprised of seven small scale claims, 167 contiguous PPMS’s and 13 MP’s that

together cover an area of 184,693.8 acres or 74,742.9 hectares, and five contiguous PL’s that cover

an area of 57,997 acres or 23,471 hectares. A list of the land tenure is given in Tables 4.1.1 through

4.1.3.

ETK has four positions of claim ownership in the Upper Puruni Area. The PL’s are wholly owned by

ETK. The seven small scale claims are located within the exterior boundaries of A-4/MP/011. Pam 1,

Pam 2, Pam 3, Joy 1, Joy 2, Joy 3 and Joy 4 located in or near the Puruni River within the exterior

boundaries of Mining Permit A-4/MP/011, described under MP’s below, and as reflected on the

records of the Guyana Geology and Mines Commission (GGMC) following verification surveys and

reviews conducted in October, November and December 2007.

The MP’s and PPMS’s identified in Table 4.1.1 and Table 4.1.2, respectively, by the prefix “A” refer

to permits which are owned by Mr. Alfro Alphonso (Alphonso), and are controlled by ETK under a

joint venture agreement (the Alphonso Joint Venture). The “Middle Ground” claims, directly north of

the Toroparu pit and south of the PL’s are held pursuant to a joint venture with Mr. Wallace Daniels

(Daniels) and are listed in Table 4.1.1 and Table 4.1.2, respectively, with the prefix “D”. The Godette

MP’s south of the PL’s and east of the Toroparu open pit are held under a joint venture agreement

with the Godette family and are listed on Table 4.1.1 and Table 4.1.2, respectively, with the prefix

“G”.

Ten parcels of land are subject to applications for the issuance of PPMS’s filed by Mr. Wallace

Daniels, a local Guyana resident. Ownership of PPMS’s covering these ten parcels of land is the

subject of a dispute between Mr. Daniels and a third party. Sandspring does not consider the

disputed parcels as having any current material value and the parcels do not form any part of the

resource estimate for the Toroparu Project and are not included in this Technical Report.

On November 9, 2011, the Company signed a mineral agreement (the Mineral Agreement) with the

Government of Guyana, which details all fiscal, property, import-export procedures, taxation

provisions and other related conditions for the continued exploration, mine development and

operation of the open pit mine at the Toroparu Project. The Mineral Agreement implements a two-

tiered gold royalty structure of 5% of gold sales at gold prices up to US$1,000/oz. and 8% of gold

sales at gold prices above US$1,000/oz., and a royalty of 1.5% on sales of copper and other





GS8 NumberPPMSNumber

Area(Acres)

LocationMap

NumberRenewal Date

A-184/001/0395/99 0265/2001 948 Ikuk River 44 March 11, 2014 A-184/002/0396/99 0266/2001 1137 Ikuk River 59 March 11, 2014 A-184/003/0397/99 0267/2001 987 Ikuk River 76 March 11, 2014 A-184/004/0398/99 0268/2001 1200 Ikuk River 78 March 11, 2014 A-184/005/0399/99 0269/2001 1200 Ikuk River 79 March 11, 2014

A-184/006/0400/99 0270/2001 1020 Ikuk River 98 March 11, 2014 A-184/007/0401/99 0271/2001 927 Ikuk River 97 March 11, 2014 A-184/008/0402/99 0272/2001 869 Ikuk River 77 March 12, 2014 A-184/009/99 0579/2002 804 Upper Puruni 149 August 15, 2013 A-184/010/99 0580/2002 804 Upper Puruni 148 August 15, 2013 A-184/011/99 0581/2002 780 Upper Puruni 147 August 15, 2013 A-184/012/99 0582/2002 1058 Upper Puruni 146 August 15, 2013 A-184/013/99 0583/2002 1170 Upper Puruni 150 August 15, 2013 A-185/001/99 0577/2002 795 Upper Puruni 5 August 15, 2013 A-185/002/99 0578/2002 1143 Upper Puruni 4 August 14, 2013 A-185/003/0411/99 0227/2001 795 Puruni River 3 March 07, 2014 A-185/004/0412/99 338/2001 1043 Puruni River 19 May 17, 2014 A-185/005/0413/99 0228/2001 1200 Puruni River 15 March 06, 2014 A-185/006/0414/99 0229/2001 1200 Puruni River 20 March 06, 2014 A-185/007/0415/99 0330/2001 1200 Upper Puruni 28 March 06, 2014 A-185/008/0416/99 0331/2001 1190 Upper Puruni 29 March 06, 2014 A-185/009/0417/99 0424/2001 1190 Upper Puruni 30 May 27, 2014 A-185/010/0418/99 0425/2001 1036 Upper Puruni 43 May 28, 2014 A-185/011/0419/99 0426/2001 1071 Upper Puruni 42 May 28, 2014 A-185/012/0420/99 0332/2001 1071 Upper Puruni 41 March 06, 2014 A-185/013/0421/99 0333/2001 1087 Upper Puruni 40 March 07, 2014 A-185/014/0422/99 0334/2001 1136 Upper Puruni 39 March 06, 2014 A-185/015/0423/99 0335/2001 1200 Ikuk River 38 March 06, 2014 A-185/016/0424/99 0336/2001 637 Ikuk River 58 March 06, 2014 A-185/017/0425/99 0337/2001 658 Ikuk River 57 March 06, 2014 A-185/018/0426/99 0338/2001 658 Upper Puruni 56 March 06, 2014 A-185/019/0427/99 0339/2001 607 Upper Puruni 55 March 06, 2014 A-185/020/0428/99 0340/2001 679 Upper Puruni 54 March 06, 2014 A-185/021/0429/99 0341/2001 637 Upper Puruni 53 March 06, 2014 A-185/022/0430/99 0342/2001 1125 Ikuk River 75 March 06, 2014 A-185/023/0431/99 0343/2001 1125 Ikuk River 74 March 07, 2014

A-185/024/0432/99 0344/2001 1125 Ikuk River 73 March 07, 2014 A-185/025/0433/99 0345/2001 1200 Ikuk River 72 March 08, 2014 A-185/026/0426/99 0346/2001 700 Putaring 71 March 08, 2014 A-185/027/99 0697/2002 675 Upper Puruni 70 October 16, 2013 A-185/028/0436/99 0347/2001 1150 Putaring 95 March 07, 2014 A-185/029/0437/99 0348/2001 1139 Putaring 94 March 07, 2014 A-185/030/0438/99 0349/2001 1035 Putaring 93 March 08, 2014 A-185/031/0439/99 0350/2001 1081 Putaring 92 March 08, 2014 A-185/032/0440/99 0351/2001 1200 Putaring 2 March 06, 2014 A-185/033-0441/99 0352/2001 1200 Putaring 1 March 06, 2014 A-185/034/0442/99 0353/2001 1104 Putaring 9 March 08, 2014 A-185/035/0443/99 0354/2001 1066 Puruni River 10 March 08, 2014 A-185/036/0444/99 0355/2001 1066 Puruni River 14 March 08, 2014 A-185/037/0445/99 0356/2001 1104 Tamakay 13 March 08, 2014 A-185/038/0446/99 0357/2001 1115 Puruni River 17 March 07, 2014 A-185/039/0447/99 0358/2001 1114 Tamakay 18 March 08, 2014

A-185/040/0448/99 0359/2001 1000 Tamakay 26 March 08, 2014 A-185/041/0449/99 0360/2001 1080 Tamakay 27 March 08, 2014 A-199/000/2000 620/2001 1016 Puruni River 64 September 19, 2013 A-199/001/2000 621/2001 1016 Puruni River 81 September 19, 2013 A-199/002/2000 622/2001 1200 Tamakay 82 September 19, 2013 A-199/003/2000 623/2001 1016 Puruni River 83 September 19, 2013 A-199/004/2000 624/2001 1016 Puruni River 85 September 19, 2013 A-199/005/2000 625/2001 1016 Puruni River 87 September 19, 2013 A-199/006/2000 626/2001 1016 Puruni River 115 September 19, 2013 A-199/007/2000 627/2001 1014 Puruni River 117 September 19, 2013 A-199/008/2000 628/2001 1085 Puruni River 118 September 19, 2013 A-199/009/2000 629/2001 1119 Puruni River 114 September 19, 2013 A-199/010/2000 630/2001 1125 Puruni River 103 September 19, 2013





GS8 NumberPPMSNumber

Area(Acres)

LocationMap

NumberRenewal Date

A-199/011/2000 631/2001 1102 Puruni River 102 September 19, 2013 A-199/012/2000 632/2001 1102 Puruni River 111 September 19, 2013 A-199/013/2000 633/2001 1076 Puruni River 112 September 19, 2013 A-199/014/2000 634/2001 1102 Puruni River 113 September 20, 2013 A-199/015/2000 643/2002 1148 Tamakay 119 October 15, 2013

A-199/016/2000 635/2001 725 Puruni River 138 September 20, 2013 A-199/017/2000 636/2001 910 Puruni River 137 September 20, 2013 A-199/018/2000 637/2001 1029 Puruni River 136 September 20, 2013 A-199/021/2000 639/2001 1011 Puruni River 32 September 20, 2013 A-199/022/2000 640/2001 995 Puruni River 33 September 20, 2013 A-199/023/2000 641/2001 965 Puruni River 34 September 20, 2013 A-199/024/2000 642/2001 958 Puruni River 48 September 20, 2013 A-199/025/2000 643/2001 1024 Puruni River 60 September 20, 2013 A-199/026/2000 644/2001 940 Puruni River 37 September 20, 2013 A-199/032/2000 649/2001 1024 Puruni River 61 September 20, 2013 A-199/033/2000 0644/2002 998 Tamakay 45 October 07, 2013 A-199/034/2000 0645/2002 998 Tamakay 47 October 07, 2013 A-199/035/2000 0646/2002 998 Tamakay 46 October 07, 2013 A-199/036/2000 0647/2002 1024 Tamakay 62 October 07, 2013 A-199/037/00 0648/2002 983 Upper Puruni 63 October 07, 2013 A-199/038/00 0649/2002 1140 Upper Puruni 49 October 08, 2013 A-199/039/00 0686/2002 912 Upper Puruni 50 October 08, 2013 A-199/040/00 0687/2002 1072 Upper Puruni 67 October 08, 2013 A-199/041/00 0688/2002 1180 Upper Puruni 89 October 08, 2013 A-199/042/00 0689/2002 963 Upper Puruni 116 October 08, 2013 A-199/043/00 0690/2002 1123 Upper Puruni 88 October 08, 2013 A-199/044/00 0691/2002 1098 Upper Puruni 86 October 08, 2013 A-199/045/00 0692/2002 1098 Upper Puruni 84 October 08, 2013 A-199/046/00 0693/2002 1123 Upper Puruni 66 October 08, 2013 A-199/047/00 0694/2002 1123 Upper Puruni 65 October 08, 2013 A-218/001/2001 0678/2002 585 Tamakay 163 October 15, 2013 A-218/002/2001 0594/2002 693 Tamakay 144 August 15, 2013 A-225/000/2001 0679/2002 1147 Tamakay 151 September 20, 2013 A-225/001/2001 0680/2002 747 Tamakay 152 September 25, 2013 A-225/002/2001 0681/2002 878 Tamakay 153 September 25, 2013 A-225/003/2001 0682/2002 484 Tamakay 154 September 26, 2013

A-225/004/2001 0683/2002 1150 Tamakay 155 September 25, 2013 A-225/005/2001 0684/2002 1140 Tamakay 156 September 25, 2013 A-225/006/2001 0475/2002 1140 Tamakay 157 July 07, 2013 A-302/001 0672/2003 1120 Puruni River 69 November 05, 2013 A-302/002 0671/2003 1120 Puruni River 140 November 05, 2013D-166/000/2004 946/04 1200 Ororiparu 167 December 07, 2013D-166/001/2004 947/04 1200 Ororiparu 170 December 07, 2013D-166/002/2004 948/04 1200 Ororiparu 171 December 07, 2013D-166/003/2004 949/04 1200 Ororiparu 172 December 07, 2013D-166/004/2004 950/04 1195 Ororiparu 168 December 07, 2013D-166/005/2004 951/04 1200 Ororiparu 169 December 07, 2013D-166/006/2004 952/04 1200 Ororiparu 173 December 07, 2013D-166/007/2004 953/04 929 Ororiparu 174 December 07, 2013D-166/008/2004 954/04 1196 Ororiparu 175 December 07, 2013D-166/010/2004 950/04 1195 Ororiparu 176 January 16, 2014D-166/011/2004 021/06 1052 Ororiparu 182 January 16, 2014

D-166/013/2004 022/06 444 Ororiparu 177 January 16, 2014D-166/015/2004 023/06 430 Ororiparu 178 January 16, 2014D-166/017/2004 024/06 445 Ororiparu 179 January 16, 2014D-166/018/2004 025/06 1052 Ororiparu 180 January 16, 2014D-166/019/2004 026/06 1052 Ororiparu 181 January 16, 2014D-181/000/2005 018/06 927 Puruni Head 165 January 16, 2014D-181/001/2005 019/06 1014 Puruni Head 166 January 16, 2014D-181/002/2005 020/06 1093 Puruni Head 164 January 16, 2014D-184/000/2005 251/06 761 Toroparu 21 May 01, 2014

Source: Sandspring, 2012





Table 4.1.2: Land Tenure – Mining Permits

GS8 Number MP Number Area

(Acres)Locati on Map Number Renewal Date

A-4/MP/000// A-4/MP/010// 1145 Mazuruni 16 April 28, 2014 A-4/MP/001// A-4/MP/011// 603 Mazuruni 24 April 28, 2014 A-4/MP/002// A-4/MP/012// 858 Mazuruni 25 April 28, 2014 A-4/MP/003// A-4/MP/013// 1098 Mazuruni 23 April 28, 2014 A-4/MP/004// A-4/MP/014// 992 Mazuruni 6 April 28, 2014 A-4/MP/005// A-4/MP/015// 1145 Mazuruni 12 April 28, 2014 A-4/MP/006// A-4/MP/016// 893 Mazuruni 7 April 28, 2014 A-4/MP/007// A-4/MP/007// 1123 Mazuruni 8 April 28, 2014 A-4/MP/008// A-4/MP/008// 1117 Mazuruni 11 April 28, 2014 A-4/MP/009// A-4/MP/009// 1200 Mazuruni 22 April 28, 2014G-6/MP/000 09/2003 960 Toroparu 195 April 9, 2014G-6/MP/001 08/2003 1120 Toroparu 194 April 9, 2014G-6/MP/002 07/2003 996 Toroparu 193 April 9, 2014


Table 4.1.3: Land Tenure – Prospect ing LicensesPL Number Area (acres) Renewal DatePL 01/2002 GS 14 E-10 11,960 September 18, 2013PL 02/2002 GS 14 E-09 11,960 September 18, 2013PL 03/2002 GS 14 E-11 11,986 September 18, 2013PL 04/2002 GS 14 E-12 10,155 September 18, 2013PL 05/2002 GS 14 E-13 11,936 September 18, 2013


Mineral claims are subject to annual rentals by the dates as indicated in Tables 4.1.1 through 4.1.3.

Sandspring acknowledges that the rentals are paid in full for all claims as of the effective date of this

report. ETK has been, and will continue to remain, responsible for the payment of rentals. Paymentson the claims are made each year prior to the renewal date of each claim. The ten units designated

as A-4/MP/007, A-4/MP/008, A-4/MP/009, A-4/MP/010, A-4/MP/011, A-4/MP/012, A-4/MP/013, A-

4/MP/014, A-4/MP/015, and A-4/MP/016, refer to MP’s that were converted from PPMS’s.

The rental rates for each of the MPs are the sum of US$1.00 per acre per annum. The rent for each

of the thirteen MP’s is fully paid.

Mineral tenures in Guyana allow for four scales of operation. These include small scale claim

licenses of 460 m x 245 m or a river claim consisting of one mile of a navigable river. PPMS’s and

MP’s cover between 150 to 1,200 acres each and are restricted to ownership by Guyanese.

However, foreigners may enter into joint venture arrangements whereby the two parties jointly

develop the property. PL’s covering between 500 and 12,800 acres are granted to foreigncompanies. Large areas for geological surveys are granted as Permission for Geological and

Geophysical Surveys with the objective of applying for PL’s over favorable ground.

Rental rates for PL’s are US$0.50 per acre for the first year; US$0.60 per acre for the second year,

and US$1.00 per acre for the third year. An application fee of US$100 and a Work Performance

Bond, equivalent to 10% of the approved budget for the respective year, is also payable. The term

for PL’s is three years with two rights of renewal for one year each. After renewing the PL’s twice,

ETK was given permission to continue renewing on an annual basis. ETK has since requested a





To the best of the writers’ knowledge there are no historical environmental liabilities on the Property.

4.2 Location

The Property is located in the Upper Puruni River Area of northwestern Guyana. The geographic

location of mining operations on the Property, are located at 06° 27’ North Latitude and 60° 03’ WestLongitude, corresponding to UTM co-ordinates of 714450 N and 826200 E. The Property location is

shown on Figure 1.1.1 and the claims area is shown on Figure 4.2.1.

Figure 4.2.2 shows in detail the projected deposit and saprolite Pit outline in relation to the

interpreted IP data.




Guyana

Figure 4.2.1

Upper Puruni Property Claim MapSource: Sandspring Resources, 2012




Guyana

Figure 4.2.2

Detail of Toro paru DepositSource: Sandspring Resources, 2010





5.4 Infrastructure Availability and Sources

5.4.1 Proximity to Population Center

With the exception of some small temporary mining camps along the access road to the Project site,

the closest village is at a distance of 230 km away. Georgetown is 385 km away by road.

5.4.2 Power

There is no nearby electricity grid. Permanent power will not be available at site prior to the

completion of construction. Construction will rely wholly on power from temporary thermal power

generators. Permanent power will be generated on site by thermal power generators.

5.4.3 Water

Water for drilling is readily available throughout the year from creeks and from rainfall run-off.

5.4.4 Mining Personnel

Laborers with a variety of experience in heavy equipment operation are available in Georgetown and

from villages situated along the nearby rivers.

5.4.5 Potential Tailings Storage Areas

KCB investigated several potential tailings facility sites, one of which was chosen for the PFS. The

selected site is located approximately 8 km to the northeast of the main Project site.

5.4.6 Potential Waste Disposal Areas

The PFS design identified appropriate areas for future waste rock disposal. The waste dumps will be

located in areas that that will not be impacted by potential future mining operations. Waste rock

produced from the Toroparu mining operations (both Toroparu and South-East pits) will be placed on

existing terrain in two designated areas. The East Dump will be located between the east of the final

Toroparu Pit and north of the South-East Pit, and the North Dump is to the north of the Toroparu Pit.

The South East Backfill Dump will be within the mined out South-East Pit area.

5.4.7 Potential Processing Plant Sites

Tt investigated several potential processing plant sites, one of which was chosen for the PFS. The

selected site is located nearby to the northeast of the planned Toroparu Pit.

5.5 Physiography

The topography is flat to gently undulating to hilly with an elevation range in the Project area of

approximately 90 to 105 masl in elevation for the mine and plant areas that is occasionally

interspersed with steep hills of meta-basic rock (up to 200 masl southwest of the mine area),

whereas the metasediments represent flatter topographies. The Toroparu Project pit is adjacent to a

very gentle valley and the area surrounding the pit has had small berms constructed to contain the

tailings from past mining operations.





6 HistoryThe following is a summary of the activities history of the Toroparu area.

Late 1880's to 1950

Historic exploitation of alluvial gold and diamonds in the Toroparu area dates back to about 1887.

Conolly (1926) described alluvial diamond operations up to about 1914, to the northwest of the

Toroparu area. Grantham (1934) described gold and diamond workings in the Majuba Hill and

Wynamu areas. The Wynamu River lies adjacent to the Toroparu pit and is labeled as “Toroparu

River” on some older maps.

Pollard and Hamilton created a geological map of the area in 1950 on which the locations of gold

and diamond workings were noted (Pollard, 1950).

6.1 Alluvial and Saproli te Exploration and Mining - 1997 to 2006

During the period of 1997 to 2006 the exploration and mining on the Toroparu property wasconducted as alluvial placer mining operations that in part mined into saprolite in the current pit area.

Exploration during this time was focused on saprolite bearing gold mineralization.

1997

Alphonso commenced alluvial mining at Toroparu in 1997; mining old placer tailings and river

alluvium by washing material into a pit with high pressure water jets and pumping the slurry up to a

sluice box. By 1999, much of the alluvial material was exhausted and work proceeded deeper into

the underlying saprolite and laterally to the west into the saprolite of the hill slope so that the surficial

alluvial area was gradually developed into a 15 to 20 m deep pit (the Toroparu open pit). The

Alphonso operations continued until 2001.

1999

ETK began auger drill sampling to the east and west of the pit area and also evaluated the possibility

of re-working the tailings. Reports by Hopkinson (1999), Uzunlar (2000) and Shaeffer (2000, 2001,

and 2003) summarize the available assay data.

2000

The Guyana Geology and Mines Commission (GGMC) carried out regional mapping and

geochemical stream sampling (Heesterman, et al., 2001) that showed an anomalous gold and

copper area in the immediate Toroparu area.

ETK entered into an exploration joint venture with Alphonso and commenced rehabilitation and

upgrading a 240 km access road into the Property to facilitate the transport of mining equipment and

supplies to the mine site.

2001

Alphonso ceased mining operations in 2001 in the Toroparu open pit. A total of 15 “land dredges”

were employed at the peak of the Alphonso mining activity in the Toroparu open pit area. It has been

estimated that 60,000 oz of gold may have been produced historically over a 70-year period from the

Toroparu area by these alluvial washing methods.





2001-2003

ETK carried out further auger drill sampling in 2001 and 2003 to the east and west of the open pit

area. This work reportedly identified an estimated 1.4 Mt of historic auriferous tailings located

southeast of the main pit area.

2003-2004

Heesterman carried out drainage geochemical sampling for ETK in the PL blocks, located north of

the Toroparu pit area and on lands granted to ETK in 2002. Further geochemical sampling was

performed around the pit area and results indicated that gold mineralization could extend at least

6 km to the northwest and 1 km to the southeast of the Toroparu open pit

2004

ETK commissioned a gravity circuit to test-mine the gold-bearing placer tailings and saprolite, and

also conducted exploration for additional gold sources defined in the GGMC regional geochemical

and prospecting survey of the Upper Puruni area.

From December 2004 to April 2007 ETK conducted intermittent, seasonal test-mining from saprolite,

in the Toroparu pit using a combination of hydraulic sluicing and a gravity circuit with screens, ball

mill, Falcon centrifugal concentrators and shaker tables.

2005

In 2005 and 2006, two phases of trench-channel sampling were completed by Meixner and Wesa to

investigate the gold mineralization in the saprolites of the pit area and to determine the suitability for

conducting further exploration. A zone of gold mineralization, over an area of about 180 m x 100 m,

was identified in the saprolitic rock of the pit area with average grades in the general range of 0.5 to

1.5 g Au/t. This zone was open in all directions.

6.2 Bedrock Exploration - 2006 to 2009

ETK initiated a bedrock exploration and drilling program in 2006 which has culminated in the initial

knowledge of the saprolite and bedrock mineralization at Toroparu. During this time, local alluvial

placer mining continued.

2006-2007

TerraQuest conducted a 5 km x 4.5 km high resolution Tri-sensor Magnetic and Radiometric

Airborne Survey around the Toroparu Pit area in October 2006 on behalf of ETK. The pit area was

found to lie within a magnetically low area just to the south of a large magnetic high area of unknown

provenance. The survey outlined a number of magnetic and radiometric anomalies in the areas

adjacent to the current Toroparu deposit

ETK initiated the Phase 1 drill program in December 2006 as recommended by Meixner and Wesa

(Meixner and Wesa, 2006). Phase 1 included the drilling of 13 NQ core drillholes (3,480 m) under

and around the Toroparu open pit; the program was completed by March 2007. Phase II drilling of an

additional 10 NQ core drillholes (3,748 m) was completed in August 2007. Phase I and II drilling

defined a mineralized block of 600 m x 300 m x 300 m around the Toroparu Pit area (Meixner, 2008).





2008

The ETK Phase III drill program, consisting of 6 NQ cored drillholes (2,590 m), was undertaken from

April to May of 2008. A total of 30 drillholes (TPD 001-030) comprising 10,218 m defined a zone of

mineralization of 650 m x 350 m x 425 m; that was open in all directions. Twenty seven holes totaling

9,492 m formed the basis for the initial mineral resource estimate published in P&E’s TechnicalReport No. 153, effective October 26, 2008, titled “Technical Report and Resource Estimate on the

Toroparu Gold-Copper Prospect, Upper Puruni River, Guyana” (P&E, 2009), as stated in Table

6.2.1.

Table 6.2.1: Toroparu 2008 Mineral Resources

Source: P&E, 2009

ETK carried out additional auger drill sampling to the northwest of the pit area over a 2 km x 3 km

area, using a mechanized auger. Nine north-easterly lines of auger samples, spaced 500 m apart,

were sampled to 5 m depths at approximately 50 m sample intervals. This survey tested the

saprolitic rocks beneath the alluvial cover for gold and copper in an area of historic placer gold

workings that lies to the northwest of the Toroparu open pit area.

An Airborne Geophysical Project was completed in the fourth quarter of 2008 by Allan Spector and

Associates Ltd., consisting of a fixed wing magnetometer and spectrometer survey, totaling

12,400 km of data along 100 m and 200 m spaced north-south oriented flight lines, and covering the

entire ETK Upper Puruni concession surface (+/- 1000 km²).

2009

The ETK Phase IV drilling program conducted between August and December 2009 comprised 21

core drillholes (10,102 m). Thirteen holes were drilled over the Toroparu open pit area with depthsupwards of 500 m and others were drilled as off-trend exploratory holes, as recommended by P&E

(P&E, 2009).

Approximately 2,500 geochemistry saprolite samples were collected using hand and power augers

during 2009, to depths from 1 to 15 m. The soil grids were oriented perpendicular to regional

structures, extending approximately 4.5 km to the WNW from the Toroparu resource area. Assay

results show several areas of gold enrichment along trend to the NW with the highest assay value

equal to 9.94 g Au/t.





SRK has not reviewed the 2008 or 2010 resource estimates, as they are not current and should not

be relied upon. The historical resource estimates are provided here for an understanding of the

progression of resource estimation at the Toroparu Project. Current Mineral Resources are stated in

Section 14 of this report.

6.3.2 2011 Programs

Resource definition core drilling

This drill program was the main activity during 2011, particularly focused on the eastern main

mineralized zone of the Toroparu deposit. A total of 120 holes were drilled totaling 42,320 m. The

objective of this drill program was to increase the overall resources and the average grades of gold

and copper, as well as the conversion of resources from the Inferred to Measured and/or Indicated

categories. Priority was given to the eastern mineralized zone of the deposit (Main Zone), which has

a higher average gold/copper grade and contains around 65% to 70% of the known global

resources.

At the end of 2011, and over a period of six years (December 2006 to December 2011), a Projecttotal of 111,668 m of resource definition drilling was realized in 225 core holes.

Metallurgy – Gold Deportment Study

This study was carried out by SGS on a 400 kg composite sample of the Toroparu mineralization

and collected in 23 different core holes. Results were received in May 2011. The objective of this

investigation was to determine the occurrence of gold, including microscopic and submicroscopic

gold in the sample, and identify and evaluate any mineralogical factors that might affect potential

gold recoveries.

Step-out Exploration Core Drilling

A total of 78 core holes, totaling 24,834 m were drilled in adjacent zones north-west and south-eastof the Main Zone deposit area in order to explore for significant and economic extensions of the

resource or nearby satellite deposits.

Exploration Core Drilling

Small recon core drilling campaigns were carried out over areas with promising surface exploration

results; including the Ameeba and Manx areas, located respectively at several kilometers north-west

and north-east of the Toroparu deposit. A total of 28 holes for 8,405 m were completed.

Geochemical Exploration

A regional saprolite geochemistry sampling campaign was started in March 2011. The survey was

focused on areas with presumed geological potential for gold. Semi-regional and detailedgeochemical sampling was performed on areas where alluvial mining activities showed gold

potential. At the end of the year a total of 4,390 samples were collected.

Ground Geophysics

Combined Gradient Array IP and Magnetometer surveys were carried out over several gold

prospects, including Ameeba, Timmermans, Manx and NW of the Toroparu deposit, completing the

grids of the 2010 surveys. An additional 17 line-km of survey were completed.





LIDAR Survey

During the course of the second quarter of 2011, a LIDAR survey was flown over an area of 250 km²

around the Toroparu deposit zone. This technology (Light Detection and Ranging) is an airborne

laser swath topographic mapping (ALSM) system. It is amongst the only methodologies which

provides accurate and high precision topography contour maps in tropical forest covered zones. Adetailed topographic contour map was produced from this data.

Road Project

After the acquisition of additional road construction equipment, Sandspring commenced, in June

2011, an improvement and rehabilitation project of the access road to the Toroparu site (total

distance 240 km).

Mineral Agreement

A Mineral Agreement was signed in November 2011 between the Government of the Republic of

Guyana and Sandspring. The Mineral Agreement defines s all fiscal, property, import-export

procedures, taxation provision and other related conditions for the continued exploration, minedevelopment and mining/processing operations at Toroparu. Furthermore, the Government of

Guyana has agreed to grant a large-scale mining license, which will allow the start of commercial

production, once economic feasibility is demonstrated.

Resource Update and NI 43-101 Technical Report No. 208

Drilling through an effective date of December 31, 2010 was incorporated by P&E Engineering into

an updated mineral resource estimate and an initial PEA NI 43-101 Technical report No. 208 dated

May 5, 2011 (P&E, 2011) (Table 6.3.2.1).





During the course of 2012 a total of 34,055 m, in 142 holes were completed for the Toroparu deposit

area. At the end of 2012, and over a period of six years (December 2006 through December 2012), a

Project total of 145,723 m of resource definition drilling was completed in 367 holes.

Exploration Drilling

Exploration drilling consisted for the larger part of Reverse Circulation (RC) holes, drill testing the

main gold anomalies, which were revealed in the area around the Toroparu deposit by the saprolite

geochemical program during the course of 2011. The total RC drill meterage amounted 15,633 m in

168 holes.

Geochemical Exploration

Regional and detailed saprolite hand-auger sampling and testing concerning regional gold potential

and local gold anomalous zones was conducted. A total of 3,480 samples were collected. Over the

course of 2011 and 2012 the geochemical surveys covered around 450 km² and a total of 7,850

geochemical samples were collected.

Airborne Geophysics Reprocessing

Re-analyses of the 2008 airborne geophysical survey data was carried out by a geophysical

consultant. The work consisted in a basic structural interpretation of the aeromagnetic and

radiometric data, and an attempt to characterize the geophysical signature of the Toroparu deposit.

This study contributed significantly to the development of a regional exploration model.

Road Rehabilitation Project

The road repair and maintenance work continued for the whole year. In 2012 a road work contract

was signed with the GGMC (Guyana Geology and Mining Commission), financing part of the total

road rehabilitation costs. Part of the work (+/-100 km) was subcontracted to MMC (Mekdeci

Machinery and Construction), a local construction company.

Preliminary Road Construction Study

After several field visits, FMG Engineering completed a Preliminary Design Study in March of 2012

on the access road reconstruction, from the Itabali port facility to the Toroparu site (230 km);

including a conceptual roadway reconstruction design plan, cost estimates and preliminary

solicitation of qualified contractors.

Environmental Permit

This permit was signed and granted to Sandspring Resources by the Environmental Protection

Agency in June 2012.

Hydro-electrical Project

A monitoring program to assess Kumurung river flow characteristics upstream and downstream and

rain fall measurements in the Project watershed started in September 2012 and is ongoing.

Updated Preliminary Economic Assessment

The culmination of all Sandspring work programs from inception of the land agreement in 2009

through October 2011 was the completion by P&E and Sandspring of an updated PEA (Scoping

Study) and a NI 43-101 Technical report No. 234 titled “Updated Resource Estimate and Preliminary





In 2003 ETK submitted a composite sample of saprolite, weighing 150 kg and with a calculated

grade of 2.80 g Au/t, to Hazen for testing (Hazen, 2003). Test work showed that with

minus 100 mesh grinding and gravity concentration, up to 80.1% of the gold could be recovered.

Recovery increased to 93.9% with froth flotation of the gravity tailings.

Beginning in 2004, ETK intermittently carried out test-mining of the saprolite from the pit, neverachieving recoveries of more than about 18%. Low recoveries were attributed to the predominance

of fine, micron-sized gold flakes in combination with high viscosity saprolite ore slurries, making

gravity capture of the gold very inefficient.

In October 2005, Meixner and Wesa (Meixner, 2006) sampled ore slurries (with average grades of

2.58 g Au/t) that were being mined by ETK. Tailings samples, taken during the same time interval as

the ore slurries were being processed returned 2.13 g Au/t, indicating significant gold loss in the

concentrator plant and gold accumulation in the tailings pond.

During the period from November 19, 2005 to February 28, 2006, accurate production records were

kept by ETK of saprolite ore processing through the pilot gravity separation plant (Shaffer, 2006). A

total of 59.625 kg (1,917 oz) of gold were recovered from 199,297 t of saprolite ore for a calculatedrecovered grade of 0.30 g Au/t. An average grade of 2.64 g Au/t was determined and analyzed at

Loring Laboratories in Georgetown. The recovery rate of the gravity circuit was calculated to be 11%.

In February 2006, R. Hyyppa, P.E., consulting mining engineer to ETK, conducted an analysis of the

gravity separation procedure (Hyyppa, 2006b) at Toroparu by testing the addition of a flotation circuit

to improve gold recoveries (Hyyppa, 2006a). Results showed that 99.5% of the gold in the

concentrate occurred in the -150 mesh fraction and that between 49% and 95% of the gold was

theoretically recoverable by flotation.

In 2009, Sandspring contracted SGS Lakefield Research Limited Metallurgical Testing, to carry out

initial metallurgical testwork on samples of saprolite ore, saprolite tailing mixture and hard rock from

the Toroparu deposit. Testing has indicated positive results for copper and gold in both the hard rock

and saprolite samples.





7 Geological Setting and Mineralization All geological data were collected and compiled and geological interpretations were performed by the

Sandspring exploration team under the management of L.W. Claessens, P. Geo and VP-Exploration

and Pascal van Osta, P.Geo and Exploration Manager. This section of the report is the result ofthese efforts and is written by the exploration managers.

7.1 Regional Geology – Guiana Shield

The Guiana Shield underlies the eastern part of Venezuela, Guyana, Surinam, French Guyana and

parts of northern Brazil. It consists of three major geological subdivisions (see Figure 7.1.1). In

Venezuela the Imataca Complex basement rocks are composed of Archean age formations of high-

grade metamorphic rocks and dispersed granitoid plutons, all older than 3.0 Ga. The younger

granitic and volcano-sedimentary terrains are of Paleo-Proterozoic age, ranging from 2.2 to 2.0 Ga,

and are unconformably overlain (covered) in the western part of the shield by the anorogenic clastic

sedimentary sequences of the early Mid-Proterozoic Roraima Formation.

The Toroparu property is located in northwestern Guyana, which is mainly underlain by alternating

volcano-sedimentary belts and large granitoid batholiths of Paleo-Proterozoic age. These

supracrustal rocks form the northern part of the Guiana Shield, which represents the northern

segment of the Amazonian Craton in South America, and is a dismembered portion of the West

African Craton. The West African Craton is well known for its gold potential and numerous tentative

correlations have been made to compare these lower Proterozoic terrains.

Over the last several decades numerous economic gold deposits were discovered in the West

African Craton, in particular in the lower Proterozoic volcano-sedimentary sequences. Most of these

deposits are in production, examples are Obuassi, Ayanfuri, Ahafo, Tarkwa, Chirano and Boguso

gold deposits in Ghana; the Sadiola, Yatela, Tabakoto, Morila and Syama deposits in Mali; theSabodala deposit in Senegal; the Essakane, Taparko, Mana and Youga deposits in Burkina Faso

and the Tongon, Ity and Bonikro gold deposits in Côte d’Ivoire.

The larger part of the Guiana Shield is geologically underexplored. Geological mapping and regional

exploration is hampered by dense tropical vegetation and thick lateritic/saprolitic weathering profiles.

Nevertheless, apart from the significant gold discoveries at Las Christinas, El Callao and others in

the Kilometre 88 district of Venezuela, Omai in Guyana, and Gros Rosebel in Suriname, increasing

alluvial mining and exploration activities since the 1990’s has demonstrated the excellent gold

potential of the Guiana Craton portion of the Amazonian Craton. Sizeable gold deposits have been

defined in metamorphosed volcano-sedimentary sequences in the Guiana sub-region. In Guyana for

example, multi-million ounce gold deposits occur at Aurora, Toroparu, and Hicks/Smart, and less

than 1 Moz gold deposits are present at Tassawini, Eagle Mountain and Million Mountain.

In the northern and northwestern parts of Guyana, the supracrustal sequences constitute the

Barama-Mazaruni Supergroup and form three curved, northwest-southeast oriented sub-parallel

belts, which show a similar regional lithostratigraphy. Limited field information seems to indicate that

each of the belts is comprised at the base of mafic tholeiitic basalts and minor ultramafic rocks,

overlain by volcanic rocks of intermediate composition alternating with terrigeneous sediments.

These sequences are interpreted to have formed as successive back-arc closure and extensional

oceanic-arc systems between 2200 and 2100 Ma (G.Voicu et al., 2001). In Suriname and French





Guyana, molasse type sediments form sequences of siltstones, greywackes and conglomerate,

unconformably overlying the volcano-sedimentary sequences. Geochronological data suggest ages

around 2125 Ma, which correlates well with the Tarkwanian, gold bearing, clastic sediments of West

Africa (Milesi, 1995). The extension of these terrigenuous facies to the west into Guyana has not

been mapped, but there are indications they exist.

Crustal shortening is reflected by polyphased deformation, which resulted in shearzone dominated

strain and tight folding, arranging the volcano-sedimentary sequences in more or less elongated

belts.

The above described supracrustal sequences are intruded by numerous, large and small calc-

alkaline felsic to intermediate granitoid intrusions, called the “granitoid complex”, with ages ranging

from 2140 to 2080 Ma (G.Voicu, et al., 2001). These plutons form large batholithic zones in between

the volcano-sedimentary belts, and as small plutons within the belts.

The lack of systematic geological mapping data and large scale remote sensing studies results in the

regional framework of the Paleo-Proterozoic terranes of the Guyana shield being poorly

documented. The region is marked by several large scale shear zones. The most prominent of thesestructural corridors stretches over several hundreds of kilometers in a west-northwesterly direction

across most of the Guyana Shield. In Guyana this feature is known as the Makapa-Kuribrong Shear

Zone (MKSZ; G.Voicu, et al., 2001). An interesting observation is that a majority of the known gold

mineralization systems are located in the vicinity of these regional tectonic features.

7.2 Regional Geology – Western Guyana

The concession package of Sandspring (1000 km²) is located in the Upper Puruni area, in between

the Cuyuni and Mazaruni rivers, in the north-west part of Guyana. The regional geology of this area

is described by Heesterman, et al., in a 2001 Guyana Geology and Mining Commission (GGMC)

report, as well as in several of Heesterman’s internal ETK reports dated 2003 and 2004, andupdated in 2005 (see references section). Voicu, et al. (1999), gives a concise description of the

regional geology of the Omai mine area, which reflects a broadly similar geology to that at Toroparu.

Figure 7.2.1 shows the most current geology sketch map of the Upper Puruni region.

Geological mapping is hampered by dense tropical vegetation and thick lateritic/saprolitic weathering

profiles, causing a general lack of bedrock exposure. As a consequence regional geology knowledge

is quite limited. In the context of these limitations Sandspring’s geology team made an attempt to

draw a comprehensive litho-structural sketch of the Upper Puruni area, using all available regional

data (Project and public data): airborne magnetics and radiometrics, topographic documents (DTM,

SRTM maps, JERS sat images), existing geological maps, and regional geochemical data. This work

tries to provide a contribution to the overall understanding of the regional litho-structural patterns.

Combining geophysical and geochemical features with topographic landscape textures along with

basic information from the official geological map, it was possible to distinguish several probable

volcano-sedimentary sequences and intrusive structures within a regional tectonic framework (Figure

7.2.1).

The northeastern half of the Upper Puruni concession is underlain by thick volcano-sedimentary

sequences consisting of alternating mafic, intermediate and to a lesser extent, felsic volcanic flows

and pyroclastics, with intercalated sedimentary successions, generally metapelites and greywackes.

These formations form the Puruni volcano-sedimentary (VS) belt which extends in a northwesterly





direction in between two large plutonic areas, the Aurora batholith located to the northeast of the

concession, and the Putareng batholith underlying most of the southwestern part of the property

(Figure 7.2.1). Regional metamorphic grade is greenschist facies and can reach the amphibolite

facies in the vicinity of the granitoid intrusions. Limited lithological information provided by scarce

outcrops and exploration drill logs suggest that the central part of the belt is predominantly occupied

by thick sequences of pyroclastics and sediments; whereas, the border zones are dominated by

mafic volcanics. Some strongly weathered rock in road cuts, and associated multi-element

geochemistry, suggest the presence of ultramafic facies, which seem to be related to the mafic

volcanic sequences.

The Putareng batholith corresponds to a calc-alkaline composite intrusive complex, ranging in

composition from granite and tonalite to diorite. In the literature it is suggested that these intrusives

developed synchronous to late in the orogenic cycle. Exploration revealed the existence of small,

more or less elongated, intra-belt plutons, generally of tonalitic to quartz-dioritic composition. The

Toroparu gold-copper deposit developed along the contact zone of one of these small intrusive

bodies (Figure 7.2.1). Reprocessed airborne magnetics data and satellite imagery interpretations

provide indications that these small plutons seem occur preferentially at “Mag low” structures along

the southern limb of the Puruni VS belt (Figure 7.2.1). Several significant gold deposits in Guyana

are related to such small intrusive bodies: Aurora, Omai and Toroparu. Petrographic,

geochronological and litho-geochemical studies are required to investigate in detail the age of the

different intrusive phases and their eventual l ink with gold (copper) mineralization.

Younger, mafic intrusions are widespread over the area and form generally irregular shaped bodies,

probably remainders of large sills and dikes, respectively. unconformably overlying or discordantly

cutting through the Paleo-Proterozoic formations. These mafic intrusives consist mainly of dolerites

or diabase and are related to the early Meso-Proterozoic Roraima basin formation.

Remote sensing imagery (SRTM and JERS) and airborne geophysical maps provide useful

information and allow preliminary interpretations of the regional tectonic framework of the Upper

Puruni area. However the general lack of bedrock outcrops and hence geological field observations

hinder further study of these structural interpretations.

The Upper Puruni area is marked by sets of NW to WNW and NNW to N-S lineaments (Figure

7.2.1). The NW oriented features seem to constitute typical belt parallel shearing structures,

following lithological contact zones and dominating the regional trend of the belt. The regional

structural pattern shows a sigmoidal flexure zone in the northwestern part of the concession, which

seem to be controlled by the set of NNW to NS lineaments. The flexure zone, if the fractures are

strike-slip shear zones, can be an area of right-hand rotational deformation. Unfortunately, there is

very little structural information available, which makes basic and reliable structural analyses difficult.

The Toroparu deposit occurs close to the crossing of the WNW trending Puruni lineament and the

NNW oriented Wynamu lineament.

7.3 Property Geology

A significant resource definition drilling program was completed from the end of 2006 to the end of

2012 on the Toroparu deposit, which provides the basis for understanding the Project geology. Drill

core logging allowed for detail understanding of the lithology, structure, and the gold–copper

mineralization in bedrock.





7.3.1 Weathering

Toroparu is located in the immediate vicinity of the confluence of the Puruni and Wynamu rivers, is in

a topographically low area, and the upper part of the lateritic profile has been eroded. The bedrock

substratum is overlain by a thin, on average 1 to 2 m residual soil layer, followed by a 25 to 35 m

thick layer of saprolite. Sap-rock is the transitional zone between saprolite and fresh rock, and formsas a gradational contact a few meters thick.

Saprolite is the result of deep tropical weathering, resulting in the larger part of the original rock

mineralogy being replaced by clays. However, part of the original rock textures can be preserved

with clay pseudomorphs. Quartz veins and veinlet networks survive quite well in saprolite, and

contain occasional free gold grains. In general, sulfides are completed leached and removed in the

saprolitic weathering layer, leaving relict voids or oxidized spots. Sulfides can be partly preserved in

the sap-rock horizon.

7.3.2 Lithology

The Toroparu gold-copper deposit occurs along the northwestern boundary of a tonalitic to quartzdioritic intrusion, close to the south-eastern edge of the pluton (Figures 7.3.2.1 and 7.3.2.2).

Throughout most of the deposit zone, hydrothermal alteration is quite intense and hampers

macroscopic as well as microscopic descriptions in order to identify and distinguish the volcanic and

intrusive rock types. Moreover most of these magmatic lithologies display comparable intermediate

mineralogical compositions.

On a deposit scale, the western part of the Toroparu mineralization system and the SE satellite

deposit are predominantly hosted by intrusive rocks (Figure 7.3.2.1: West Zone, Main West Zone

and SE Zone). The abundant presence of xenoliths of volcanic rocks indicates that the zones

correspond to the roof of the intrusive. In the eastern part of the deposit area (Main East Zone), the

mineralization forms an elongated cloud along a contact zone of a greenschist metamorphic volcanicsequence, draped over a deeper seated tonalitic intrusive (Figures 7.3.2.1 and 7.3.2.2).

The different zones of mineralization are interpreted to be separated by WNW and NNW oriented

fault sets.

The intrusive lithologies are tonalite to quartz diorite in composition and display a medium grained

granular (hypidiomorphic), massive, but often porphyritic texture.

The tonalites intrude a sequence of greenschist metamorphic volcanics of intermediate to mafic

composition, consisting of fragmental pyroclastics (possibly volcanic breccia or debris flows with

predominantly felsic clasts) alternating with fine grained tuffaceous layers, grading locally into

coarser lapilli and local intermediate to mafic lava flows, often porphyritic. North of the deposit areathe pyroclastics grade into fine grained and laminated arenaceous and pelitic sediments.

At depth, in the vicinity of the above described intrusive contact, and associated with zones of higher

grade gold, several core holes intersected dacitic to quartz-andesitic composition rocks, displaying a

fine grained, massive but often fine to micro-porphyritic texture. These rocks, described as “probably

of sub-volcanic (hypabyssal) origin” in petrographic analyses, seem to form irregular bodies in the

meta-volcanics and intrusives. The volcanic and intrusive rocks are intruded by sets of discontinuous

sub-vertical mafic dikes of variable widths, from a few tens of cm to over 10 m. The dikes show a





relict original mafic intrusive mineralogy and texture overprinted by a greenschist metamorphism

mineral assemblage.

7.3.3 Structure

Detailed core logging shows that the volcano-sedimentary sequence and the intrusive rocks did notundergo a strong overall deformation.

The volcano-sedimentary sequence of alternating coarse and fine volcanoclastics, and lava flows

appear as massive non-foliated layers. Unit boundaries are generally not well expressed, which is

probably due to strong alteration, and that makes the observations and/or interpretations of eventual

fold systems difficult. The tonalites show an overall massive texture and appear as an undeformed

intrusive rock.

Foliation occurs locally and is probably associated to small local shear fractures. Foliation is

relatively frequent in the transition zone between the Main Eastern and Western zones and is related

to the NW-SE fault system separating the two main mineralized bodies.

On a deposit scale relatively dense fracture networks seem to occur by preference in elongated E-W

oriented and west plunging lenticular bodies, which, in particular in the Main Eastern and the SE

zones appear as higher grade mineralization features. Dense fracturing associated to higher grade

gold and copper mineralization seems to develop more or less along the intrusive contact and cross-

cuts as well the meta-pyroclastic sequences, the hypabbysal intrusives or subvolcanic facies, and

the tonalities/quartz diorites. Around these higher grade core features and towards the borders of the

deposit, fracturing intensity gradually decreases and gold and copper grades drop. A similar

structure, but less well expressed because of lower grades, has been detected in the Main West part

of the deposit.

Most of the Upper Puruni region is characterized by a pattern of conjugated sets of WNW to NW and

NNW to NS lineaments, which are probably shear fractures. The Toroparu deposit is located close toand between two major lineaments: the WNW-oriented Puruni lineament, to the south-west and the

NNW striking Wynamu linear structure. The Wynamu affects and off-sets the south-east part of the

deposit and is post-mineralization. Within such a regional structural pattern, the mineralized zones of

the Toroparu deposit can be interpreted as east-west oriented, west plunging, dilational zones within

a NW to WNW oriented, oblique sinistral strike-slip fault zone. It is clear that more structural

evidence is needed to fully support this interpretation of higher grade E-W lenses within the overall

WNW oriented deposit geometries.

7.3.4 Alteration

Over most of the deposit area the volcanic and intrusive facies are affected by a quite stronghydrothermal alteration. Core logging defines irregular zones of silicification and

sericitization/chloritization, with associated epidote. Carbonate is ubiquitous in most lithologies as

small disseminated grains in the groundmass, sometimes giving the finer grained facies a micro-

porphyritic aspect, and is abundant and associated to quartz in veinlets.

Microscopically the most common alteration assemblage, overprinting the original rock mineralogy, is

propylitic/phyllic in nature: albite – actinolite - (tremolite) – chlorite – sericite – carbonate – epidote –

± local quartz. Petrographic analyses describe a hydrothermal assemblage containing secondary




Guyana

Figure 7.2.1

Upper Puruni DistrictRegional Geological Sketch

Source: Sandspring Resources, 2013

(L.W. Claessens, 2012)





8 Deposit TypeThe existing exploration results suggest that the Toroparu deposit is a gold-copper-bearing

mineralized system hosted by a sequence of metamorphosed pyroclastics and minor volcanic flows

and sediments adjacent to an altered granodiorite pluton. The mineralization consists ofdisseminated sulfides in a veinlet and fine fracture/stockwork, which could be shear-zone related.

The genesis of the mineralized system and related alteration is not well understood and still based

mainly on macroscopic observations (core logging). Additional geological, petrographical,

mineralogical and chemical work is required to help define the deposit model and its geological

context.

The Las Cristinas and Las Brisas deposits, forming a large gold-copper mineralization system, are

located in the southeastern part of Venezuela (Channer, et al., 2005; Cristinas and Brisas Tech. Rep.

resp. 2007-2008; total reserves 27 Moz) and 150 km west of the Toroparu deposit in a similar volcano-

sedimentary belt. These deposits share the same economic constituents (gold, copper and silver) and

similar volcanoclastic host rocks. However, there are substantial differences with respect to thegeological context and the mineralization style. The Venezuelan deposits form stratiform and foliation

parallel to elongated mineralized lenses within sheared mafic pyroclastics and volcanics, and are

marked by the absence of intrusive stocks and quartz-(carbonate) vein stockwork. (Channer, et al.,

2005; Addison, et al., 2006, for Gold Reserve Inc.)

The Toroparu deposit shows a better resemblance to the Archean-aged Boddington deposit. Both

deposits are hosted by greenschist metamorphosed volcanics, sub-volcanics, and intrusives, and

show a similar mineralization style. The Boddington deposit in Australia is interpreted as a

structurally controlled, low-sulfidation, intrusion-related Au-Cu deposit formed by two overprinting

magmatic-hydrothermal events. The bulk of the mineralization and associated alteration are

genetically related to a K-rich post tectonic magmatic suite of intrusions (McCuaig, et. al., 2001).Table 8.1 shows the primary geological and mineralogical features of the Toroparu Au-Cu deposit in

relation to other similar deposit of the Guiana Craton.



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project

PC/MLM Toroparu_NI43-101_TechnicalReport_349800.020_044_MLM.docx

Table 8.1: Primary Geological and Mineralization Features of Several Gold and Gold –Copper Deposits

Toroparu Aurora Omai Eagle Mt Gros R

Company Sandspring Res. Guyana Gold Fields IamGold Eagle Mt. Corp. IamGo

Country Guyana Guyana Guyana Guyana Surina

Commidit y gold-copper-(silver) gold gold gold gold

Host lithologies pyroclastics; lava flows tonalite-diorite Q-diorite ; granodiorite porphyry clastic Ass oci ated rocks intermed. -mafic sediments, mafic volc. andesite-basalt andesite : sediments felsic v

tonalite; Q-diorites felsic volc.; sediments volc.

Metamorphic grade lower greenschist lower greenschist lower greenschist lower greenschist middle

(of host rocks)

Structural setting oriented brittle fracture vein stockwork (intr) stockwork vein stockwork (intr) brittle-d

network in volc./intr. brittle-ductile brittle shearing shearing( volc-sed)

shearing (volc-sed)

Ore/gangue mi nerals chalcopyrite-bornite- pyrite pyrite, galena, chalcopyrite, magnetite, pyrite,

pyrite-molybd.-chal- quartz-ankerite pyrite, pyrrhotite, molybdenite, chalcopyrite-q

cite, quartz, carbonate sphalerite, molybd. pyrite, scheelite feldspa

quartz, biotite

Hydrothermal alter. sulfidation; carbonatiz.; silicification; sericitiz.; sulfidation; carbonatiz.; silicific.; sulfidation; potassi

silicific.; sericitiz.; albitisation; carbonatiz. silicific.; sericitiz.; sericitiz.; chloritiz;. carbon

chloritiz.; albitization sulfidation chloritiz.; albitization argillic Structural ti mingof mineralization

Late to post- tectonic syn- to late tectonic late to post tectonic late to post tectonic syn- to

Age o f m ineral izat ion unknown unknown 2002 Ma unknown unknow

Source: Voicu et al., 2001; modified by Sandspring Resources, 2013)





8.1 Geological Model

The existing exploration results suggest that the Toroparu Deposit is a gold-copper-bearing

mineralizing system hosted by a sequence of metamorphosed pyroclastics and minor volcanic flows

and sediments adjacent to an altered granodiorite pluton. The mineralization consists of

disseminated sulfides in a veinlet – fine fracture stock work, which could be shear-zone related.

During 2012, a re-logging exercise was performed of most of the existing holes within the resource

zones with the objective to standardize geological descriptions and develop a reliable geological

model of the Toroparu gold-copper deposit, including the definition of geological limits for the

resource modeling.

Geological cross-sections (Figure 7.3.2.2) were generated by Sandspring geologists and examined

with the drillhole database. It is difficult to identify clear litho-structural boundaries for the

mineralization system. The drilling delineates zones of Au-Cu veinlets and mineralized fractures

containing minor sulfides (pyrite and chalcopyrite primarily), defined as stockworks or possibly shear-

related WNW to West trending and near vertical structural zones, and located dominantly in

greenstone metamorphosed felsic to intermediate volcano-sedimentary rocks at or near the contact

with granitoid (diorite/granodiorite) intrusive rocks. The structurally controlled mineralization cuts

across all rock types, except for late mafic cross-cutting dikes (less that 10% by volume of the total

rock volume), and thus the mineralized structural zones define the modeled domains, not lithologies.

Lithology wireframes were not constructed as the gold copper mineralization is cross-cutting. The

geological model is therefore composed of oriented grade shells around the mineralization defined in

drilling and from limited trench sampling, and represent the structurally controlled zones of

veinlet/stockwork Au-Cu mineralization.





9 ExplorationExploration carried out at the Toroparu from 2003 to 2012 is briefly described in Section 6.0 (History)

of this report. For detailed information about ETK/Sandspring exploration activities prior to 2011 the

reader is referred to previous NI 43-101 Technical Reports that are publically filed on SEDAR byP&E and Sandspring listed in the References Section 27 of this report (P&E, 2009, 2010 and 2011).

This Section 9 is devoted to discussion of the exploration programs of Sandspring during 2011 and

2012; in particular, drilling that has been completed since P&E and Sandspring completed the PEA

NI 43-101 Technical Report dated March 12, 2012 (P&E, 2012).

9.1 Exploration – 2011 and 2012

Up to the end of 2010 all exploration was focused on the Toroparu deposit and immediate

surrounding areas. At the end of the 1st quarter of 2011 an important exploration campaign was

launched with a main objective to test the gold potential of the Company’s 1000 km² Upper Puruni

Concession. This program was consisting of systematic regional and semi-regional geochemistrysampling, geological mapping, ground geophysics and reconnaissance exploration drilling.

Geochemistry (Saprolite) Sampling

During 2011 and 2012 a total of 7,390 samples were collected (Table 9.1.1). A saprolite sample is

taken by hand auger (3 1/4 inch diameter) at a depth of 1 to 3 m, below the soil or lateritic crust or

alluvial layer. Sampling is carried out following a regular grid pattern of 1000 m x 100 m in the

regional surveys, and 400 m x 50 m or 200 to 250 m x 50 m for the semi-detailed and detailed

surveys, respectively. Locations of hand auger holes are recorded by a Garmin handheld GPS.

Blank and standard QA/QC samples are inserted. Duplicate samples are prepared in the field

(sample splitting). The samples are sent for analyses to the Acme Vancouver laboratory for ICP 34

elements 1F03 (30 g samples). As of December 2012, approximately 400 km² were covered withsaprolite auger sampling within the concession lands.

Table 9.1.1: Saprolite Sampling in 2011 and 2012

Source: P&E, 2013

Saprolite Geochemistry sampling

Zone Total in 2012 Total in 2011 To tal

Ameeba 258 894 1152

Toroparu 555 149 704

Sona Hill (Toroparu Creek) 348 0 348

Puruni 540 1028 1568

Manx‐Timermans 230 289 519

Red Dragon 603 1997 2600

Makapa

(north

Ameeba) 499 0 499Total samples collected 3033 4357 7390

Blank 187

Duplicate 176

Standard 197

Total samples analyzed 7950





A total of 10 gold anomalies have been detected, which form a cluster around the Toroparu deposit

(Figure. 9.1.1).

Exploration drilling, primarily RC drilling, was carried out on the most prospective anomalies:

Ameeba, Toroparu W, Red Dragon North, Road, and Sona Hill.

Geological Mapping

The Upper Puruni concession is located in a poorly understood early Paleo-Proterozoic granite-

volcano-sedimentary terrain of the Guiana Shield in western Guyana. The geological knowledge is

limited due to large areas of virgin tropical forest with minimal to no road access and intense tropical

weathering. Rock outcrops are very rare, which significantly hampers systematic geological

mapping. As a result existing geological maps are quite inaccurate and lack detail. As a

consequence systematic mapping is not possible.

Sandspring has created a litho-structural sketch map of the Upper Puruni area using all available

regional data (Project and public data): airborne magnetics and radiometrics, topographic documents

(DTM, SRTM maps, JERS sat images), existing geological maps and regional geochemical data.

This contributed to the overall understanding of the regional litho-structural patterns.

Over 144 samples were collected from outcrop and road saprolite exposures. Fourteen samples

were submitted for petrographic description.

Exploration (Reconnaissance) Drilling

In 2011 exploration consisted mainly of core drilling since no reverse circulation (RC) rigs were

available in Guyana. Reconnaissance core holes were drilled at Ameeba, Timmermans and Manx,

respectively for 5,964 m, 1,116 m and 2,441 m; for a total of 9,521 m. These programs did not reveal

any significant mineralization systems on outlying targets.

An RC rig was commissioned by the end of 2011 and during the course of 2012 a 15,400 m

reconnaissance RC program was completed for several prospects (Table 9.1.2). The exploration

drilling program was designed to test the gold anomalous surface features revealed by the regional

and semi-regional saprolite geochemical sampling programs. The RC program was focused on the

following drill target areas: the Toroparu NW targets containing several zones NW of the Toroparu

deposit; the Ameeba Zone; the Red Dragon north and Road sectors; and the Sona Hill area

(Figure 9.1.1).

Table 9.1.2: 2012 RC Reconnaissance Drilling Program


QAQC

Duplicate Blank Standard

Toroparu twin holes 3 384 5

Toroparu NW 71 5242 5

Tororparu W 7 705 10

Ameeba 29 2879 5 1754 85 58 58

Red Dragon North 21 2057 10

Red Dragon Road 15 1146 10

Sona Hill 30 2969 10 944 145 62 62

TOTAL 176 15382 7362 569 344 338

1120 170 73 74

3544 169 151 144

Samples

collectedPropsect name hole metrage

sample

interval





Relatively small areas of mineralization were intercepted in the Toroparu NW and Ameeba areas,

sufficient to justify further test drilling. The Red Dragon Zones did not encounter significant

mineralization. RC drilling on the Sona Hill, located 5 km to the SE of Toroparu area, showed

encouraging and significant gold intercepts which required follow-up drilling. A follow-up preliminary

core drilling program of 810 m in five holes continues to show encouragement; additional core drilling

is warranted for this satellite area.

Geophysics

Combined Gradient Array IP and Magnetometer test surveys were carried out over several gold

prospects; Ameeba, Timmermans, Manx and NW of the Toroparu deposit, completing the grids of

the 2010 surveys. An additional 17 line-km were added and a total of 102 line-km of ground

geophysics were realized.

Reprocessing o f Airborne geophys ical data

Re-analyses of the 2008 airborne geophysical survey data was carried out by a geophysical

consultant. The work consisted of a basic structural interpretation of the aeromagnetics and

radiometrics and identification of the geophysical characteristics of the Toroparu deposit. This study

contributed significantly to the development of a regional exploration model.

LIDAR Survey

During the 2nd quarter of 2011 a LIDAR survey was flown over an area of 250 km² around the

Toroparu deposit zone. This technology is an airborne laser swath topographic mapping (ALSM)

system. It is amongst the only methodologies, which provides accurate and high precision

topography contour maps in tropical forest zones. The resultant topographic map has sufficient detail

for feasibility level study and engineering.

Resource Definition Infill Drilling

Project drilling previously reported in the PEA of 342 holes for 147,529 m; through hole TPD-265

included drilling through August 2011 (effective date of the PEA of January 30, 2012), as deposit

definition drilling.

An important part of the 2012 drilling activities was focused on completing the Resource Definition

drill grid over the Toroparu deposit. This targeted conversion of Inferred resources to Measured &

Indicated resources. During 2012 a total of 34,055 m in 142 holes were performed on the Toroparu

deposit area.

9.2 Relevant Exploration Work – Post-PEA Drill ing

Sandspring conducted additional drilling from September 2011 through December 2012, which isincluded in the current updated resource estimate as presented in Section 14 of this report. That

additional drilling was composed of the following:

Post-PEA drilling form August 2011 to September 2012: 166 holes for 44,096 m; through

hole TPD-426; and

A program of targeted infill drilling conducted from September through December 2012,

resulted in 48 holes for 12,163 m in both the Main Zone and the Southeast Zone.





9.2.4 Signif icant Results and Interpretation

As of an August 2012 interim resource model, approximately 27% of the total resources within the

Main Zone and approximately 39% of the Southeast Zone within an ultimate pit shape were

classified as Inferred resources; combined representing nearly 2.0 Moz of Inferred classified material

– the target of the in-fill drilling.

Figure 9.2.4.1 shows drill collar locations and drillhole traces for all exploration drillholes at Toroparu

as of August 2012. Figure 9.2.4.2 shows the drill collar locations (without drillhole traces) for the PEA

drilling. Figure 9.4.2.3 shows the drillhole collar locations (without drillhole traces) for the post-PEA

drilling, and Figure 9.4.2.4 shows the targeted in-fill drilling collar locations (without drillhole traces).

Figure 9.2.4.5 shows a cross-section, looking horizontal along Azimuth 297, of drillholes and an

interim 2012 resource block model, color coded by classification. Targeted areas of Inferred

classification (blue) are noted as are proposed in-fill drillholes as of August 2012. Figure 9.2.4.6

shows the same cross-section showing the Au grade of the targeted Inferred classification

mineralization. Figure 9.2.4.7 shows the location of targeted in-fill drilling assay data.

Figures 9.2.4.8 and 9.2.4.9 show the results on classification and grade of blocks as a result of the

targeted in-fill drilling program, which was successful in converting a substantial amount of Inferred

mineralization to Measured and Indicated classification.




Guyana

Figure 9.1.1

Saproli te Auger SamplingToroparu

Source: Sandspring Resources,

2013




Guyana

Figure 9.2.4.1

Toroparu – 2013 Drillhole CollarLocations and Traces – All Holes(486), Current Mineralized Shape

Source: SRK Consulting, 2013




Guyana

Figure 9.2.4.2

Toroparu – Drillhole Collar Location – PEA Dri ll ing through August 201

(342 holes)Source: SRK Consulting, 2013




Guyana

Figure 9.2.4.3

Toroparu – Drillhole Collar Location – Post PEA and Pre Targeted In -Fil

(166 holes)Source: SRK Consulting, 2013




Guyana

Figure 9.2.4.5

Cross-Section Showing BlockClassification and Areas of Targete

Inferred MineralizationSource: SRK Consulting, 2013




Guyana

Figure 9.2.4.6

Cross-Section Showing Block AuGrade and Areas of Targeted

Inferred MineralizationSource: SRK Consulting, 2013




Guyana

Figure 9.2.4.7

Cross-Section Showing TargetedInferred Blocks and In-Fill Drill DatSource: SRK Consulting, 2013




Guyana

Figure 9.2.4.8

Cross-Section showing Current BlockClassification and areas of Targeted

Inferred Mineralization that have been

Converted to Measured and Indicated






BHID FROM TO AU g/t CU % BHID FROM TO AU g/t CU %TPD022 242 243.5 0.71 0.12 TPD022 299 300.5 2.73 0.69TPD022 243.5 245 0.65 0.19 TPD022 300.5 302 2.92 0.67TPD022 245 246.5 1.13 0.22 TPD022 302 303.5 3.43 1.37TPD022 246.5 248 0.96 0.38 TPD022 303.5 305 5.21 1.30TPD022 248 249.5 0.80 0.24 TPD022 305 306.5 3.65 1.26

TPD022 249.5 251 1.03 0.24 TPD022 306.5 308 2.30 0.68TPD022 251 252.5 0.84 0.29 TPD022 308 309.5 3.18 0.51TPD022 252.5 254 2.59 0.50 TPD022 309.5 311 1.03 0.22TPD022 254 255.5 1.55 0.54 TPD022 311 312.5 0.99 0.27TPD022 255.5 257 2.02 0.43 TPD022 312.5 314 1.10 0.30TPD022 257 258.5 1.61 0.45 TPD022 314 315.5 2.17 0.54TPD022 258.5 260 2.04 0.39 TPD022 315.5 317 1.62 0.46TPD022 260 261.5 2.31 0.46 TPD022 317 318.5 1.07 0.28TPD022 261.5 263 5.99 0.54 TPD022 318.5 320 1.18 0.34TPD022 263 264.5 27.07 0.86 TPD022 320 321.5 1.42 0.50TPD022 264.5 266 8.18 0.77 TPD022 321.5 323 0.49 0.18TPD022 266 267.5 2.25 0.48 TPD022 323 324.5 0.14 0.14TPD022 267.5 269 1.89 0.44 TPD022 324.5 326 0.57 0.16TPD022 269 270.5 3.82 0.66 TPD022 326 327.5 0.70 0.20

TPD022 270.5 272 3.61 0.58 TPD022 327.5 329 0.07 0.03TPD022 272 273.5 2.95 0.62 TPD022 329 330.5 0.06 0.02TPD022 273.5 275 2.30 0.42 TPD022 330.5 332 0.86 0.24TPD022 275 276.5 2.72 0.50 TPD022 332 333.5 0.23 0.09TPD022 276.5 278 2.19 0.30 TPD022 333.5 335 1.63 0.47TPD022 278 279.5 2.78 0.36 TPD022 335 336.5 2.74 0.38TPD022 279.5 281 2.00 0.38 TPD022 336.5 338 0.35 0.06TPD022 281 282.5 4.62 0.78 TPD022 338 339.5 0.17 0.04TPD022 282.5 284 1.26 0.26 TPD022 339.5 341 0.01 0.02TPD022 284 285.5 1.49 0.27 TPD022 341 342.5 1.44 0.27TPD022 285.5 287 1.49 0.53 TPD022 342.5 344 0.84 0.21TPD022 287 288.5 1.61 0.62 TPD022 344 345.5 0.08 0.03TPD022 288.5 290 3.63 0.68TPD022 290 291.5 3.93 0.88





Guyana

Figure 10.1

Toroparu – Plan Map of Drillholesand Traces, Showing Mineralized

Envelope

Source: Sandspring Resources,

2013




Guyana

Figure 10.3.1

Example of Uncut Core Photograpfor Drillhole TPD-103





In 2005 the Santiago, Chile laboratories received ISO 9001:2000 registration with the preparation

facilities in Mendoza, Argentina and Guyana following in 2006. Acme’s Lima, Peru facility has

completed its registration audit in 2012.

In 2011 all Acme laboratories across South America were recertified under ISO 9001:2008. Their

preparation laboratory facility in Georgetown (East Coast Demerara), Guyana was recertified inJanuary 2012 after a successful audit.

Bureau Veritas completed the acquisition of AcmeLabs worldwide on February 23, 2012. The

acquisition greatly enhanced the abilities of AcmeLabs because of the capital investments made by

Bureau Veritas.

Another key benefit was the merger of AcmeLabs’ metallurgical division with that of Inspectorate’s

(also acquired by Bureau Veritas), thereby creating one of the most potent metallurgical divisions in

the industry today.

At Acme in Georgetown (East Coast Demerara), the samples are dried and the entire sample is

crushed to better than 80%, passing -10 mesh. A 250 gram split is taken and pulverized to better

than 85% passing -200 mesh. The pulps are sent to Acme in Santiago, Chile or Vancouver, British

Columbia where they are analyzed for gold and copper.

All samples were analyzed for copper by four-acid digest with AAS finish. The majority of samples

were analyzed for gold by lead collection fire assay method with AAS finish (50 gram charge). All

samples with results >10 g/t Au were further analysed by lead collection fire assay method with a

gravimetric finish.

11.5 QA/QC Procedures

Sandspring has maintained a QC program that had been initiated with drillhole TPD001. Every

sample batch prepared for analyses consists of 32 regular samples, one Coarse Duplicate, one PulpDuplicate, two Certified Reference Materials (CRM, Standards) and a blank sample.

For 2012, Sandspring initiated a check assay program using Actlabs Guyana Inc. as an outside

laboratory to provide check assays for pulp splits prepared by the primary analytical laboratory,

Acme. A total of 1,749 regular samples, 57 blanks, 58 pulp duplicates, and 56 coarse material

duplicates were prepared. Sandspring inserted 123 standards into that sample flow for a total of

2,043 check sample analyses.

11.5.1 QA/QC Resul ts

The 2012 check assay results are shown in Figures 11.5.1.1 through 11.5.1.4, which are scatter

plots of Actlabs versus Acme results, respectively for Au, Cu, Au in pulp duplicates, and Au in coarseduplicates. The results indicate no particular bias of Au or Cu between the assay labs; satisfactory

confirmation of assays.

11.5.2 QA/QC Actions

Sandspring staff in 2012 prepared monthly reports of QC samples submitted and the results of any

failures for standards or blanks, based on plots of the data against expected values and showing a

two and three standard deviation line for each standard sample value. Failures of over two standard

deviations from the expected values resulted in that sample batch being re-run.





The test program included single-pass tests at three different pressure settings in order to determine

the optimum operating parameters for the test apparatus. The test program also incorporated locked

cycle testing in order to simulate the product size distributions to be expected in an industrial sized

HPGR circuit. The results of the locked cycle testwork are summarized in Table 13.4.3.1.

Table 13.4.3.1: Summary of HPGR Test FindingsDescription Unit Value

Wet Bulk Density Kg/L 1.75Feed Particle Size, F80 mm 10Product Particle Size, P80 mm 2.3Pressure of Operation bar 60Moisture (% H2O) 3.6Dry Net Throughput t/h 1.5Circulating Load % 72.4Gross Specific Energy Requirement kWh/t 3.70Net Specific Energy Requirement kWh/t 3.06Specific Grinding Force N/mm 3.01Specific Throughput t*s/m *h-(mf ) 220Specific Throughput Rate t*s/m

3*h-(mc) 196

Ratio mc/mf 0.89

The results indicate that the sample material is amenable to the HPGR process.

13.5 Gravity Separation (ACO/LCO)

Gravity separation tests performed at Resource Development Inc. (RDi) and SGS Lakefield resulted

in Au recoveries from 13.4% to 52.1%. Grind sizes between P 80 50 µm and P80 300 µm were tested

and showed a general trend of higher recoveries at finer grind sizes. Gravity recoveries were

estimated to be about 38% for the P80 150 µm primary grind size.

The results of the three phases of gravity testwork are summarized in Table 13.5.1.

Table 13.5.1: Gravity Separation Result s Summary for Phase 1, Phase 2, Phase 2 Extension

GravityTest No.

Feed SizeP80, µm

Feed Weightkg

ProductMass

% Assays Au, g/t

Distribution%

Phase1

MC-04 300 2 Mozley Concentrate 0.07 282 32.5Knelson/Mozley Tailing 99.93 0.39 67.5

MC-05 150 2 Mozley Concentrate 0.12 189 36.5Knelson/Mozley Tailing 99.88 0.4 63.5

Phase2

50 20 Mozley Concentrate 0.03 652 31.0Knelson/Mozley Tailing 99.97 0.39 69.0

75 20

Mozley Concentrate 0.04 838 52.1

Knelson/Mozley Tailing 99.96 0.33

47.9

125 20 Mozley Concentrate 0.03 517 24.8Knelson/Mozley Tailing 99.97 0.44 75.2

175 20

Mozley Concentrate 0.03 370 13.4Knelson/Mozley Tailing 99.97 0.67 86.6

Phase2 Extension

G-51 228 2Mozley Concentrate 0.24 122 47.7Knelson/Mozley Tailing 99.8 0.33 * 52.3

G-52 149 2Mozley Concentrate 0.20 145 43.4Knelson/Mozley Tailing 99.8 0.37 * 56.6

Knelson/Mozley tailings is calculated from cyanidation test





test ACO-02 produced a higher grade concentrate of 25.3% Cu, 88.5 g/t Au, 224 g/t Ag and 26.1% S

at recoveries of 73.9%, 59.2%, 55.1% and 62.5%, respectively.

The ACO Composite was subjected to a six cycle LCT. The test was performed using 4 kg charges

of the ACO Composite gravity tailings ground to P80 156 µm (from test G-5). The reagents used

included PAX in conjunction with the promoter Aerofloat 208. In the two cleaning and cleanerscavenger stages, CMC was used as a non-sulfide gangue depressant. Lime was used to modify the

pH and MIBC was used as a frother and added on an as-required basis.

The results from the LCT test are shown in Table 13.6.4.2. The Cu concentrate obtained from the

last three stages (D to F) produced a concentrate containing 21.0% Cu, 56.0 g/t Au, 180 g/t Ag and

20.9% S at recoveries of 91.3%, 67.2%, 65.0% and 78.4%, respectively. The first cleaner scavenger

tailing contained 0.091% Cu, 0.77 g/t Au and 2.1 g/t Ag and constituted 10.6% of the total mass,

12.6% of the Au, and 10.2% of the Ag.

Table 13.6.4.2: Locked Cycle Test Results

Product

Wt Assays, %, g/t % Distribution

g

% Cu Au Ag S Cu Au Ag SCopper Con 190 0.8 21.0 56.0 180 20.9 91.3 67.2 65.0 78.41st Cl Sc Tailings 2599.6 10.6 0.091 0.77 2.1 0.34 5.4 12.6 10.2 17.3Rougher Tailings 21718.4 88.6 0.007 0.15 0.6 0.01 3.3 20.2 24.8 4.3Head (Calculated) 24508.0 100 0.18 0.65 2.1 0.21 100 100 100 100

An examination of the stage-by-stage results indicated that the Cu and S grades dropped abruptly

for the last two stages of the test (E & F), due to a build-up of non-sulfide gangue. From this

observation, it appears that the added CMC to the cleaning circuit was not sufficient or an extra

stage of cleaning should be applied.

Au and Ag recoveries from gravity separation test (G-5) combined with LCT at P80 150 µm was

calculated to be 79.1% Au and 69.2% Ag. The results can be found in Table 13.6.4.3.

Table 13.6.4.3: ACO Combined Result s from Gravit y Separation and LCT Tests

Grind SizeCampaign

Gravity recovery , % Cleaner Flotation, % Comb. Recovery, %

Au

Ag

Au Ag Au Ag

150 µm 38.5 11.9 67.2 65.0 79.1 69.2

13.6.5 Gold Ore with Low Copper (LCO)

The LCO metallurgical testwork program involved testing of a Master Composite and the four

Variability Composites (A to D) to evaluate response to gravity separation, rougher flotation, and

cyanidation of gravity separation concentrate. Evaluation of the effect of grind size on metallurgical

performance was limited during this testwork program and included three scoping rougher flotation

tests at grind sizes of 125, 175 and 225 µm to confirm the findings from the ACO test program.

Three scoping rougher flotation tests were performed on the three gravity tailings produced at three

different P80 grind sizes to observe the effect of grind size on rougher flotation performance.

The recovery of copper and gold increased as a function of finer grind size, while the recovery of

silver did not display this trend. The copper recovery increased from 93.6% at P 80 252 µm to 97.2%





at P80 125 µm. Similarly, the gold recovery increased from 79.7% at P80 252 µm to 84.3% at P80 125

µm. In order to compare results of these tests to the testing of the ACO sample, a P 80 grind size of

150 µm was selected for the subsequent bulk rougher flotation tests.

A total of eight bulk rougher flotation tests were performed on the gravity tailing produced from the

Master and Variability Composites, all at the P80 grind size of 150 µm. The objective of the tests wasto produce concentrate for the subsequent cyanidation testwork.

The results show that the recovery of copper ranged from 97.0% to 92.4% for the different

composites. The recovery of gold shows significant variation between samples from as low as 65.7%

to as high as 80.2%. The average of three tests of the Master Composite showed a gold recovery of

75%.

The overall combined gravity and flotation gold recovery from the Master and Variability composites

is shown in Table 13.6.5.1.

Table 13.6.5.1: Combined Gravity and Flotation gold Recovery for the LCO Composites

Composite Au GravityRecovery, % Au Flotat ionRecovery, %* Au Comb.Recovery, %

Master 67.5 75.1 91.9Ore A 42.5 69.0 82.2Ore B 39.0 80.8 88.3Ore C 57.9 68.1 86.6Ore D 46.2 76.3 87.3

*Average of all applicable tests

13.7 Cyanide Leaching (ACO/LCO)

Cyanide leaching for the 2011 test program conducted at SGS Lakefield was split into three separate

phases. Subsequent testwork in 2012 by SGS was conducted on both ACO and LCO composites.

13.7.1 Phase 1

Phase 1 cyanidation testwork was presented in two separate flowsheet processes, first was the

cyanidation of whole ore (P80 target of 150 µm and 75 µm) and the second was rougher concentrate

cyanidation.

The cyanidation testwork involved three whole ore tests, one at P80 150 µm and two at P80 75 µm.

One of P80 75 µm leaches was performed as a CIL test while the other did not have any carbon

added. Following a 48-hour leach residence time, extractions of 90.0% Au at P80 160 µm, and 89.1%

Au at feed P80 of 102 µm were observed. The CIL had a total extraction of 91.1% Au, with 88.7%

contained in the carbon and 2.4% remaining in the barren leach solution. The cyanidation testingalso included leaching of rougher concentrate in two tests where one was tested as received while

the other was reground. Extraction of 96.4% and 97.8% Au was achieved from the as received and

reground samples respectively. Cu extraction and cyanide consumption increased from 18.9% Cu

and 14.7 grams of NaCN for the as received sample test to 29.6% Cu and 27.4 grams of NaCN for

the reground sample. It appears that regrinding of the sample may have led to increased leaching of

copper resulting in a higher copper concentration in solution.





13.7.2 Phase 2

The cyanide leaching in Phase 2 included leaching of four separate products. The first set of leach

tests was performed on the gravity tailings resulting from four target grinds. Following direct

cyanidation of the gravity tailing, the leach residues were combined and then split for performing two

CND tests that were completed to a weak acid dissociable (WAD) content of 0.5 and 2.0 ppm. TheCND test residues were then tested for recovery of Cu using flotation.

The second set of leach tests was performed on rougher tailings following the rougher flotation of

each size campaign.

The third set of leach tests was performed on the rougher concentrates from the 75 µm campaign.

The rougher concentrate was split into thirds and the effect of regrind was evaluated. The three

leach residues were then combined and a CND test was completed to a WAD content of 0.5 ppm

after which the recovery of Cu using flotation was again examined.

The fourth set of leach tests was performed on Mozley vanner gravity concentrate from the 75 µm

campaign.

The purpose of this highly integrated testwork program was to evaluate the effect of grind size as

well as to evaluate the effect of the processing flowsheet on the recovery of Au and Cu.

The data shows direct cyanidation of the gravity tailing considerably improves Au recovery with an

increase of 8.5%. Alternatively, the recovery of Cu is shown to be much improved through rougher

flotation.

Despite the significant improvement in Au recovery from direct leaching the capital and operating

costs of such a process is expected to be significant due to the requirement of CND to allow

subsequent recovery of Cu. Therefore, the Phase 2 Extension program evaluated extension of the

flotation process to evaluate cyanidation of cleaner concentrate and cleaner tailing thus eliminating

the need for the intermediate CND step between cyanidation and flotation.

13.7.3 Phase 2 Extension

The cyanide leaching testwork in the Phase 2 Extension was performed on three separate process

streams. The first was the direct cyanidation of the gravity tailings resulting from two target grind

sizes P80’s of 228 μm and 149 μm. The final Au extraction value for both tests was approximately

89% which indicates grind size has little effect in Au leaching of the samples.

The second was cyanidation of cleaner concentrate and cleaner tailing from MC-32 to MC-35. The

results showed that cyanidation of the cleaner tailing was excellent with an Au extraction of 88.2%,

while cyanidation of the cleaner concentrate was poor with an Au extraction of 59.0%.

The third cyanidation test was performed on the cleaner scavenger tailings from flotation tests MC-

37 to MC-40. The result showed the final extraction of gold was about 81%.

Flotation tests MC-37 to MC-40 investigated the possibility of deporting a greater proportion of the

Au to the cleaner tailing using a larger dosage of lime to increase the pH further. The results showed

that the proportion of Au reporting to the cleaning circuit increased from 3% to approximately 5%. It

is possible that the proportion of Au reporting to the cleaner tailing could be further increased by an

even greater lime dosage, or reduced collector dosage. Also, performing a locked cycle test may

increase the proportion of Au in the Cleaner Scavenger Tail.





13.7.4 Locked Cycle Test Cleaner Tailing Cyanidation (ACO)

One Cyanidation test was performed on the cleaner tailings obtained from the locked cycle test

conducted with the ACO composite. The amount of copper and gold reporting to the first Cleaner

Scavenger Tailing in this test was 5.4% and 12.6% respectively. The extraction of gold and silver

was 72% and 64% respectively and was accompanied by a copper extraction of 34%.

The overall recovery of the gravity separation coupled with cleaner flotation and cyanide leaching of

the Cleaner Tailing is summarized in Table 13.7.4.1.

Table 13.7.4.1: ACO Gravity , Cleaner Flotation, and Cleaner Tail Leach Summary

Grind SizeCampaign

Gravity Cleaner Con Cleaner tail Cleaner Tail CN Combined Au Ag Au Ag Au Ag Au Ag Au Ag

150 µm 38.5 11.9 67.2 65.0 12.6 10.2 72.2 64.2 85.4 74.9

From comparison of the results of the ACO test program, cyanide leaching of the gravity separation

tailing offers higher gold and silver recovery than rougher and cleaner flotation combined with

leaching of the cleaner tailing. However, the cyanide consumption and copper extraction from the

cleaner flotation processing route is 0.11 kg/t and 1.8% Cu respectively. This is considerably lower

than the gravity tailing leaching route that resulted in cyanide consumption and copper extraction of

1.24 kg/t and18.0% Cu respectively.

13.7.5 Gravity Concentrate Intensive Cyanidation (ACO)

The response of the gravity concentrate stream, gravity tailings stream, and cleaner scavenger

tailings stream to cyanide leaching was examined in a series of tests.

The Mozley concentrate obtained from gravity tests G-1 to G-4 were submitted for intensive

cyanidation. The tests were performed at 5% solids with 20 g/L of NaCN for 24 hours.

Au extraction was 99% for all cases and Ag extraction was 97%. Results show that the feed size has

no effect on leaching recovery.

The gravity tailings of gravity test G-1 to G-4 at four P80 sizes were submitted for bulk cyanidation

tests. The tests were performed at 40% solids with 0.5 g/L of NaCN as Carbon-in-Pulp (CIP) tests in

which the carbon was added after 48 hours of leaching. Carbon contact time was about six hours.

The final Au extractions at 54 hours ranged from 71% for CN-12 at P 80 240 μm to 87% for CN-5 at

P80 72 μm and CN-10 at P80 179 μm. Overall, the gold extraction occurred in a narrow range and

showed a general trend of an increase in gold extraction with decreasing feed size.

The final Ag extractions at 54 hours were distributed in a very narrow range from 72% for CN-12 at

P80 240 μm to 77% for CN-5 at P80 72 μm and CN-7 at P80 117 μm.

The final copper extractions for all the leaches ranged from 16% for CN-9 at P 80 166 μm to 20% for

CN-7 at P80 117 μm. The copper extraction for all tests was low and does not appear to be

influenced by varying feed size.

The combined extraction of gold and silver from the gravity separation stage and cyanide leaching

ranged from 82% for CN-11 and CN-12 (both P80 240 μm) to 94% for CN-5 at P80 72 μm for gold and

from 75% for CN-12 at P80 240 μm to 80% for CN-5 at P80 75 μm for silver.





The copper extraction for all the leaches did not exceed 14% at 54 hours. The iron extractions were

very low and did not exceed 1.5%; however, there was still a significant amount of iron present in

solution which resulted in high cyanide consumptions.

The combined recovery of gravity separation, rougher flotation and cyanidation of the rougher

concentrate for gold and silver are shown in Table 13.7.8.1. The combined recovery for gold rangedfrom 79% (CN-32R) to 85% (CN-31) and for silver ranged from 30% to 50% for CN-35 and CN-33

tests, respectively.

Table 13.7.8.1: LCO Combined Result s from Gravity Separation, Rougher Flotation andRougher Concentrate Leaching

CN TestNo.

OreType

Recovery, %

Gravity

Gravity TailFlotation

Leach Combined*

Au Ag Au Ag Au Ag Au Ag

CN-30 Master 50 11.4 75.7 50 92 74 84.8 44.2CN-31 Master 50 11.4 77 49.2 92 72 85.4 42.8CN-36 Master 50.2 12.2 72.7 48.2 93 60 83.9 37.6

CN-36R Master 50.2

12.2 72.7 48.2 95 81

84.6

46.5CN-32 A 42.5 8.8 72.3 41.2 95 66 82.0 33.6CN-32R A 48.1 11 65.7 35.8 91 62 79.1 30.8CN-33 B 38.8 11.2 80.8 55.1 91 80 83.8 50.3CN-34 C 57.9 4.1 68.1 45.6 94 82 84.8 40.0CN-35 D 46.1 11.2 76.3 29.8 89 70 82.7 29.7

*Leach of rougher float concentrate

13.7.9 Comparison of Cyanide Leaching of Gravity Tailing and Flotation Products for ACO and LCO

The main objective of the ACO and LCO testwork program was to compare the performance of direct

cyanide leaching of the gravity tailing and cyanide leaching of a flotation product. The flotation

product of interest was flotation cleaner scavenger tailings for ACO program while it was a rougher

concentrate for the LCO program.

Tables 13.7.9.1, 13.7.9.2, and 14.7.9.3 show comparisons of the overall gold recovery for three

different combinations for ACO testwork program:

Table 13.7.9.1: Overall Gold Recovery for ACO Composi te, Gravity and Gravi ty TailingCyanide Leaching

GrindSize

Campaign

Au Recovery, %Gravity Gravity Tail CN

Leach*

Combined

Au

Ag

Au Ag Au Ag75 µm 54.9 12.7 86.3 76.0 93.8 79.0

125 µm 43.9 8.3 83.3 76.4 90.6 78.3175 µm 39.1 10.5 82.8 73.8 89.6 76.5225 µm 37.3 9.6 71.6 73.8 82.2 76.3





Table 13.7.9.2: Overall Gold Recovery for ACO Composi te, Gravity and Gravi ty TailingRougher Flotation

Grind SizeCampaign

Au Recovery %

Gravity recovery, % Flotat ion Recovery, %* Comb. Recovery, %

Au Ag Au Ag Au Ag

75 µm 54.9

12.7 77.8 78.8 90.0 81.5 125 µm 43.9 8.3 80.6 81.0 89.1 82.6

175 µm 39.1 10.5 80.5 81.1 88.1 83.1

225 µm 37.3 9.6 76.1 77.9 85.0 80.0

Table 13.7.9.3: Overall Gold Recovery for ACO Composi te, Gravity , Gravity Tailing CleanerFlotation and Cleaner Tail cyanide Leaching

Grind SizeCampaign

Au Recovery %

Gravity Cleaner Con Cleaner tailCleaner TailCN Leach

Combined

Au

Ag

Au Ag Au Ag Au Ag Au Ag

150 µm 38.5 11.9 67.2 65.0 12.6 10.2 72.2 64.2 85.4 74.9

The results show that while the gravity separation and leaching of the gravity separation tailing for all

grind sizes offers higher overall gold recovery, it is not a large enough difference to compensate for

the loss of saleable Cu and the increased cyanide consumption.

Similarly Tables 13.7.9.4 and 13.7.9.5 show comparisons of the overall gold recovery for two

different combinations for LCO testwork program:

Table 13.7.9.4: Overall Gold Recovery for LCO Composi tes, Gravity and Gravity TailingCyanide Leaching

Sample

Au Recovery, %

Gravity Gravity Tail CN Leach Combined Au

Ag Au Ag Au Ag

Master* 58.6 7.9 88.4 58.4 95.2 61.7Ore A 39.2 7.3 94.7 36.2 96.8 40.7Ore B 47.7 12 90.9 51.3 95.2 56.9Ore C 60.8 11.6 92.1 42.2 96.9 48.7Ore D 37.7 9.8 87.8 29.0 92.4 36.0

Table 13.7.9.5: Overall Gold Recovery for LCO Composi tes, Gravity and Gravity TailingRougher Flotation and Rougher Concentrate Cyanide Leaching

Sample

Au Recovery,%

Gravity Gravity Tail Flotation Ro Conc Leach Combined

Au

Ag

Au Ag Au Ag Au AgMaster* 50.1 11.8 75.1 49.1 92.7 71.4 84.9 42.7Ore A* 45.3 9.9 69.0 38.5 92.9 64.1 80.4 32.2Ore B 38.8 11.2 80.8 55.1 91.4 79.8 84.0 50.3Ore C 57.9 4.1 68.1 45.6 94.1 82.5 84.9 40.1Ore D 46.1 11.2 76.3 29.8 89.1 70.3 82.8 29.8

Similar to the ACO program, the overall gold recovery for the LCO samples is also higher for the

gravity separation and leaching of the gravity separation tailing option. However, the cyanide





Table 13.7.9.9: Copper Extraction and Cyanide Consumption for Rougher ConcentrateLeaching of LCO

ProductRougher Concentrate CN Leach Corrected for % Wt and Cu dist .

Cu Extraction, %

CN Consumption, kg/t

Cu Extraction, %

CN Consumpt ion, kg/t

Master* 10.8

13.1

10.5

1.09

Ore A* 9.6 8.77 8.9 0.72

Ore B 12.7 13.1 12.2 1.11

Ore C 14.5 11.0 13.6 0.75

Ore D 6.6 7.93 6.4 0.52

*Average of all tests

13.8 Saprol ite Test Work Program 2012-2013

The metallurgical test program conducted by Inspectorate 2012 and 2013 focused on the extraction

and recovery of gold from saprolite. The testwork included gravity separation, cyanidation, flotation,

and cyanide detoxification.

The bulk of the work was done on a Master Composite that represents the average grade of the

material and which is derived from drillholes that are spatially represented throughout the deposit.

Approximately 150 to 200 kg was the intended sample size to ensure that sufficient material is

available in the event that repeat or confirmatory work is required.

A set of variability composites were provided that represent low, medium and high grade material

and spatial variability throughout the deposit. Samples size was about 40 to 50 kg per sample to

ensure that sufficient material is available for possible repeat or confirmatory work. All variability

samples were collected from ½ core intervals of archived drill core; most of the samples were from

drillholes completed in 2011 and 2012.

13.8.1 Gravity Separation Testwork

Both the coarse and fine saprolite samples were subject to gravity concentration using a 3 inch

Knelson® centrifugal concentrator followed by a Mozley vanner. The gravity test for the fine saprolite

sample was performed on the sample without any prior processing. The coarse saprolite sample was

ground to P80 200 µm prior to the test. Gold recoveries for the fine and coarse samples were

approximately 50% and 27%, respectively. Intensive cyanide leaching of the gravity concentrates

resulted in recoveries of 97% from both samples.

13.8.2 Flotation Testwork

Four flotation tests were performed on the fine saprolite sample to investigate different reagentschemes on the recovery of Au.

Recoveries for the fine saprolite sample were between 70% and 80% for all four tests.

Four additional tests were performed on the coarse saprolite sample to investigate the impact of

grind size on the recovery of Au. The four different P80 grind sizes used were 270 µm, 209 µm, 142

µm, and 88 µm. Results showed that a decreasing grind size had a positive effect on the recovery of

Au, with recoveries up to 86.7% for the finer grind size of 88 µm.





13.8.3 Cyanidation Testwork

Cyanide leach and CIL tests were performed on the saprolite fines sample to investigate the effect of

cyanide dosage in both leach and CIL circuits. Three separate NaCN dosages (0.25 g/L, 0.5 g/L, and

1.0 g/L) were tested on both the leach and CIL tests. Results from these test showed that NaCN

dosages have no effect on Au recovery, with all three dosages achieving the same recoveries. Theleaching method did appear to affect recoveries, with the CIL tests achieving a slightly higher

recovery of 98% versus the leach recovery of 96.8% (nominally the same).

Intermediate samples were taken from the leach tests to examine the leaching kinetics of the

sample. The tests showed that the leaching kinetics of the sample were slow, with the full 72 hours

needed to approach maximum recoveries. Figure 13.8.3.1 shows the Au leaching kinetics curve for

the sample.

Cyanidation tests were also performed on the coarse saprolite sample to investigate the grind-

recovery behavior of the sample. Four different grind sizes (P80 277 µm, 213 µm, 129 µm, and 88

µm) were tested. The tests show that the sample requires grinding to a P80 of at least 213 µm in

order to facilitate acceptable recoveries. Grinding the sample from P80 213 µm to P80 129 µm showedslight improvement to the final recovery, indicating that the optimal grind size is between P 80 213 µm

and P80 129 µm. The Au leaching kinetics curve for different grind sizes is shown in Figure 13.8.3.2.

One cyanide detoxification test was performed on the tailings from one of the CIL tests for the

saprolite fines using the air/SO2 process. Results from the test show that the air/SO2 process can

detoxify the tailings to achieve effluent standards.



0.000

0.010

0.020

0.030

0.040

0.050

0.060

0.070

0.080

0.090

0.100

0 20 40 60 80 100 120 140 160 180

R o u g h e r T a i l i n g s G r a d e

( % ,

g / t )

Sample k80 (µm)

Cu

Au

Au

Cu


Guyana

Figure 13.6.2.1

Effect of Sample P80 on RougherTailings Grade of Au and CuSource: TetraTech, 2013



20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

R e

c o v e r y ,

%

Leach Time, hours

Au Cyanide Leach Kinetics

277 um 213 um 129 um 88 um


Guyana

Figure 13.8.3.2

Au Cyan ide Leach Kinetics fo rCoarse Saproli teSource: TetraTech, 2013





Main Eastern lens, containing the larger part of the resource and displaying in its core zone

the highest average Au and Cu grades;

Main Western lens, marked by lower average gold grades and very low grades of Cu; and

SE lens, carrying mainly gold mineralization, forms a near-by satellite body, 1.2 km SE of the

main eastern lens.

On a deposit scale, relatively dense fracture networks seem to occur by preference in elongated E-W

oriented and west plunging lenticular bodies, which, in particular in the Main Eastern and the SE

Zones appear as higher grade mineralization features. Dense fracturing associated with higher grade

gold and copper mineralization seems to develop more or less along the intrusive contact and cross-

cuts lithologies. Around these higher grade core features and towards the borders of the deposit,

fracturing intensity gradually decreases and gold and copper grades drop. The Main Zone lenses are

surrounded by a larger, 2.75 km long and 200 to 400 m wide, WNW oriented mineralization envelop,

which is marked by scattered medium-to-low grade gold and barely any copper mineralization. A

similar structure, but less well expressed because of lower grades, has been detected in the Main

West part of the deposit. The objective of a 2012 re-logging exercise was to homogenize the

geological descriptions and develop a reliable geological model including the definition of geological

limits for the resource modeling.

In spite of an improved knowledge of the geological framework, it was difficult to identify clear litho-

structural boundaries for the mineralization system and limits to the mineralization have been

modeled as essentially grade “shells”. These were developed with a sequence of steps incorporating

interpreted geology (extent, shape, structures) as controls. A three dimensional grade contour

threshold was selected which produced a coherent geometry and a geologically realistic

representation of the overall mineralized extents of the deposit. Preferential orientation of the

continuity of mineralization (anisotropy) was interpreted as having a variation related to these overall

geometries. Anisotropy models were constructed and used not only as geologic controls for the

assignment of grades but also for the delineation of mineralized and non-mineralized zones internalto the overall domain wireframes.

14.3.1 External Domain Envelope

Envelopes were constructed by Sandspring to represent the overall limits of potential possible

mineralization using Leapfrog® and GEMS software and all drillhole data, the Toroparu domain

boundaries, and grade boundary interpretation from visual inspection of drillhole sections. Four

domains were created named Fresh Rock Main, Saprolite Main, Fresh Rock South East, and

Saprolite South East. The South East Zone domains were constructed separately as the zone is

geographically separated from the Main Zone by 1.2 km. These domains were created with

computer screen digitizing on drillhole sections and plan views in Gemcom by Sandspring. Theoutlines were influenced by the selection of mineralized material above 0.2 g/t Au in fresh rock and

saprolite that demonstrated zonal continuity along strike and down dip. In some cases mineralization

below 0.1 g/t Au was included for the purpose of maintaining zonal continuity. Smoothing was

utilized to remove obvious jogs and dips in the domains and incorporated a minor addition of Inferred

mineralization. This exercise allowed for easier domain creation without triangulation errors from

solids validation. On each section, polyline interpretations were digitized from drillhole to drillhole but

not typically extended more than 20 m into untested territory. Minimum constrained true width for

interpretation was 3.0 m. The interpreted polylines from each section were “wireframed” in Gemcom





into three-dimensional domains and are displayed in plan and perspective on Figures 14.3.1.1 and

14.3.1.2, respectively.

14.3.2 Anisotropy Model

A visual examination of the distribution and locations of drillhole intercepts indicates that thepreferential orientation of the continuity of mineralization (anisotropy) appears to be related to the

overall geometry of the elongated mineralization. Surfaces (digital terrain models) were created by

Sandspring from polylines following the higher grade assay occurrences throughout the entire

mineralized envelope and are displayed in red on Figure 14.3.2.1. These interpretations are used as

controls for grade estimation as discussed in section 14.10 and also acted as controls for the further

refinements of the mineralized shells internal to the overall shape.

14.3.3 Internal Domain Envelopes

With added relatively closely spaced information from the targeted infill drilling program, delineation

of non-mineralized (or extremely low grade, below 0.2 g/t Au) zones internal to the overall domain

wireframes was possible for the Main Zone and portions of the South East. Within the overall

mineralized envelope the Toroparu low grade domain boundaries were determined based on

lithology, structure and grade boundary interpretation from visual inspection of drillhole sections.

These domains were created by Sandspring with computer modeling in LeapFrog® 3D using

constant (ordinary kriging) and taking into consideration statistically calculated nugget effect for this

deposit. The outlines represent mineralized material below 0.2 g/t Au. The raw material for this

modeling exercise was selected based on filtering the existing assay database within the mineralized

envelope and creating two datasets: first above 0.2 g/t Au, and second below 0.2 g/t Au. The first

dataset was used to create internal higher grade wireframes following the orientation of the

anisotropy model and structural components containing higher grade material. The second (low

grade material) dataset was used to fill the gaps between the higher grade and the overall domain

wireframe. As a result a body that reflects the three-dimensional spreading of the low grade was

developed. Subsequently a series of volume filters were used to eliminate small bubbles and

volumes smaller than a block-size in the model. Figure 14.3.3.1 is the non-mineralized component of

the Main Zone that is internal to the overall mineralization displayed on Figure 14.3.2.1.

The Domain wireframes were imported into the Datamine Studio3® mining software package. SRK

has reviewed these wireframes and considers them appropriate; the saprolite/fresh rock boundary is

essentially a sub-topographic-parallel surface while the areal extent is a modified (or smoothed)

grade shell at a 0.2 g/t Au cut-off. Grade shells were constructed for alternative (higher) cut-offs as

well however for the purpose of a “global in-pit” resource, identifying solely a single global population

of grades, the 0.2 g/t shell was selected as reasonable and realistic. Figure 14.3.3.2 and 14.3.3.3 are

representative model plans and sections displaying the final delineation of mineralization. Internal

non mineralized units are intended to be of sufficient size to represent zones that can be differentially

mined with the anticipated methods.

14.4 Domain Analysis & Grade Capping

All assays were assigned relevant domain codes via an intersection with the wireframes. Assay

statistics are reported for both the fresh rock and saprolite mineralized domains for both Au and Cu

on Table 14.4.1.





Table 14.4.1 Assay Summary Statis tics by Domain

Item Au g/t assays Cu % assays

Al l Fresh Saprol ite Al l Fresh Saprol ite

Number of Values 42,203 39,593 2,610 42,203 39,593 2,610Maximum Value 562.600 562.600 48.600 4.658 4.658 1.862

Minimum Value 0.005 0.005 0.005 0.000 0.000 0.000Mean 0.807 0.815 0.678 0.078 0.078 0.070Variance 20.194 21.306 3.307 0.014 0.014 0.013Standard Deviation 4.494 4.616 1.818 0.118 0.118 0.113Coefficient of Variation 5.57 5.66 2.68 1.52 1.51 1.61

Source: SRK, 2013

Figures 14.4.1 and 14.4.2 are lognormal cumulative frequency (CF) distribution diagrams for Au and

Cu for the fresh rock and saprolite domains. Using these lognormal probability diagrams as a guide,

in conjunction with an examination of the distribution of drillhole data, “thresholds” were selected for

each domain type; an inflection point was selected to identify assays that are to be considered

“outliers” to the general distribution and “capped” or set back to the defined threshold. The thresholds

selected are tabulated on Table 14.4.2.

Table 14.4.2: Assay Capping Thresholds

Item Au g/t Cu %

Fresh Assay Cap 15.00 1.10Number Of Values Affected 113 29Maximum Assay Value 562.60 4.66Saprolite Au g/t Cu %

Assay Cap 8.00 0.50Number Of Values Affected 19 31Maximum Assay Value 48.60 1.86

Source: SRK, 2013

Alternative methods to the capping applied could be developed to mitigate the impact of outliers and

allow their inclusion in the assay data population; multiple populations potentially could be defined

representing different styles of mineralization. However, given the intent of modeling the primary

mineralization as a single population for the deposit, the raw assays were capped or “set back” to the

respective threshold values noted above prior to compositing. Table 14.4.3 summarizes the statistics

for capped assays; as expected there is a reduction of the coefficient of variation (CV) for both

grades within all populations. For fresh rock the CV for Au was reduced from 5.66 for uncapped to

1.86 with a decline in the standard deviation.





Wax sealing of the core was not done, and is not necessary, as the core in fresh rock is solid dense

rock with no porosity and very few open fractures.

SRK was provided an Excel Spreadsheet; Master_Bulk_Density_November_24, 2012, which has a

total of 2815 bulk density measurements for core data from various lithologies and drill depths. There

are 277 measurements that have a designation as saprolite, 213 defined as other specific rocktypes, and 2325 measurements for which the rock type is not specified.

SRK considers the amount and quality of bulk density data is sufficient for use in resource estimation

at feasibility level.

The data were examined in three ways; globally for the entire data set, by rock type, and by drill

depth. Figure 14.6.1 shows the histogram distribution of all the data, clearly indicating a tri-modal

population; saprolite (1.0 to 2.1), transition (2.1 to 2.5), and fresh rock (2.5 to 3.9). The transition data

represent in part intermediate densities of the true transition from saprolite to fresh rock, but also

includes some 161 measurements of lower density rock than the average fresh rock (2.75), yet the

data occur at depths of 100 m to 850m and are therefore not saprolite, or saprolite-fresh transition

rock.

An examination of bulk density data by the lithology codes results in the summary in Table 14.6.1

Table 14.6.1: Bulk Densit y by Major Rock types

Type No. Low High Ave Mod. Ave Comment

Saprolite 277 1.45 2.97 1.93 1.88 Ignore 1 @ 1.45 and 18 at 2.5 to 3.0

All other specifiedRock Types

213 2.25 3.92 2.77 2.76 Ignore 1 @ 2.25 and 4 above 3.3

Non-specifiedRock type

1325 1.01 3.98 2.69Presumed to be a mix of saproliteand fresh rocks of all types

Source: SRK, 2012

Individual spread sheet tabs for specific rock types present the following bulk density data Table

14.6.2.

Table 14.6.2: Bulk Densit y Data by Rock Code

Type (Rock Code) No. Low High AverageID, MD 9 2.25 2.90 2.74GB 7 2.84 3.04 2.97GRDR 26 2.70 2.80 2.73

AI 56 1.01 2.96 2.75MIV, P-MIV, FIV, F-MIV 111 2.64 3.92 2.78SAP 28 1.52 2.07 1.81

Source: SRK, 2012

SAP = Saprolite, is defined in the geological model as a separate solid shape, and GB = Gabbro,

which are dikes and are not separately modeled as the volumes are considered minimal. Fresh rock

lithologies range from 2.73 to 2.78 with the exception of Gabbro at 2.97.

In summary Table14.6.3 is a comparison of the density data currently in use in the resource block

model versus the averages as determined by the histogram and the specific lithologies.





Table 14.6.3: Summary of Bulk Density Data

Type Current Block ModelDatabase Histogram

InterpretationDatabase by

Major Rock TypeData by

Specific Rock type

Saprolite 1.84 1.85 1.88 1.81

Transition Not modeled2.30 (2.1 to 2.5),

(6-7% of the data)N/A N/A

Fresh Rock 2.76 2.75 2.76 2.77 (2.73 to 2.97)

Source: SRK, 2012

True transition rock from saprolite to fresh rock is not modeled, but is estimated at 0 to perhaps 5

meters in thickness and insignificant to the modeling, as are the gabbro dikes of nearly 3.0 density.

SRK used the two average bulk densities, 1.84 for saprolite and 2.76 for fresh rock in the resource

model; interpolation of density data was not deemed necessary.

It is estimated that 10% or less of the total volume of mineable material would be affected by using

either a mid-range density of 2.1 to 2.5 or the higher density for gabbro at 2.97, and in fresh rock

these densities will likely result in a negligible net change in tonnage

14.7 Block Model

SRK constructed block models using the Datamine Studio3® mining software package, for the

Toroparu deposit with data provided by Sandspring. The South East Zone, being approximately

1.2 km offset from the Main, was modeled separately. The models have the following characteristics

and limits.

Table 14.7.1: Toroparu Model Limits

Direction Minimum (m) Maximum (m) Blocks

Main ZoneEasting 824,200 826,800 260 ColumnsNorthing 713,800 715,700 190 RowsElevation -640 160 160 LevelsSouth East ZoneEasting 826,400 827,500 110 ColumnsNorthing 712,400 713,400 100 RowsElevation -300 140 88 Levels

Source: SRK, 2013

The block size of 10 m square in plan and 5 m vertically was considered appropriate with the

assumed mining selectivity expected for open pit mining in the area. No sub-cells (or part cells) were

used except at the saprolite/fresh rock interface and the primary resolution is to the full block size

which is adequate for the global resource model.

14.8 Search Orientation/ Anisotropy Model

Variograms, indicator variograms and correlograms were constructed for raw and composited assay

values for both Au and Cu; reasonable results were achieved and the variograms are relatively well

behaved. No preferential orientation (anisotropy) of the continuity of mineralization could be

observed but given the distribution of drilling and the geometry of the mineralization this is not un-

expected. Displayed on Figures 14.8.1 and 14.8.3 are ordinary, isotropic and anisotropic variograms





To preserve local grade variation, and to fill the volume of the mineralized domains, a search

neighborhood strategy with three search ellipse (SVOL) volumes was used (Table 14.9.2). Only

blocks not estimated with the first set of parameters were estimated with a subsequent expanded

search. In order to preserve this local variation of grades and also have a requirement for gradeassignment using data from more than one drillhole, a minimum of three 1.25 m composites was

required, with a maximum of two from any given hole, for estimation with the first two search

volumes. This results in the constraint that for “Measured & Indicated” confidence at least three 1.25

m composites from at least two drillholes are required for grade assignment.

Table 14.9.2 Search Neighborhood Strategy

SVOLSearch Distance (m) Minimum Number

Of CompositesMaximum From

One DrillholeSearch Or ien tat ion X Y Z

Au1 Dynamic 26 26 13 3 22 Dynamic 52 52 26 3 2

3 Dynamic 156 156 76 2 2Cu1 Dynamic 33 33 16 3 22 Dynamic 66 66 32 3 23 Dynamic 198 198 96 2 2

Source: SRK, 2013

Confidence classifications are initially assigned (latter smoothed as discussed in section 14.10) using

the search volume criteria and constraints of the minimum number of samples and maximum from

one drillhole. This is modified as tabulated below (14.9.3) by the absolute minimum distance to the

nearest composite. A block estimated with the second search volume that is within 10 m of a

composite is categorized as Measured and a block estimated with the third search volume that is

within 15 m of a composite is categorized as Indicated.

Table 14.9.3 Confidence Classification Scheme

Isotropic Absolute DistanceMinimum Number

Of CompositesMaximum From

One DrillholeClass SVOLMinimum Distance to

Nearest Composite

Measured 1 3 2Measured 2 10 m 3 2Indicated 2 3 2Indicated 3 15 m 2 2

Inferred 3 2 2

Source: SRK, 2013

On Figures 14.9.1 and 14.9.2 are plan views and sections that along with the perspective view of

(Figure 14.9.3) of the block model visually demonstrates the effects of the modeled anisotropy in

maintaining the mineralization continuity in preferential orientations.

14.10 Mineral Resource Classification/Confidence Assignment

For many resource models the block-by-block resource classifications should be smoothed into

geologically sensible and coherent zones that reflect a realistic level of geological and grade





estimation confidence taking into account the amount, distribution, and quality of data. A common

way of implementing this “smoothing” process is to create resource classifications based on block

estimation attributes and the broader geological and data considerations and then to adjust the

classifications of all blocks. This process includes geological rather than purely mathematical input

and is seen as an integral part of the resource classification process. Subsequent to an initial pit

optimization exercise (utilizing all blocks including Inferred) the confidence classification of all blocks

falling within the pit were examined and modifications were made to minimize the existence of

“spots” of, for example, blocks classified mathematically as Inferred that are encompassed by those

classified as Indicated, within areas with reasonable geological continuity and sufficient sampling. On

Figure 14.10.1 are representative plan views of the final modified classification with Measured and

Indicated blocks displayed in green and Inferred in red.

14.11 Resource Statement

The resource model was further investigated with a Whittle™ pit optimization to ensure a reasonable

stripping ratio was applied and a reasonable assumption of potential economic extraction could be

made. Whittle™ software was used to generate pit shells (Figure 14.11.1) for both the Main and SEZones using the operating cost inputs described in the footnote to Table 14.11.2. Incremental cut-

offs for resource reporting were produced with the economic parameters on Table 14.11.1; the

price/cut-off relationship is displayed on Figure 14.11.2.

Table 14.11.1: Resource Report ing Cut-offs

Source: SRK, 2013

Table 14.11.2 summarizes the resource for the Main, SE and both zones at a 0.30 g/t Au cut-off

within the global optimal pit designs. On Tables 14.11.3 and 14.11.4 are summary resources at

various cut-offs for all zones for Measured plus Indicated and Inferred, respectively.





Table 14.11.2: Resource Statement @ 0.30 g/t Au cut -off as o f March 31, 2013

Resource Classi ficat ion Tonnes Au Au oz Cu Cu(All rock types) (000’s) (g/t) (000’s) % (Mlb)

Main ZoneMeasured 41,542 0.98 1,307 0.109 100

Indicated 185,957 0.87 5,203 0.082 334Measured & Indicated 227,500 0.89 6,510 0.087 434Inferred 127,756 0.74 3,045 0.042 118South East ZoneMeasured 2,905 0.97 91 0.037 2Indicated 9,836 0.93 294 0.035 8Measured & Indicated 12,741 0.94 384 0.036 10Inferred 1,768 0.78 45 0.041 2 Al l ZonesMeasured 44,447 0.98 1,398 0.104 102Indicated 195,793 0.87 5,497 0.079 342Measured & Indicated 240,240 0.89 6,894 0.084 444Inferred 129,525 0.74 3,090 0.042 120

Source: SRK, 2013

1. Mineral resources are inclusive of mineral reserves;2. All resources in the revised mineral resource statement are In-Pit resources reported within an optimized pit shellabove an economic cut-off grade of 0.30 g/t Au. The economic cut-off grade was determined using a gold price ofUS$1,350/oz Au, an average metallurgical recovery of 95.9% for gold, Processing + G&A costs of US$11.49/t, andincludes US$112/oz Au for freight, smelting, refining and royalties. Copper metallurgical recovery used was 91%.Pit slopes used in the pit optimization were 45 degrees, and the mining costs used were US$2.06/t for fresh rock.

3. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is nocertainty that all or any part of the Mineral Resources estimated will be converted into Mineral Reserves;

4. Mineral Resources are reported in accordance with Canadian Securities Administrators (CSA) National Instrument43-101 (NI 43-101) and have been estimated in conformity with generally accepted Canadian Institute of Mining,Metallurgy and Petroleum (CIM) "Estimation of Mineral Resource and Mineral Reserves Best Practices" guidelines;

5. The grades for Au and Cu were estimated separately, and presented as associated average metal grades at the Aucut-off;

6. Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate, andnumbers may not add due to rounding;

7. The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there has beeninsufficient exploration to define these Inferred resources as an Indicated or Measured mineral resource and it is

uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category;and

8. The mineral resource estimate for the Project was calculated by Frank Daviess, MAusIMM, R.M. SME, AssociateResource Geologist of SRK Consulting, Inc. in accordance with the Canadian Securities Administrators NationalInstrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and generally accepted CanadianInstitute of Mining, Metallurgical and Petroleum “Estimation of Mineral Resource and Mineral Reserves BestPractices” guidelines (“CIM Guidelines”).





14.12.2 Comparative Statistics

On Tables 14.12.2.1 and 14.12.2.2 are comparative statistics for the grades of model blocks (at a

zero cut-off within the shells) and the composited assay values within the relevant shells.

The average estimated Au and Cu grades of the resource model blocks are marginally lower than

the average grades of the composited values used for the estimation for each of the areas. The

coefficient of variation, already relatively low for the composites, is lower for each of the modeled

areas. In general the model is a “smoothed” representation of the composited data and is adequate

for global resource estimation.

Table 14.12.2.1: Fresh Rock Composi te/Model Statist ics

Population Au g/t Cu %

Model Assay Model AssayBlocks Composites Blocks Composites

Maximum Value 12.90 15.00 0.96 1.10Minimum Value 0.00 0.00 0.00 0.00Mean 0.71 0.76 0.06 0.07

Standard Deviation 0.654 1.235 0.067 0.102Coefficient of Variation 0.92 1.62 1.10 1.47

Source: SRK, 2013

Table 14.12.2.2: Saprolite Compos ite/Model Statisti cs

Population

Au g/t Cu %

Model Assay Model AssayBlocks Compos ites Blocks Compos ites

Maximum Value 5.52 8.00 0.46 0.50Minimum Value 0.00 0.00 0.00 0.00Mean 0.61 0.65 0.06 0.06Standard Deviation 0.537 0.938 0.071 0.080

Coefficient of Variation 0.89 1.43 1.20 1.31

Source: SRK, 2013

For comparative purposes the resource block model was assigned grades using alternative

methodologies to the ordinary kriging scheme (OK) with dynamic anisotropies that is reported above.

These include inverse to the distance squared (ID2) and nearest neighbor (NN). ID2 and OK methods

were used both with the dynamic anisotropy option and with an isotropic (uniform spherical) search

ellipsoid. On Table 14.12.2.3 for Au and Table 14.12.2.4 for Cu, are Measured and Indicated

resources for fresh rock at a zero cut-off estimated with alternative methods. At the zero cut-offs1, for

all cases, the choice of estimators does not have a major impact on the global inventory tonnages or

grade. As expected nearest neighbor (NN) result in the highest estimated grade as there is little

smoothing with this method; differences are minimal. The effects of variable anisotropy are seen as

the increase of grade from 0.542 (Isotropic) to 0.555 (anisotropic); this “preservation of metal” was

the intended outcome of the procedure.

1 Note: a zero cut-off is for comparative purposes only for the entire mineralized datasets, and does not relate to a

resource. Resources cannot be reported at a zero cut-off grade.




Guyana

Figure 14.3.1

Toroparu PlanSource: SRK, 2013




Guyana

Figure 14.3.1.1

Toroparu Plan “ Fresh Rock” (grey) “ Saprolite” (orange) DomainsSource: SRK, 2013




Guyana

Figure 14.3.1.2

Toroparu Perspective “ Fresh Rock(grey) & “ Saproli te” (orange)

DomainsSource: SRK, 2013




Guyana

Figure 14.3.2.1

Toroparu Perspectives “ Fresh Rock(grey) & “ Anisotropy “ (red)Source: SRK, 2013




Guyana

Figure 14.3.3.3

Toroparu Model Cross Sections,Mineralized/non-Mineralized

(green/grey)Source: SRK, 2013




Guyana

Figure 14.8.1

Variogram, Au (g/t) Modeled Anisotropic & Isotrop ic VariogramSource: SRK, 2013




Guyana

Figure 14.8.2

Variogram, Au (g/t) Modeled Anisotrop ic VariogramSource: Company, Year




Guyana

Figure 14.8.4

Variogram, Cu (%) Modeled Anisot ropic Var iogramSource: SRK, 2013




Guyana

Figure 14.9.1

Resource Model Plans160 Elevation -250 ElevationSource: SRK, 2013




Guyana

Figure 14.9.2

Resource Model RepresentativeCross SectionsSource: SRK, 2013




Guyana

Figure 14.9.3

Resource Model PerspectiveSource: Company, Year




Guyana

Figure 14.10.1

Resource Model ConfidenceClassif ication -225, -200,

-100 Plan ViewsSource: SRK, 2013




Guyana

Figure 14.11.1

Optimized Resource PitsSource: SRK, 2013



-

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

$1,070 $1,250 $1,300 $1,350 $1,400 $1,450 $1,500 $1,550 $1,600 $2,030

A u

G / t

Au cut off grade (g/t)


Guyana

Figure 14.11.2

Cut-off/PriceSource: SRK, 2013




Guyana

Figure 14.12.3.2

North-South Swath DiagramSource: SRK, 2013

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

XC 825400 826400

AUKRG

AUNN




Guyana

Figure 14.12.3.3

East-West Swath DiagramSource: SRK, 2013

0

0.5

1

1.5

2

2.5

3

3.5

YC 714890

AUKRG

AUNN





15 Mineral Reserve Estimate

15.1 Conversion Assumptions, Parameters and Methods

All conversion assumptions such as model dilution, mining recovery, cut-off grade calculation, pitoptimization and costs were taking into consideration to calculate the reserve estimate.

The following steps were used to calculate the reserves:

1. Apply mining dilution to resource block model (using 3D techniques);

2. Gather costs and process recoveries;

3. Input optimization parameters into pit optimizer to calculate nested pits using different gold

selling prices (only Measured and Indicated resources were included as ore);

4. Chose pit optimization shell based on strip ratio, revenue, grade distribution, discounted

cash flow, cash costs, equipment selection sizes, pit footprint, depth of pit, minimum mining

widths, cut-off grade, processing plant size and many other factors;

5. Detailed phase design with ramp access to all benches;

6. Multiple trade off mine plans based on different processing rates (quarterly periods for five

years and yearly until the end of the mine life);

7. Detailed truck haulage estimates;

8. Detailed mine cost estimates based on detailed mine plan;

9. Discounted cash flow based on all capital and operation cost inputs; and

10. Choose final mine plan and cash flow followed by reported reserves.

There follows a description of how reserve dilution was applied and how the in-pit cut-off grade was

calculated.

Model Grade Dilution

The mineralized ore body shell was developed by using two gold cut-offs (0.1 g/t and 0.2 g/t shells).

SRK used the 0.2 g/t shell for grade estimation while the 0.1 g/t shell was used to calculate the

dilution outside of the 0.2 g/t shell. SRK calculated the dilution with the following method:

0.2 g/t Au shell and internal low grade dykes triangulations were used to calculate the

percentage of the block inside of the triangulations;

Because all material within the block model block were within the 0.1 g/t Au shell, SRK

applied a 0.15 g/t Au dilution grade for gold and 0.015% Cu to the volume outside of the 0.2g/t Au shell. This output was then used to calculate on a block by block basis the new 3D

dilution. Blocks that did not touch the 0.2 g/t Au shell or the dykes were not affected; and

This is a purely mathematical manipulation and there was no involvement of lithology or

other geologic criteria. SRK calculation shows a drop of 2% of gold grades and close to 3%

by mass.




Guyana

Figure 15.1.1

Cut-off Grade Calculation Graph(by different Gold selling price)Source: SRK, 2013





16 Mining MethodsMore accurate resource and geologic models produced over the course of 2011/2012 during the

prefeasibility definition drilling campaigns identified two geographically distinct populations of gold

bearing saprolite and fresh rock ores, distinguishable by their copper sulfide contents, ore withrecoverable copper being defined as “Au/Cu Ore”, and without recoverable copper content as “Au

Ore”. The mine plan and production schedule defined in the PFS were optimized for higher

metallurgical recovery by processing these ores separately in different circuits as they are mined

from the Toroparu and SE pits. During the metallurgical testwork program these two ore types were

given the names of ACO (Average Copper Ore) for the “Au/Cu Ore” and LCO (Low Copper Ore for

the “Au Ore”.

The PFS mine plan provides for the excavation of 127.1 Mt of ore (containing 4.107 Moz of gold at

an average grade of 1.00 g/t Au) and 468.9 Mt of waste for a combined total of 596.0 Mt of material

at a life of mine (LoM) stripping ratio of 3.69:1. Mining will be conducted with conventional open pit

mining techniques over a 16-year mine life in two pits, the Toroparu Pit which will be mined in 13

phases, and the nearby South-East Pit (1.2 km to southeast of Toroparu), which will be mined in four

phases. The mine plan includes:

5 Mt of saprolite Au Ore containing 148,000 oz of gold at an average grade of 0.91 g/t Au

that will be processed via conventional cyanide leach;

52 Mt of fresh rock Au/Cu Ore containing 1,953,000 oz of gold with an average grade of

1.17g/t Au and 0.18% Cu that will be processed via flotation concentration; and

70 Mt of fresh rock Au Ore containing 2,006 Moz of gold with an average grade of 0.89 g/t

Au and 0.05% Cu that will also be processed via cyanide leach.

Processing facilities will be developed in three phases:

Pre-production is designed to process 1.18 Mt/y (3,250 t/d) of saprolite Au Ore during that

period;

Phase 1 is designed to process 5.475 Mt/y (15,000 t/d) of fresh rock Au/Cu Ore via flotation

concentration, and a combination of 1,500 t/d of saprolite Au Ore and fresh rock 2,600 t/d of

Au/Cu Ore flotation tailings over the first three years of production; and

Phase 2 is designed to treat 5.745 Mt/y (15,000 t/d) of fresh rock Au Ore, saprolite Au Ore

and Au/Cu Ore flotation tailings via cyanide leach and 2.738 Mt/y (7,500 t/d) of Au/Cu Ore

via flotation concentrate starting in the fourth year of production and continuing for the

balance of the mine life.

Mining operations are planned to commence during the second year of construction of the Project in

the center of the Toroparu pit, with mining of saprolite Au Ore to support pre-production of gold in

saprolite Au Ore processing. The following year mining will be expanded to include mining and

stockpiling of fresh rock ore to support the start-up of fresh rock Au/Cu Ore processing in the first

year of production. Total mining during the two pre-production years is estimated at 15.2 Mt with a

stripping ratio of 1.05:1. From this point, the mine plan calls for 14 years of mining out of a total 16

year production life. The first five years of mining (after pre-production) will continue in the center of

the Toroparu Pit at a rate of 40 Mt/y utilizing small scale mining equipment fleet based on 50 t

capacity haul trucks with a stripping ratio of 3.29:1. Mined fresh rock Au/Cu Ore will be processed in

the concentrate circuit, and fresh rock Au Ore will be stockpiled for later feed into the expanded





cyanide leach circuit beginning in Year 4 (Phase 2). Loading operations will be accomplished with a

fleet of hydraulic excavators (to enhance ore selectivity) and hydraulic shovel, with operational

flexibility provided by a wheel loader.

In the sixth year of the mine plan the annual mining rate is expected to be ramped up to an average

of 50 Mt/y and the main truck fleet will be switched over to 133 t capacity haul trucks to support theexpanded processing capacity. In Year 13, the annual mining rate will reduce to 21 Mt/y. In Year 14,

open pit operations will be completed for the mineral reserves defined in the PFS mine plan. The

processing plant will continue operating for another two years processing from low grade ore

stockpiles.

16.1 Proposed Mining Method

For the PFS, it is assumed that a conventional open pit operation including drilling and blasting,

loading and hauling.

Drilling and blasting are planned to be performed on 10 m benches in both pits. This matches a

multiple of the block size in the geological block model. Due to the expected selective mining that willbe required for ore mining, loading and hauling are planned to be performed using a half-bench

height for ore, and a full bench heights for waste handling.

The Toroparu pit is planned to be developed first, with the process facility to be constructed adjacent

to this pit. This will minimize the ore haulage requirements during the early years of the Project.

The Project plans to use proven technology, with no requirement for untried or untested technology.

16.2 Geotechnical Mine Design Parameters

Knight Piésold Ltd. (KP) completed a geotechnical site investigation program and pit slope design for

the Toroparu Project. The Project site includes a 30 to 40 m thick sequence of surficial saprolite and

a massive volcanic and metasedimentary rock assemblage. Two major geotechnical domains,

saprolite and Fresh Bedrock, were defined for the pit slope geotechnical assessment.

Five major pit design sectors, namely, Northeast, East, South, Southwest, and Northwest, were

defined for the proposed Toroparu Pit based on the orientations of pit walls and the wall geology.

Sub sectors were also delineated to differentiate the saprolite and Fresh Bedrock domains in each

sector.

It is assumed that moderately sized mining equipment will be used for the Toroparu open pit mining

operations, which allows for a 10 m high single bench to be developed in the pit walls.

The inter-ramp slope angles are typically determined by the bench geometry. A shallow slope angle

of 38 degrees is recommended in the upper saprolite slopes. A steeper inter-ramp slope angle of 53

degrees is recommended for areas with less kinematic controls (South, Southwest, and Northwest

Sectors), while a flatter inter-ramp angle of 50 degrees is appropriate for the East Sector where the

potential for minor planar failure has been identified. The Northeast Sector is limited to an inter-ramp

angle of 45 degrees due to adverse planar structures.

A 16 to 20 m wide catch bench is recommended along the saprolite/Bedrock contact, to intersect

surface run off water and provide additional containment capacity for potential saprolite raveling





16.3.4 Optimization Parameters

The pit optimizations have been carried out using Whittle™ optimization software (Whittle™ Version

4.4). Revenue, mining costs, processing values and other factors as described in this section were

input to the Whittle™ software.

Mining Dilution

The block model as imported into Whittle™ was diluted. The optimization process included factors of

0% mining dilution and 100% ore recovery (as this was pre-coded into the block model). These

parameters were supplied by the client but considered by SRK to be reasonable.

Discount Rate

The pit optimization process did not utilize a discounting factor. Inflation was not factored into the

costs, which represent an indication of the “Current Prices” in the analysis.

The Lerchs-Grossmann algorithm (on which the Whittle™ software is based) produces a series of

mathematically optimum pit shells directly linked to the Revenue Factor utilized if the maximum

undiscounted cash flow is the selection criterion for optimization.

Geotechnical Parameters

For the Toroparu optimization, three geotechnical domains were utilized. For the upper Saprolitic

zones, an overall wall angle of 28 degrees was used. Below the saprolite, in the North-East corner of

the proposed pit, an overall wall angle of 38 degree was used. For all other areas of the pit, an

overall wall angle of 45 degrees was used. These parameters are much shallower than the proposed

inter-ramp to account for ramp systems within the pit optimization runs.

For the South-East Pit optimization, two geotechnical zones were utilized. For the upper Saprolitic

zone, an overall wall angle of 28 degrees was used. Below the saprolite, an overall wall angle of 40

degrees was used. The overall wall angle includes any allowances for ramps within each wall.

Royalties

Royalties have been defined by the Sandspring. Royalties to a total of 8% for gold sales and 1.5%

for copper sales have been applied.

Mining Costs

SRK reviewed the proposed costs and modified the input values based on prior experience with

similar projects. SRK has not applied an incremental cost to account for the increased cost of mining

at depth.

Material has been classified either as saprolite or fresh rock, and a unique cost per tonne has been

applied for each material type. For saprolite, the cost per tonne is US$1.54/t and for Fresh (or

Sulfide) Rock, the cost per tonne is US$1.86/t.

Processing Costs and Recoveries

The estimated processing costs for both deposits were supplied by Tt. Three processing methods

have been identified with unique costs for each stream. The processing costs are defined in Table

16.3.4.1.





Table 16.3.4.1: Processing Parameters

Material Type Processing Method Unit Value Recovery

Saprolite (SAP) Carbon in Pulp (CIP) US$/t ROM 7.713 Au – 96.06%

ACO (Au/Cu Ore) Flotation / CIP / Intense Leach US$/t ROM 11.31 Au – 85.46%

Cu – 91%

LCO (Au Ore) CIP / Intense Leach US$/t ROM 10.12 Au – 95.90% Source: SRK

Other Costs

Due to the processing methods, a series of other costs also require inclusion in the optimization.

Table 16.3.4.2 summarizes the optimization parameters used.

Table 16.3.4.2: Optimization Parameters (Base Case)

Parameter Unit Value

Mining Dilution % 0Mining Dilution Grade 0.00

Mining Recovery % 100Toroparu Overall Slope Angle (Above Saprolite / Fresh interface) (

o) 28

Toroparu Overall Slope Angle (Below Saprolite / Fresh interface) (o) 38

Toroparu Overall Slope Angle (Below Saprolite / Fresh interface) (o) 45

South East Overall Slope Angle (Above Saprolite / Fresh interface) (o) 28

South East Overall Slope Angle (Below Saprolite / Fresh interface) (o) 40

Mining Cost (Saprolite Material) US$ / t 1.54Mining Cost (Fresh Material) US$ / t 1.86Mining Rate Mt/y 60Processing Rate (Saprolite) Mt/y 2.0Processing Rate (ACO) Mt/y 15.0Processing Rate (LCO) Mt/y 15.0Process Recovery Au (SAP) % 96.06Process Recovery Au (ACO) % 85.46

Process Recovery Cu (ACO) % 91.00Process Recovery Au (LCO) % 95.90Processing Costs (SAP) US$ / t ore 7.713Processing Costs (ACO) US$ / t ore 11.31 Processing Costs (LCO) US$ / t ore 10.12 General and Administration US$ / t ore 1.37 Sustaining Capital Cost US$ / t ore 0.65Gold (Au) Price US$ / oz 1,400Gold Royalty US$ / oz 112Copper Royalty US$ / lb 0.051Doré Au NSR Deductions / Losses % of Au Sales 0.1Doré Au NSR Transport and Insurance US$ / oz 2.45Doré Au NSR Refining Charges US$ / oz 0.65Cu Concentrate Au NSR Deductions / Losses % of Au Sales 3.0Cu Concentrate Au NSR Smelting and Refining US$ / oz 6.5Cu Concentrate Cu NSR Deductions / Losses % of Cu Sales 4.76Cu Concentrate Cu NSR Treatment and Refining US$/ lb 0.28Freight and Marketing US$ / lb 0.34

Source: SRK

16.3.5 Optimization Process

To optimize both deposits, a series of nested pit shells were calculated over a range of Revenue

Factors (RFs). Each of the nested pit shells were generated based on the maximum undiscounted





Geotechnical Parameters

Table 16.4.2 shows the geotechnical parameters used for the Preliminary Economic Assessment

(PEA) pit design. Due to no changes in the geotechnical parameters, SRK applied the same inter-

ramp angles to the new pit design. Table 16.4.3 shows the pit geotechnical parameters for the

saprolite and fresh rock. SRK used the PEA geotechnical parameters and applied smoothingbetween the proposed inter-ramp angles to ensure that smooth sector transitions.

Table 16.4.2: Toroparu Pit Final Geotech Pit Design Parameters Used in Updated PEA

RocktypePit DesignSector

Pit WallOrientation

(º)

KinematicFailure

Mode

BenchFace

Angle (º)

BenchHeight

(m)

BenchWidth

(m)

Inter-ramp

Angle (º)Saprolite - - 65 12 10 38

Fresh Bedrock

Northeast 220 Planar 65 24 13 45East 250 Planar 70 24 11.5 50South 355 Toppling 75 24 11.5 53

Southwest 40 Toppling 75 24 11.5 53Northwest 120 75 24 11.5 53

Source: Knight Piésold*The geotech parameters above does not include the South East pit area. SRK used 45 degrees inter-ramp angles for thesouth-east pit design.

Table 16.4.3: Toroparu and South -East Pits Geotech Pit Design Parameters

Pit RocktypeStart

AzimuthEnd

AzimuthBerm

Width (m)Batter

Angle (º)BenchHeight

DoubleBench

Inter-ramp Angle (º)

Toroparu Fresh

0 5 10.00 70 20 YES 495 10 11.25 70 20 YES 47

10 15 11.75 70 20 YES 4715 20 12.25 70 20 YES 4610 45 12.75 70 20 YES 45

45 48 12.25 70 20 YES 46

45 50 11.75 70 20 YES 4750 55 10.75 70 20 YES 4855 60 10.00 70 20 YES 4960 115 9.50 70 20 YES 50

115 120 9.00 70 20 YES 51120 125 8.50 70 20 YES 52125 130 8.20 70 20 YES 52130 345 7.80 70 20 YES 53

345 350 9.00 70 20 YES 51350 355 9.25 70 20 YES 50355 0 9.75 70 20 YES 50

Toroparu Saprolite 0 0 9.25 70 10 NO 37South-East Fresh 0 0 12.75 70 20 YES 45South-East Saprolite 0 0 9.25 70 10 NO 37

Source: SRK

Figure 16.4.1 shows the Toroparu pit geotechnical sectors used to design the phase designs and

final pit design. Blending sectors were added to ensure smooth transition between 45 to 53 degrees

sectors. Saprolite material was excluded from this methodology.





Table 16.6.1.2 shows the planned mill feed schedule.

Table 16.6.1.2: Production Mill Schedule (Mill Feed)

Years

Saproli te Oxide Ore Fresh Au/Cu Ore (ACO) Fresh Au Ore (LCO)

OreProcessed

(kt)

Mill Au Grade

(g/t)

Contained Gold

(koz)

OreProcessed

(kt)

Mill Au Grade

(g/t)

Contained

Gold (koz)

Ore Processed

(kt)

Mill Au Grade

(g/t)

ContainedGold

(koz)

OProcess

Year -2 1,186 1.25 48 0 0.00 0 0 0.00 0 1,1

Year -1 1,186 0.95 36 0 0.00 0 0 0.00 0 1,1

Year 1 548 0.74 13 5,475 1.74 306 0 0.00 0 6,0

Year 2 548 0.61 11 5,475 1.34 236 0 0.00 0 6,0

Year 3 517 0.65 11 5,475 1.24 218 0 0.00 0 5,9

Year 4 64 1.89 4 2,738 0.98 87 5,475 1.32 232 8,2

Year 5 64 1.54 3 2,738 1.63 143 5,475 0.78 138 8,2

Year 6 64 1.85 4 2,738 1.01 89 5,475 0.77 135 8,2

Year 7 64 0.92 2 2,738 1.63 143 5,475 0.77 136 8,2

Year 8 64 1.48 3 2,738 1.73 153 5,475 1.31 231 8,2

Year 9 64 0.76 2 2,738 0.80 70 5,475 0.71 125 8,2

Year 10 64 0.77 2 2,738 1.03 91 5,475 1.13 199 8,2

Year 11 64 0.80 2 2,738 1.32 116 5,475 1.39 244 8,2

Year 12 64 0.76 2 2,738 0.51 45 5,475 0.78 137 8,2Year 13 64 0.48 1 2,738 0.98 86 5,475 0.66 115 8,2

Year 14 64 0.48 1 2,738 0.99 87 5,475 0.89 157 8,2

Year 15 64 0.48 1 2,738 0.49 43 5,475 0.48 85 8,2

Year 16 271 0.48 4 2,505 0.49 40 4,609 0.48 71 7,3

Totals 5,022 0.91 148 51,780 1.17 1,953 70,309 0.89 2,006 127,1

Source: SRK*ACO terminology: Material where the copper content is above 0.09%. – Material subject to Flotation – a.k.a Au/Cu Ore*LCO terminology: Material where the copper content is below 0.09%.– Material subject to Cyanide Leaching – a.k.a Au Ore





16.8 Mining Equipment Requirements

16.8.1 Summary

Mining methods will be open pit mining using hydraulic excavators, shovels and wheel loaders loading

haul trucks for waste and ore haulage. The operations are described further in the following sections.Mining activities will include removal of any growth medium (topsoil), free-digging, drilling, blasting,

loading, hauling and mining support activities. Material within the pit will be generally blasted on a 10

high m bench, but it is planned for ore to be mined by excavators in two 5 m flitches (lifts). Saprolite

material (approximately 12% of the total material to be mined) can be loaded directly with hydraulic

excavators without the need for blasting. Waste dumps will be used for material below the cut-off

grade, and stockpile for ore above the cut-off grade. The stripped waste material will be placed in

dumps, either to the north or east of the Toroparu Pit, and lower-grade ore placed in a stockpile, near

to the primary crusher location. Some ore will be sent directly to the primary crusher.

16.8.2 General Parameters and Fleet Selection

Specific requirements dictated the selection of mining equipment types and sizes. Loading equipment

selection focused on generally having diesel-powered hydraulic excavators (backhoes), together with

a front end loader available for added operational flexibility. A hydraulic shovel was also included in a

larger equipment fleet placed into service in Year 6 of full production.

Hydraulic excavators will be primarily used for loading in the open pits (Toroparu and SE pits) and the

front end loader for loading in the low grade stockpile. Trucks will be matched to the loading

equipment units. Additional equipment units were provisioned when required, in keeping with the

planned mine production schedule requirements.

The major mine equipment fleet requirements were based on the annual mine production schedule,

the mine work schedule, and shift production estimates. The mine equipment requirements andcosting were based on the purchase of new equipment. The equipment fleet selection and

requirements are further discussed in the individual sections that follow in this report.

It was planned that all mine mobile equipment would be diesel-powered, in order to avoid the

requirement to provide electrical power into the pit working areas.

The mine operations schedule is proposed to include two twelve-hour shifts per day, seven days per

week for 355 days per year, which includes an annual allowance of 10 days downtime for weather

delays for most of the mine operations, and 15 days downtime for weather delays for the drilling

operations. Mine productivity and costing included estimating the productive operating time per

twelve-hour shift. Non-productive time per shift includes shift change (travel time), equipment

inspections, fueling, and operator breaks. It was estimated that the total time per shift for these itemswill be 1.68 hours. The scheduled production time (scheduled operating hours) was therefore

estimated at 10.32 hours per shift, representing a (shift) utilization of 86% of the twelve-hour shift

period (and excludes mechanical availability and work efficiency factors).

In addition, allowances were made for work efficiencies including equipment moves (production

delays while moving to other mining areas within the pit), and certain dynamic operational

inefficiencies. These work efficiencies are further discussed in the respective sections for drilling,

loading and hauling.





Equipment fleet mechanical availability was estimated for the various major mine equipment fleets,

including drills, hydraulic excavators, hydraulic shovels, front end loaders, trucks, etc. (with

replacement equipment units assumed to be new). For the large haul truck fleet annual adjustments

were made depending on the age (in terms of operating hours) of the equipment as averaged over the

entire fleet. Table 16.8.2.1 shows the estimated mechanical availabilities used, and the planned

equipment life for each equipment type.

Table 16.8.2.1: Mechanical Availabil ities and Planned Equipment Life

Op Hr Intervals0 -

5,0005,000 -10,000

10,000 -15,000

15,000 - 30,000 - Over Unit

30,000 50,000 50,000 Life

Mech. Mech. Mech. Mech. Mech. Mech. Oper

Major Mining Equipment Avail. Avail. Avail. Avail. Avail. Avail. Hrs

Type (%) (%) (%) (%) (%) (%) (op hrs )

Blasthole drill - AC PV-235 75 75 75 75 75 - 50,000Control blasthole drill – AC D65 75 75 75 75 75 - 50,000Front end loader - Cat 88H 80 80 80 80 80 - 50,000Front end loader - Cat 993K 80 80 80 80 80 80 60,000Hydraulic excavator - Cat 390 85 85 85 85 85 - 40,000Hydraulic excavator - Cat 6018 85 85 85 85 85 85 80,000Hydraulic excavator - Cat 6040 85 85 85 85 85 85 90,000Hydraulic shovel - Cat 6040 85 85 85 85 85 85 90,000Haul truck - Cat 740B 85 85 85 85 85 - 50,000Haul truck - Scania 10x4 85 85 85 85 - - 18,000Haul truck - Cat 785D 92 90 88 84 80 75 75,000Track dozer - Cat D9T 75 75 75 75 75 75 60,000Wheel dozer - Cat 844H 75 75 75 75 75 75 60,000Motor grader - Cat 16M 75 75 75 75 75 - 50,000Water truck - Scania 8x4 30kL 75 75 75 75 - - 20,000Excavator - Cat 374DL 75 75 75 75 - - 30,000Compactor - Cat CS/CP-54 75 75 75 75 - - 30,000

Source: SRK

Table 16.8.2.2 shows the mining equipment requirements for selected years of the mine plan.





Table 16.8.2.2: Planned Mining Equipment Requirements for Selected Years

Equipment Units Used Make Model Size -2 -1 1 2 4 6 8 10 12 14

Drilling

Blasthole drill - new Atlas Copco PV-235-D 165 mm 0.6 2.8 3.0 3.2 3.2 4.3 3.9 4.3 0.7

Control blasting drill Atlas Copco SROC D65 110 mm 1 1 1 1 1 1 1 1 1

LoadingFront end loader Caterpillar 988H 6.4 m3 1.4 0.6 1.3 1.8 2.1 1.8 0.6 1.1 1.4 1.2

Front end loader Caterpillar 993K 12.2 m3 0.7 0.6 0.9 0.1

Hydraulic excavator Caterpillar 390DL 5.0 m3 1.7 2.6 7.7 7.5 7.2 6.9 3.4

Hydraulic exc/shovel Caterpillar 6018EX/FS 10.0 m3 0.9 2.9 2.9 2.9 3.0 2.9 2.0 2.8

Hydraulic exc/shovel Caterpillar 6040EX/FS 22.0 m3 0.8 2.0 1.9 1.9 0.6

Hauling

Haul truck - new Caterpillar 740B 40 t 0.2 3.7 3.7 3.7 3.6 3.2 2.4 1.4 2.0 1.3

Haul truck - new Scania G460CB 10X4 50 t 3.6 8.5 42.1 47.6 57.9 57.5 36.3 19.9 19.8

Haul truck - new Caterpillar 785D 133 t 6.8 24.5 24.8 24.7 12.2

Other Mine Equipment

Crush/Screen Plant Manufacturer Jaw/Cone 335 kW 1 1 1 1 1 1 1 1 1 1

Track dozer - new Caterpillar D9T 306 kW 3 4 4 4 4 5 5 5 5 3

Wheel dozer - new Caterpillar 844H 468 kW 0 2 2 2 2 3 3 3 3 2

Motor grader - new Caterpillar 16M 221 kW 2 3 3 3 3 3 3 3 3 3

Backhoe loader Caterpillar 450E 102 kW 1 1 1 1 1 1 1 1 1 1

Water truck - new Scania P410CB 8X4 30,000L 1 2 2 2 2 2 2 2 2 2

Excavator - new Caterpillar 374DL 355 kW 1 2 2 2 2 3 3 3 3 2

Compactor - new Caterpillar CS/CP 54 97 kW 2 2 2 2 2 2 2 2 2 2

Support Equipment

Transport/mover Manufacturer Model 360 t 1 1 1 1 1 1 1 1 1 1

Truck crane Manufacturer Model 120 t crane 1 1 1 1 1 1 1 1 1 1

Recovery truck Scania G460CB 8X8 360 kW 1 1 1 1 1 1 1 1 1 1

Secondary blast drill Manufacturer 75 kW 64 mm 1 1 1 1 1 1 1 1 1 1

Fuel/lube truck Scania P410CB 8X4 30,000 L 1 1 1 1 1 2 2 2 2 1

HD mechanic's truck Scania P360CB 6X4 1 1 2 2 2 2 2 2 2 2

Flatbed truck Scania P360CB 6X4 19 t crane 1 1 1 1 1 1 1 1 1 1

Welding truck Manufacturer Model 1 1 1 1 1 1 1 1 1 1Tire service truck Scania P360CB 6X4 1 1 1 1 1 1 1 1 1 1

Forklift Manufacturer Model 1 1 1 1 1 1 1 1 1 1

Pit pumps/generators Flygt/Gen BS2290-434 82 kW 4 4 4 4 8 8 8 8 8 4

Pit pumps & engines Godwin/Cat HL260M/C18 430 kW 2 2 3 6 6 7 7 8 5

Personnel van/bus Manufacturer Model 5 5 5 5 5 5 5 5 5 5

Service pickup Manufacturer 4x4 5 15 15 15 15 15 15 15 15 15

Light plant Manufacturer Portable 8 kW 10 10 10 20 20 20 20 20 20 20

Blasting

Blasting flatbed truck Scania G360CB 4X4 1 1 1 1 1 1 1 1 1

ANFO/Emulsion truck Scania P360CB 6X4 13 t 1 1 1 1 1 1 1 1 1

Blasters crew truck Manufacturer 4x4 1 1 1 1 1 1 1 1 1

Blasthole stem truck Scania P360CB 6X4 1 1 1 1 1 1 1 1 1

Source: SRK

16.8.3 Dril ling

The planned drilling equipment fleet will consist of Atlas Copco PV-235 units. This fleet was based on

drilling 165 mm blastholes to an average depth of 11.5 m (including a 1.5 m sub-drill) for development

of 10 m high benches. The drills can single-pass drill (no rod changes) such holes.





The planned nominal production blasthole pattern is equivalent to a 5.9 m x 5.9 m pattern (spacing

and burden) in waste and ore, however, in practice the burden and spacing will vary. (The planned

nominal 5.9 m square pattern would be approximately equivalent to a 5 m x 7 m pattern.) For the main

production drilling an instantaneous drilling rate of 0.65 m/minute was estimated for waste and ore.

Allowances were made in the drilling productivity estimates for re-drills (5%) and moving to new

working areas. Fleet requirements were based on drilling all of the fresh rock within the planned openpits (88% of material).

Allowance was also made to have a dedicated drill (for drilling wall control blasting blastholes and

other specialized drilling situations. Table 16.8.3.1 shows selected drilling statistics based on the

planned drilling equipment and drilling patterns for waste and ore.

Table 16.8.3.1: Drilling Statistics Per Unit

Item Unit Value

Rock Type Waste & OreWaste/Ore Pattern Size m x m 5.9 x 5.9Drilling Tram and Set Up Time min/op hr 8.3Drilling Penetration Rate m/min 0.65

Drilling Time per Blasthole min 17.7Moving and Delay Time min/op hr 10Production per Unit (100% Available) * t/op hr 2,144

Source: SRK*Includes allowance of 5% for re-drills.

Table 16.8.3.2 shows selected drilling productivity information based on the planned drilling

equipment. Annual production capacity for per drill is 11.6 Mt/y.

Table 16.8.3.2: Drilling Product ivit y Per Unit

Item Unit Value

Rock Type Waste & OreProduction per Unit (100% Available) t/op hr 2,144Planned Operating Hours per Shift scheduled op hrs 10.32Planned Operating Hrs per Year* scheduled op hrs 7,224Estimated Mechanical Availability** % 75%

Actual Operating Hours per Year op hrs 5,418 Annual Production Capacity per Unit Mt/y 11.6

Source: SRK*Includes allowance of 15 days downtime for weather delays.**Typical mechanical availabilities for drills used.

16.8.4 Blasting

Bulk emulsion explosives will be used for blastholes. Blasting requirements were based on blasting allfresh rock within the planned open pits (88% of material).

The powder factor for production blasting was estimated to be 0.233 kg/t (kg explosives per tonne of

rock), based on an estimate by Orica Mining Services. As previously mentioned, a 5% contingency

allowance was made for additional blasthole drilling (closer drilling to achieve proper fragmentation),

and this contingency also includes the necessary explosives.





The explosives provider for the mine will have a dedicated bulk emulsion plant, which will be capable

of sufficient production for the planned mining operations. Blasting accessories will be transported to

site and stored in suitable explosives magazines.

The mine will have a 13 t emulsion truck, which will deliver bulk explosives to the blast sites during

daylight hours. The blasting equipment fleet will include a dedicated stemming truck, a flatbed truck

and blasting crew truck. Stemming material will be mainly drill cuttings. The mine blasting crew will

manage and conduct the blasting operations.

16.8.5 Loading

Loading equipment selection included having a combination of diesel-powered hydraulic excavators,

hydraulic shovels, and front end loaders for operational flexibility. The hydraulic excavators are

capable of mining more selectively, and will be used for mining most of the ore, and part of the waste.

The hydraulic shovels will be primarily used for loading waste. The front end loaders (6.4 m 3

Caterpillar 988H class) will be used for stockpile re-handling loading duties.

The loading equipment fleet for the earlier years of the mining operations was planned to be a

combination of equipment consisting of up to eight smaller hydraulic excavators (5.0 m 3 Caterpillar

390 DL class, Excav1), up to two medium size hydraulic excavators (10.0 m3 Caterpillar 6018 class,

Excav2), one medium size hydraulic shovel (10.0 m3 Caterpillar 6018 class), and two front end

loaders (6.4 m3 Caterpillar 988H class wheel loaders, FEL1). This equipment will load a fleet of 50 t

capacity haul trucks (Scania G460CB 10x4 class units). A fleet of four 40 t capacity articulated dump

trucks (ADTs) was planned, which can be placed in service for either partial mining of saprolite

material within the pit, or re-handling main stockpile ore to the primary crusher.

Mid-way through the mine life, the smaller hydraulic excavators will be phased out, and large loading

equipment units will be brought into operation. These large units will be one hydraulic excavator (22.0

m3 Caterpillar 6040 class, Excav3), one hydraulic shovel (22.0 m3 Caterpillar 6040 class), and one

large front end loader (12.2 m3 Caterpillar 993K class, FEL2). These units will load a fleet of 133 tcapacity haul trucks (Caterpillar 785D class units).

The medium size hydraulic excavators and shovel (10.0 m3 Caterpillar 6018 class units) will be

capable of loading both the 50 t and 133 t capacity haul trucks.

The hydraulic excavators and shovels were estimated to be able to free-dig approximately 12% of the

total material within the planned open pit (saprolite waste and ore). Dry density for saprolite was

estimated to be 1.84 t/ m3 and for fresh rock 2.76 t/ m

3. Saprolite moisture content was estimated to

be 20% on average (varying with season and depth), and swell in loading to be 20%. Fresh rock

moisture content was estimated to be 6% on average, and swell in loading to be 40%.

Table 16.8.5.1 shows selected loading statistics for the planned loading units in saprolite.




Guyana

Figure 16.5.1

Toroparu Pit Phase DesignSource: SRK, 2013




Guyana

Figure 16.5.2

South-East Pit Phase DesignSource: SRK, 2013




Guyana

Figure 16.5.4

South-East Final Pit Design -Measured and Indicated BlocksSource: SRK, 2013




Guyana

Figure 16.5.5

Mine Plan Progress MapsYear 5 EndSource: SRK, 2013




Guyana

Figure 16.5.6

Mine Plan Progress MapsYear 10 EndSource: SRK, 2013





17 Recovery Methods

17.1 Summary

The Toroparu processing facility will be developed in three phases over the life of mine (LoM). This is

to accommodate for variation in ore types over the production schedule. .

The first period is characterized as the Pre-production Phase. This phase is estimated to consist of

two years, during which the facility will process 3,250 t/d of saprolite through a carbon-in-pulp (CIP)

leach circuit using a refurbished ball mill already in the possession of Sandspring. The CIP circuit and

downstream process equipment will be designed to expand to the full production rate for processing a

combination of flotation tailings and saprolite during Phase 1.

Phase 1 consists of processing 15,000 t/d of Gold Ore with Average Copper (ACO) via copper

flotation of gravity tailings with cyanide leaching of the cleaner scavenger flotation tailings via a CIP

circuit alongside saprolite. It is estimated that this will occur for the first five years of the mine life.

Phase 2 consists of processing 15,000 t/d of Gold Ore with Low Copper (LCO) via CIP leaching and7,500 t/d of ACO via flotation with CIP leaching of the cleaner scavenger tailing. This phase continues

over the remaining LoM.

17.2 Overview

The estimated mineral reserves included in the mine plan as developed by SRK total approximately

51.8 Mt of ACO grading 1.17 g/t Au, and 18% Cu; 70.3 Mt of LCO grading .89 g/t Au, and 5.0 Mt of

saprolite grading .91 g/t Au. The final process plant will be designed to treat 8.2 Mt/y of ore at a

throughput rate of 22,500 t/d dry solids at peak production during Phase 2. Instantaneous milling

throughput is dependent on ore hardness as well as which phase of production the facility is operating

under. The operating basis is 365 days available per year with continuous two shifts per day and

twelve-hour shifts. The annual operating hours for design for continuous process and water treatment

areas is 8,322 hours.

A simplified schematic drawing of the proposed facility is provided in Figure 1.7.1. Note that this figure

effectively combines both Phase 1 and Phase 2 into one diagram, thus not all equipment shown will

run simultaneously, but all equipment shown will operate at least in one phase.

17.3 Design Basis

17.3.1 Preproduction-Phase

During preproduction only saprolite ore will be processed. The design basis for the principal process

areas during the pre-production phase is summarized as follows:

A front-end loader will deposit RoM saprolite into a hopper. This hopper will discharge to a screen that

separates and sends oversized saprolite to a stockpile for future processing in later years. Screened

undersize reports to a refurbished ball mill already in the possession of Sandspring. The ball mill

grinds the ore to P80 200 μm, and operates in closed circuit with a cyclone. The cyclone overflow

reports to a CIP circuit. The CIP circuit is designed for a throughput of 3,250 t/d of saprolite ore. This

equipment will be utilized in the Phase 1 expansion.





Desorption and gold room areas are designed to recover the maximum gold, calculated nominally by

month over LoM. Carbon regeneration area is designed to reactivate five tons of carbon per day.

Detox area will be designed to treat 3,250 t/d of CIP circuit saprolite tailings at a feed density of 30%

solids and will achieve a final CNWAD concentration of 0.5 ppm. This target remains constant over

the LoM for all phases regardless of feed material

17.3.2 Phase 1: 15,000 t/d ACO and Saprolite

Phase 1 is designed to receive and process ACO in a two stage crushing circuit feeding into an HPGR

circuit. The saprolite ore runs in parallel to the ACO process maintaining the preproduction flow circuit

as described previously.

The crushing circuit will receive 15,000 t/d of ACO, which undergoes two stage crushing. Stage 1

consists of a jaw crusher. Stage 2 is comprised of a screen and cone crusher operating in open

circuit. The product of Stage 2 reports to the ACO flotation feed stockpile. Material from this stockpile

is reclaimed and fed to an HPGR operating in closed circuit with edge recycle being recirculated via

conveyors. The HPGR product undersize reports to a ball mill feed conveyor.

The ball mill circuit will grind ACO material at a throughput of 15,000 t/d and achieve a product size of

P80 150 μm. The ball mill circuit is equipped with a gravity concentrator to enhance gold recovery, the

concentrate of which undergoes intense cyanide leaching and recovery through electrowinning.

The ball mill cyclone overflow reports to two parallel trains of rougher flotation, the concentrate of

which reports to the regrind and cleaner flotation circuit. The rougher flotation tailings report to a

dedicated tailings thickener.

The regrind circuit consists of a single tower mill operating in closed circuit with a cyclone cluster, the

overflow of which reports to cleaner flotation. The cleaner flotation circuit consists of two parallel trains

of float cells. The concentrate from each bank of 2nd cleaner floatation cells is discharged to a

common concentrate thickener. Concentrate is dewatered via thickening and filtration to produce afinal product. The tailings from the 1st cleaner scavenger float cells are sent to the CIP circuit for gold

recovery via cyanide leach.

The CIP circuit consists is expanded to a single train of five tanks that receives both the saprolite and

ACO cleaner scavenger flotation tailings. Carbon is recovered from the CIP tanks and is sent to the

desorption circuit where it undergoes and acid wash and elution with a barren solution. The resulting

pregnant solution is processed through an electrowinning circuit to produce a gold sludge. The sludge

is then further refined to produce gold doré via a furnace. The barren CIP tailings report to tailings

thickener for dewatering prior to detoxification via the air/SO2 method.

17.3.3 Phase 2: 15,000 t/d LCO with 7,500 t/d ACOPhase 2 is designed to receive and process LCO and ACO and minor amounts of saprolite until the

end of the mine life. The crushing circuit is carried forward from Phase 1 and operated for a longer

period to accommodate the extra tonnage throughput. Crushing area will receive 15,000 t/d of LCO

and 7,500 t/d of ACO material, which are processed independent of each other in separate crushing

circuits. The product from the Stage 2 crusher reports to either the ACO flotation feed stockpile or the

LCO leach stockpile. Material from each stockpile is reclaimed and fed to an HPGR operating in

closed circuit with a screen. The LCO and ACO material have separate dedicated HPGR units. The





LCO material is sent to the previous Phase 1 HPGR unit; A new HPGR unit is needed to process the

ACO material. The configuration of each circuit follows as previously described for Phase 1 with each

of the screen undersize materials reporting to their respective ball mill feed conveyor.

The grinding area for each feed material is comprised of a ball mill operating in closed circuit with a

cyclone cluster. The Phase 1 ball mill is dedicated to processing LCO material and a new ball mill is

required to process the ACO material. The grinding mills will grind both materials to P80 150 μm.

Each grinding train also includes a gravity concentrator which receives a feed from the cyclone

underflow. The gravity concentrate reports to intense cyanide leaching as a batch process for gold

recovery. The leach solution from both circuits is subsequently pumped to the gold room

electrowinning area. Each ball mill will be sized for 15 k and 7.5 k.

The flotation circuit is the same circuit as in Phase 1, however only one train of float cells is used as

the throughput has been reduced to half. The regrind mill remains the same but will operate at a

reduced load.

The CIP area is expanded into two trains of five CIP tanks each, now receiving feed as fresh LCO

material and a reduced amount of ACO cleaner scavenger flotation tailings. Each train shares a single

loaded carbon recovery and a single barren carbon dewatering screen.

Tailings from the CIP process are sent to a designated tailings thickener. From here, they are

thickened to 45% solids prior to detox. Additionally, the thickener serves as a wash system to reduce

the CNWAD concentration prior to detoxification. Rougher flotation tailings report to a separate tailings

thickener, the underflow of which can bypass cyanide detoxification.

Desorption and gold room circuits are the same as in Phase 1

Detox area will treat 16,250 t/d of thickened adsorption circuit tailings using the air/SO2 method.

17.4 Mass and Water Balance

A detailed mass and water balance covering all streams within the process plant has been prepared

by Tt based on daily throughput requirements, plant availability, and the design feed grades for gold

and copper.

17.5 Process Design Criteria

17.5.1 Primary/Secondary Crushing and Stockpile

ROM material is deposited into a feed bin, which is discharged via an apron feeder to a grizzly at a

design throughput of 1,340 t/h.

Table 17.5.1.1: Primary Jaw Crusher

Crusher duty, h/d 16.8Crusher daily throughput, t/d (average) 22,500Crusher hourly throughput, t/h (design) 1,340Crusher availability, % 70Crusher type Jaw





Table 17.5.2.3: HPGR Phase 2

Number of HPGR’s 1Design Capacity, t/h 510 Roll width, m 1.2 Roll diameter, m 1.7 M-dot, ts/hm 200

Motor power, kW 895

Table 17.5.2.4 Ball Mill, Phase 2

Number of mills 1Mill size, m 5.5 x 7.9 EGL Mill power, kW 2,983 Ball mill product size, P80, µm 150 Classification type Hydrocyclone Recirculating load, % (nominal) 200 Cyclone diameter, mm 660 Number of cyclones 8 operating/4 standby

17.5.3 Gravity Circui t and Intense Cyanide Leaching

A gravity concentrator is placed on a split stream of each grinding train’s cyclone underflow stream.

The gravity tails reports back to the ball mill, while the concentrate is intermittently collected and

undergoes batch intense cyanide leaching. The leach solution is pumped directly to the pregnant

eluate tank in the electrowinning area.

17.5.4 Rougher Flotation Circui t

The rougher flotation circuit consists of two parallel trains of seven 100 m3 rougher flotation cells

operating at a feed density of 30% solids to achieve a plant retention time of 50 minutes. During

Phase 2, this is reduced to a single train of seven float cells. The concentrate from both trains ofrougher flotation are discharged to a common regrind cyclone feed pumpbox.

Table 17.5.4.1: Rougher Flotation Circu it

No. of parallel trains 2No. of float cells per train 7 Volume per float cell, m 100Plant retention time per train, min 50

17.5.5 Regrind Circui t

The regrind circuit consists of a single VTM1500 tower mill operating in closed circuit with a regrind

cyclone. The regrind cyclone feed box receives input from the tower mill discharge as well as rougher

flotation concentrate and products from the cleaner flotation circuit. The equipment associated with

the regrind circuit remains constant between Phase 1 and Phase 2. The grinding circuit has an

assumed recirculating load of 200% of the rougher concentrate.

17.5.6 Cleaner Flotation Circui t

The cleaner flotation circuit consists of two parallel trains of 1st cleaner, 2nd cleaner, and 1st cleaner

scavenger float cells. Each cleaner flotation train is sized to handle 71 t/h of feed solids with a feed





density of 25% solids. The 1st cleaner concentrate reports to 2nd cleaner flotation, while the tailings

report to 1st cleaner scavenger flotation. The scavenger concentrate and 2nd cleaner tailings report to

a pumpbox, which in turn discharges to the regrind cyclone feed pumpbox for recirculation. The 2nd

cleaner concentrate advances to the concentrate dewatering circuit. During Phase 2, only a single

train of cleaner flotation cells is operational due to the decreased availability of ACO material.

A summary of the cleaner flotation circuit criteria is presented below.

Table 17.5.6.1: 1st Cleaner Flotation Circuit

No. of parallel trains 2No. of float cells per train 5Volume per float cell, m 20Plant retention time per train, min 20

Table 17.5.6.2: 1st Cleaner Scavenger Flotation Circuit

No. of parallel trains 2No. of float cells per train 4

Volume per float cell, m 20Plant retention time per train, min 9

Table 17.5.6.3: 2nd Cleaner Flotation Circuit

No. of parallel trains 2No. of float cells per train 3Volume per float cell, m 10Plant retention time per train, min 7

17.5.7 Concentrate Dewatering Circui t

The concentrate dewatering circuit consists of a 11.0 m diameter concentrate thickener with

downstream pressure filters. The concentrate from 2nd cleaner flotation reports to the concentrate

thickener, where it is thickened to an underflow density of 60% solids. The moisture content is

subsequently reduced further to 92% solids in the pressure filters, which discharge into a concentrate

stockpile. Reclaimed water is recycled into the process.

The concentrate thickener and filter presses are sized to handle the anticipated Phase 1 demand of

approximately 8 t/h. During Phase 2, the same thickener is used with the option to operate with a

single filter press.

17.5.8 CIP Circuit

The CIP leach circuit for Preproduction and Phase 1 consists of one train of five tanks, two of which

serve as CIP stages. The solids feed rate is approximately 147 t/h and 176 t/h at continuous operation

for Preproduction and Phase 1, respectively. During Phase 2, the circuit is expanded to two parallel

trains, five tanks per train, of CIP receiving a combined feed of 713 t/h from both LCO feed and ACO

cleaner scavenger tails.

A summary of the CIP circuit criteria is presented below.





17.5.10 Electrowinning and Gold Room

Gold is recovered from pregnant eluate by electrowinning in a single train of three electrowinning

cells. Gold sludge is collected onto woven wire stainless steel mesh cathodes. The gold sludge is

recovered from the cathodes and further refined in a furnace to produce a gold doré. Despite the

absence of Hg in the deposit, a wet scrubber is included in the design to mitigate any potential

discharge into the atmosphere.

Table 17.5.10.1: Electrowinning

Number of electrowinning cells 3Cell Type SludgingNumber of Cells Operating 3Cell Arrangement Series

Table 17.5.10.2: Smel ting

Method Gas FiredHours per Smelt 4

Pours per Week 2Strips per Smelt 3-4Slimes Treatment Method Dry & Direct SmeltMaximum Bar Size 800 oz

17.5.11 Carbon Regeneration

Stripped carbon is treated in a rotary kiln operating at nominally 700° C to remove adsorbed organics

and restore carbon activity. The regeneration capability is designed to treat five tons of stripped

carbon in a 24 hour period.

A summary of the carbon regeneration criteria is presented below.

Table 17.5.11.1: Kiln Carbon Dewatering Screen

Screen type VibratingScreen deck material PolyurethaneScreen aperture, mm 0.8Screen feed rate, m /h 3.1Screen drainage rate, m /h/m 25

Table 17.5.11.2: Barren Carbon Dewatering Screen

Screen type VibratingScreen deck material PolyurethaneScreen aperture, mm 0.8

Screen drainage rate, m /h/m 25

17.5.12 CIP Tails Detoxification and Tails Dewatering

The CIP tailings report to a tailings thickener where they are thickened to approximately 38% solids in

Phase 1 and 45% in Phase 2. In addition to thickening, the underflow is washed such that the cyanide

content has been reduced prior to detoxification. The washed thickener underflow is subsequently

treated in an air/SO2 detox reactor to reduce CNWAD levels to below 0.5 ppm. The rougher flotation





tailings are treated in a separate dedicated tailings thickener such that the underflow can bypass the

detoxification circuit.

17.5.13 Reagents

The major reagents to be used in the process plant will be:

Lime (quicklime);

Sodium Cyanide;

Sodium Hydroxide;

Nitric Acid;

Flocculant;

Copper Sulfate;

Sodium Metabisulfite;

A3148A Flotation Collector; and

MIBC Flotation Frother.

The reagent facilities are described as follows.

Quicklime (Lime)

A wet slaking system in which lime pebbles will be fed to a tower mill in closed circuit with a cyclone.

The cyclone overflow product reports to a surge tank, from which lime slurry is pumped throughout the

process as needed. The demand for lime varies between LoM phases.

Sodium Cyanide

Sodium cyanide will be delivered in Isotainers in solid form. Dissolution and unloading will be carried

out simultaneously by pumping diluted solution from the mixing tank through the Isotainer and back to

the mixing tank. Cyanide will be added to the CIP and Elution Circuits via metering pumps. The

demand for cyanide varies over the LoM with each phase of production.

Sodium Hydroxide

Sodium hydroxide (caustic soda) will be delivered to the site in solid form in bulk bags and mixed in

the caustic soda mixing tank. The caustic soda solution will be pumped to the desorption and

electrowinning circuits.

Copper Sulfate

Copper sulfate will be delivered as needed for makeup and mixed in the copper sulfate mixing tank.

The mixed solution reports into the storage tank, from which it is pumped directly to the cyanide

detoxification circuit. The demand for copper sulfate varies each phase over the LoM, particularly due

to changes in copper in the RoM material.

Sodium Metabisulfite

Sodium metabisulfite (SMBS) will be delivered in bulk bags and mixed in the SMBS mixing tank and

subsequently transferred to the SMBS storage tank. This solution is pumped directly to the cyanide

detoxification circuit. The demand for SMBS varies dependent on the ore type processed.





A31418A Flotation Col lector

A31418A flotation collector will be delivered as a solution in a bulk container. This container is

offloaded into a holding tank, from which it is pumped directly to the flotation circuit.

MIBC

MIBC will be delivered as a solution in a bulk container. This container is offloaded into a holding tank,

from which it is pumped directly to the flotation circuit.

Flocculant

Flocculant will be delivered in bulk bags and mixed in the flocculant mixing tank, from which it is

pumped to the mixed flocculant holding tank for distribution to the thickeners throughout the process

facility.

Other Reagents

Other reagents and consumables used are:

Activated carbon for the CIP circuit; Borax, nitre and silica for gold room fluxes;

Diesel fuel for the elution heaters and regeneration kilns; and

Grinding media for ball mills and the tower mill.





Truck Fuel Facility and Equipment Ready Line

The vehicle fueling facility and ready line will be located at the entrance to the plant area adjacent to

the main haul road access to the mine pits. The fueling facility will store about 800,000 L of diesel for

the initial years and will be doubled (1,600,000 L) when the operation begins using the CAT 785D

trucks. Smaller tanks will hold a variety of oils, service fluids and lubricants. The ready line will be

located adjacent to the fueling facility and will be well lighted for 24- hour use.

Explosives Storage

An explosives storage pad is provided for supplier’s explosives magazine. An area of approximately

100 m x 100 m has been identified on high ground approximately 400 m north of the entrance station.

The explosives storage area is at an approximate elevation of 128 m. It is accessed by the main

access road to an area near the Puruni Bridge. The explosives storage area is isolated from the mine

support, plant and man camp facilities, but is in close proximity to the main and satellite pit operations.

18.1.5 Process Support Facilit ies

The process plant will be located on high ground, at an approximate elevation of 106 m, adjacent tothe run-of-mine stockpile and low grade ore stockpile. Adjacent to the plant will be the process

operations support facilities that include:

Administration Building;

Laboratory;

Workshop, warehouse and storage yard; and

Entry station.

Administrat ion Bui ld ing and First Aid Faci li ty

The administration building will be located in the plant area and south of the process facilities. The

building is sized as a 25 m x 18 m, single-story, pre-engineered, steel-framed structure with walls and

roof to be erected upon a spread footing foundation. The building will provide offices for the process

operations staff, conference/training facilities, toilets, break room, and safety.

Adjacent is a dedicated first aid facility (10 m x 5 m) for early care treatment.

Laboratory

The laboratory building will be approximately 50 m x 30 m, single-story, pre-engineered, steel-framed

structure with walls and roof to be erected upon a spread footing foundation. It will be located adjacent

to the main process facilities on the south side. The laboratory will house sample preparation,

assaying, testing facilities along with supporting sample and chemical storage rooms.

Warehouse, Workshop and Storage Yard A separate building is planned for process operations workshop/warehousing activities. The

warehouse/workshop will have about 1,800 m2 of covered floor space. This work shop is set-up to

perform major repair service to the process equipment. For equipment were the shop cannot

accommodate, repairs will be outsourced. Adjacent to the building will be a fenced compound that will

store bulk items and equipment as necessary.





Gatehouse and Weigh-Scale

An entry area including a guard house and weigh-scale will be located at the main entrance to the

property just south of the mine for traffic and inventory control. The entry area is graded above the

Puruni River floodplain at elevation 98 m. The entry area is approximately 100 m x 50 m to allow room

for a lay down, parking, and turn-around.

18.1.6 Man Camp

The man camp will be located approximately 700 m east of the plant area. The camp is defined by a

rectangular pad approximately 400 m x 200 m at elevation 105 m. The pad slopes down to the east at

2%. The camp facility is up-wind and up-gradient from the plant facility, and is shielded from the plant

area by a broad ridge with a maximum elevation of approximately 123 m.

A dormitory style man camp facility is proposed for the Project. The facility includes 490 single person

rooms with shared bath facilities, two 24 unit single dorms with private bath facilities, a recreation

complex, a commercial laundry facility, a medical unit, an office unit that includes a greeting area and

four offices, and a kitchen and dining unit.

18.1.7 Additional Suppor t Facilities

Airstr ip

The Toroparu site includes an existing airstrip which will remain in service for the life of the Project.

The existing airstrip location will interfere with the main pit ultimate boundary, later in the mine life. To

alleviate this interference, it will be extended approximately 200 m toward the southeast.

Communications

Radio Communications Opportunities and Risks

A radio base station will be provided for plant wide site-to-office communications within a 10km radius

from the communications center for use in the following areas:

Process plant;

Plant office;

Laboratory;

Mine workshop and offices;

Central control building; and

Each substation.

This will reduce wiring costs and allow voice messaging integration with e-mail. End-to-end IP video

connectivity with business quality transmission will provide video conferencing capabilities.

Satellite Communications

A satellite communications network will be provided for site-to-site communications between Toroparu

operations, Georgetown and outside of Guyana. The system includes voice/data/video/fax, internet,

and VPN services, including bidirectional links between the mine site and Georgetown.

Satellite phones will be installed at strategic areas for emergency communications.





Fiber Optic

The IT system will be based at the communications building and connected throughout the site by a

fiber optic network. The connection between IT devices and end-users will provide high through put,

secure, reliable and redundant service for data and voice. The network system will be connected to

protocol independent multicasts (PIMS) and business networks through routers with firewalls and will

provide remote access as required. The system will have security and encryption to prevent

unauthorized access.

18.1.8 Power Supply and Distr ibut ion

Due to the remote location of the Project, a site-based power plant is required. The power supply will

be provided via a series of interconnected 7.4 MW gen-set units generating power at 13.8 kV, 60 Hz.

Power distribution for the Project site will be by wooden pole overhead power lines from the substation

at the generator plant routed to the process plant, mine support area and camp. Transformers and

switchgear will be located at each of the buildings/facilities in the process plant with individual

transformers and switchgear and Motor Control Centers for the local power loads.

The power plant will be fueled by intermediate fuel oil specifically, IFO 180. The power plant will be

constructed in modules with the ability to add generators into the system as power demand grows.

Phase 1 of the Project will require a peak power demand of 22.4 MW to support the operation. This

peak demand increases to 26.7 MW during Phase 2 of the Project through the life of mine.

The power plant will have its own fuel storage facility with a one month on-site fuel storage capacity

for IFO 180 for continuous operation of the power station. One 1,000 m3 fuel oil storage tanks will be

built next to the power plant in Phase 1 and a second identical tank will be incorporated to support

Phase 2.

Typical voltages will be 4160 volts for motors greater than 200 kW and at 13.2 kV for the larger ball

mills. Additional transformers and electrical equipment will be provided for those electrical loads lessthan 200 kW at 480 volts.

Power for the entry station will be provided by a standalone diesel generator to be utilized on an as

needed basis. The Project will reuse the existing 1.5 MW unit for this supply.

18.1.9 Water Supply

Process Water

Initial fill for the process water requirements will come from the Wynamu River. Make-up water for the

process plant will be obtained by recycling decant water from the Tailings Management Area (TMA)

via a water reclaim system. A 190 m3 (50,000 gallon) water storage tank will be located at the plant

site.

Fire Water

Fire water for the plant facilities will come from the Wynamu diversion pond. The fire water shall be

distributed around the site in an arrangement to cover all the ancillary buildings, process plant, mine

support facilities, workshops and yard equipment. The systems are sized to provide four hours of fire

water demand and are fully integrated with a fire alarm and detection systems. Costs for a fire water

and protection system have been allocated in the estimate.





Potable Water

Drinking water for the man camp and plant site will come from rain water that will be collected from

camp building roofs. The building gutter system will convey the collected water to a screening and

filtration system where contaminants are collected and separated from the roof rain water collected.

Collected rain water from the site will be conveyed to a centralized potable water storage tank. A 284

m3 (75,000 gallon) water storage tank will be located at the camp site.

The water supply can be supplemented with river water from the Wynamu diversion pond during dry

periods. However, water from the river will need to be treated using a package plant to bring the water

up to drinking water standards. Ultrafiltration has been identified as the most effective means of

treating the Wynamu river water.

18.1.10 Waste Water Treatment and Solid Waste Disposal

Waste Water Treatment

Wastewater from the camp and process facilities will be treated using a package plant. Two dedicated

sewage treatment systems will be provided, one system for the man camp site and one system for theplant facilities area. The collected sewage will be transferred to packaged wastewater treatment

systems. Each system will be capable of treating 50,000 L of wastewater per day. Treated effluent

from the man camp facility will piped east of the camp facility and released into a local unnamed

tributary which will eventually flows into the Puruni River while the treated effluent from the plant area

location will be piped to the north and released into the Wynamu River. Sludge produced from the

wastewater treatment systems will be disposed of in the landfill with other waste material.

Solid Waste Disposal

It is assumed that non-recyclable, non-toxic solid waste will be disposed of at an onsite, lined landfill

to be sited and designed at a later stage.

18.2 Tailings Management Area

The TMA will be located on the northeast side of the Toroparu property. This facility will be staged and

operated in three independent cells as shown in Figure 18.2.1. The cells will operate separately at

different stages of the service life of the TMA as follows:

Cell 1 will operate up to the first quarter of Year 5 of full production;

Cell 2 will operate from the second quarter of Year 7 to second quarter of Year 9; and

Cell 3 will operate from the third quarter of Year 11 to end of the mine life at Year 16.

Detoxed tailings will be discharged into the TMA cells from the crests of the tailings dams, and

supernatant water volumes reclaimed to plant by a floating pump barge positioned on the decantponds of the cells. Excess water volumes will be discharged through spillways into three collections

ponds located adjacent to the cells for monitoring and control previous to release to the environment.

The design is based on the near surface ground conditions being sufficiently impermeable to prevent

ground contamination without a geosynthetic liner, and that the ground has sufficient strength to

support that perimeter dam slopes. Fill and compaction of in-situ saprolite was included in cost

estimates. The TMA has a storage capacity of up to 143 Mt. The tailings dry density estimate used for

the design was 1.30 t/m3.





The geotechnical conditions of the tailings site have been assumed from the geology information

previously described and from the results of the investigation program carried out by KCB (2012) at

the location of the tailings site defined in the PEA. A specific site investigation program for the tailings

site should therefore be carried out for the new location defined in this study during feasibility

engineering of the project.

18.3 Off-Site Infrastructure and Logist ic Requirements

18.3.1 Off-Site Infrastructure

Port Facilities

The port facility at Pine Tree Landing will be a key feature in the mine supply chain. Pine Tree Landing

is located on the right bank of the Cuyuni River, approximately 2 miles above the confluence with the

Mazaruni River, and has historically been used by logging companies as a shipping point for lumber

products and supplies. Pine Tree Landing is linked to the Toroparu Mine by the existing Upper

Mazaruni Development Authority & Toroparu Mine Road.

Currently, Pine Tree Landing has approximately 225 m of waterfront, and is cleared of vegetation for

approximately 500 m from the Cuyuni River, for a total area of approximately 32 acres. The existing

facilities are very limited and will not be reutilized.

The proposed port facility will generally include the following items:

Wharf;

Man camp;

Aggregate surfacing;

Container storage areas;

Bulk fuel handling infrastructure and storage;

Equipment laydown areas; Maintenance building;

Administration building;

Access control gate;

Lighting;

General civil infrastructure;

Security; and

Communications systems.

Mine fuel supply logistics and infrastructure will be provided by a contracted fuel supplier. Delivery will

be via supplier barge, and the supplier will construct the required infrastructure within the port facility

for offloading and storage of HFO and diesel. Initially, two 5,000 bbl tanks will be constructed,sufficient for an approximately 1 month supply of fuel during the mine construction phase. Expansions

of the depot will be made by the supplier as the mine progresses into start-up and expansion phases.

The tanks and infrastructure will include fire suppression and spill containment features.

The port facility will accommodate 1,000 ton barges which will transfer cargo between Georgetown

and Pine Tree Landing via the Essequibo, Mazaruni and Cuyuni Rivers. Containerized cargo is

anticipated to be approximately 45 containers import and 45 containers export per week. Cargo at the





port will be handled primarily with 40-ton forklifts and reach stackers. Heavy or oversized cargo will be

handled by a mobile harbor crane or as roll-off cargo.

Access Road

The existing roadway begins at Pine Tree Landing on the Cuyuni River and intersects the Upper

Mazaruni Development Authority (UMDA) road about 10 km west of Itaballi. At an intersection knownas Camp 4, located ~200 km (120 miles) west of Itaballi, the roadway continues north to the mine via

the Toroparu Mine Road, an approximately 30 km (20 miles) section constructed in the mid- 2000’s by

Sandspring.

The UMDA, established to administer infrastructure development funded by the United Nations

Development Program, constructed the existing road from the Mazaruni River village of Itaballi to

Kurupung in the mid-1970’s. The road is traveled today by a number of users, primarily mining and

logging related, and serves several small communities. The road is now identified as the Itaballi-

Puruni-Papishao Road by the Government of Guyana (GoG).

Upon cancellation of the UMDA hydroelectric projects in the late 1970’s, sections of the UMDA road

from Puruni Landing to Kurupung fell into disrepair. The establishment of the Toroparu explorationcamp by Sandspring’s Guyanese subsidiary, ETK, in 2001 was followed by the re-construction of

these road sections and the construction of a new road connecting the UMDA road to Toroparu.

The upgrade of the approximately 230 km (140 miles) roadway will generally include:

Brush back of vegetation;

Subgrade stabilization;

Installation or repair of culverts and bridges;

General grading and subgrade preparation earthwork;

Slope stabilization;

Fine grading; and

Production and placement of aggregate surfacing.

FMG conducted a reconnaissance of the roadway alignment in December of 2011. Escorted by

Sandspring personnel, an FMG engineer traveled the roadway from Itaballi to the Toroparu Mine over

the course of several days. The existing roadway conditions and features, including bridges, culverts,

soft subgrades, potential aggregate sources, and other pertinent information, were noted and

recorded. The Sandspring personnel provided insight into potentially applicable construction

techniques and specific upgrade concerns.




Guyana

Figure 18.2.1

Tailings Management AreaSource: KCB, 2013





19.4 Contracts and Status

Currently there are no material contracts in place other than those disclosed in this document. It is

anticipated that the following contracts will be in place upon Project commencement.

19.4.1 Metal Treatment, Refining, and Transpor tation

Agreement for the secure transport of doré by air from Toroparu to refinery in North America

or Europe;

Agreement for the refining of gold doré and delivery to Sandspring’s designated bullion

account;

Agreement for the treatment & refining of copper concentrates;

Agreement for the transportation of containerized copper concentrates from Pine Tree

Landing to Georgetown, and transshipment for delivery to offshore custom smelter; and

Agreement for transportation insurance for export of precious and base metal cargoes.

19.4.2 Supplier & Service Contracts

Barge transportation of supplies between Georgetown Harbour and Pine Tree Landing;

Diesel and fuel oil supply and delivery to Pine Tree Landing;

Process reagents, consumables, and supply contracts;

Equipment preventive maintenance services;

Air transportation (Georgetown to site) services; and

Site security services.

19.5 Indicative Terms

Terms used in the development of financial estimates of revenue and costs are as follows.

19.5.1 Doré Net Smelter Return

Based on actual costs from other gold producers with similar sized operations:

Gold Payable: 99.9%

Refining Charge $0.65/oz Au

Secured air transport and Insurance $2.45/oz Au

19.5.2 Copper Concentrate Net Smelter Return

Based on indicative proposal received from multinational copper company:

Copper Payable 96.5% of contained Cu s.t. min deduction of 1%

Gold Payable: 97% of contained gold

Silver Payable: 97% of contained gold

Treatment Charge $95.00/t

(2012 benchmark + US$30/t)

Copper Refining Charge $0.065/payable lb Cu

Gold Refining Charge $6.50/ payable ounce of gold

Silver Refining Charge $0.65/payable lb Cu





20 Environmental Studies, Permitt ing and Social orCommunity ImpactSRK has prepared this section based on the work conducted by ETK and by Sandspring’s

consultants. ETK performed the baseline studies. KCB performed the geochemistry studies anddesigned the TMA and SRK designed the waste rock management facilities.

The Toroparu Gold Project is subject to a number of regulatory permits and licenses, issued by

several different governmental agencies. The Project has received its environmental permit for gold

and copper mining and processing based on a permit application dated May 2, 2008, and the

approved ESIA (ETK, 2012). The permit is valid from June 2012 through May 2017 and was issued to

ETK. Sandspring acquired ETK as part of its acquisition of GoldHeart Investment Holding Ltd. in May

2009, and thus the environmental permit is valid for Sandspring. The permit includes design,

operational and reporting compliance items.

Sandspring has signed a binding Memorandum of Understanding with the Government of Guyana that

grants exclusive right to evaluate and develop and hydroelectric power plant on the Kurupung River,

approximately 30 miles from the Project. A separate environmental permit will be required for the

hydroelectric power plant.

The applicable primary permit or license requirements are summarized in Table 20.1.





variations in rainfall. The entire area identified for development into the mine, like the rest of Guyana

has a tropical climate and is not subject to extreme variations in temperature and humidity.

Sandspring maintained a weather station at the Project during 2005 and during discrete periods of

2005. Very little climate data are available locally. During the 2005 monitoring the maximum annual

precipitation was 2100 mm, and temperatures ranged from a low of 68 to a high of 108 degrees

Fahrenheit. The dry periods were January through March and August through October.

Winds were primarily from the north-northeast and the average maximum speed was 9.1 m/s. It is

generally windier during the short wet season.

It is recommended that monitoring at the weather station be re-established, and that the data collected

include evaporation information. It would also be helpful to have precipitation data recorded at

intervals less than 1 hour to understand the severity of storms.

Air Quali ty

In the area surrounding the Project site, there are no major industries that serve as significant sources

of fixed or mobile atmospheric emissions. Aerial emissions are mostly attributable to the gases from

rotting trees and other vegetative matter although some background emissions will inevitably berelated to the operation of various small to medium sized motorized equipment in the area. Due to the

high humidity and significant rainfall, dust levels on the roadways are generally low. Airborne

discharges and particulate matter are not monitored in the area but are not expected to exceed the

emission guidance established by the World Bank or WHO Ambient Air Quality guidelines.

Surface Water

The Toroparu Gold Project area is drained by the Puruni River and by several tributaries, the main

one being the Wynamu River. The total estimated drainage area of the Puruni River is approximately

4170 km2. Approximately 375 km2 of that drainage area is located upstream of the proposed mine

site.

In 2010, water quality samples were obtained at three upstream locations to assess background water

quality for the proposed mine. The surface water sampling locations are shown in Figure 20.1.1.1. The

water quality samples are demonstrative of water quality impacts from the former open pit mining

operation and are also indicative of water quality prior to the commencement of additional mining

operations. For comparison purposes the results were compared to the International Finance

Corporation (IFC) effluent requirements for mining operations (2007), because the Guyana EPA

guideline does not present permissible limits for water quality standards. The majority of the sample

results were below the IFC effluent requirements; however oil and grease and iron were exceeded in

one or more samples.

Groundwater

The site occurs in the Precambrian crystalline basement rock section of Guyana. Five groundwater

monitoring wells were installed to varying depths and at different locations during the first round of

investigations conducted at the concession. The locations of these wells are depicted as MW-1, MW-

2, MW-3, MW-4 and MW-5 on Figure 20.1.1.2. The location of each well, the top of casing elevation

and the static groundwater level, measured during each phase of baseline collection work at the site

are detailed in Table 20.1.1.1.





Table 20.1.1.1: Monitor Well Locations and Groundwater Elevation Data

WellIdentifier

Location (UTM) Top ofCasingElevation(masl)

Groundwater Elevation Measurement (masl)

Easting Northing 16/07/07 28/10/07 3/1/008 4/11/08 7/11/10

MW-1 824353 715217 32.87 31.70 31.04 32.03 30.48 32.20MW-2 825149 713946 31.27 30.26 No data 31.14 28.33 31.14MW-3 825948 713313 32.04 30.06 29.68 30.78 29.37 31.88MW-4 826293 714540 32.50 30.01 29.73 30.63 29.14 31.14MW-5 826854 713914 32.87 30.06 29.12 30.77 28.98 30.61

Source: ETK, 2012.Masl= meters above mean sea levelElevations are rounded to the nearest hundredth of a meter.

Groundwater levels recorded for the dry seasons are generally lower than those recorded for the wet

seasons. This can be interpreted as being indicative of precipitation being the primary source of

groundwater recharge in the Project area. No water was present in MW-3 during the second phase of

the field work, coinciding with conditions recorded in the long dry season.

Very little data are available on groundwater flow parameters for that section of Guyana. Observations

of remnant mines in the area indicate some groundwater inflow through the weathered unconsolidated

material overlying intact rock. That flow may, however, be reflective of recharge by precipitation.

Rising head in situ hydraulic conductivity tests were conducted in each monitor well after each

baseline sampling event. The hydraulic conductivities ranged from 9.75 x 10-5 cm/sec in MW-3 to

7.43 x 10-8 cm/sec in MW-5.

Groundwater samples were collected and analyzed according to the parameters mandated by the

Guyana EPA guidelines. The baseline results during the dry seasons were similar to those for the wet

seasons. The Guyana EPA guideline does not present permissible limits for water quality standards,

so for comparison purposes the results were compared to the International Finance Corporationeffluent requirements for mining operations (2007). Groundwater samples had slight exceedances of

Iron and pH and very high exceedances of total suspended solids. Iron ranged from non-detectable at

a detection limit of 0.03 mg/L to a high of 8.35 mg/L compared to the guidance value of 2.0 mg/L,

while pH ranged from 5.62 to 8.91 in comparison to the guidance value of 6 – 9 pH units. Total

suspended solids had very high exceedances ranging from 100 to 26,774 mg/L compared to a

guidance value of 50 mg/L. Since naturally-occurring groundwater typically would not exhibit high

levels of total suspended solids, SRK recommends that the sampling methodology and water

construction and development procedures be further reviewed to see if the well filter pack is

appropriate, the well development was adequate and the sampling technique is acceptable to

international standards.

Archaeological Resources

The site is located in the Middle Mazaruni area. The mine site would have been encompassed by

former Akawaio settlements and is contained within the Mazaruni Amerindian Reserve demarcated in

1904. This demarcation was reduced to the Upper Mazaruni Amerindian District in 1959. There are no

records of any immigrant activity in the area.





Herpetofauna

The relative species abundance determined during the survey indicates that the habitat may be ideal

for amphibians and reptiles. None of the species documented are endemic species or listed by

CITES.

Fauna Species of Interest

No locally rare, threatened or endangered species were recorded during the surveys; however a

number of species identified are listed by CITES. These along with other species listed by CITES and

their status are provided in Table 20.1.1.2.

Table 20.1.1.2: International Status of Species

Species Common Name International StatusLocal Status(Unofficial)

Caiman crocodiles Spectacle Caiman CITES Appendix II/III Fairly Common

Paleosuchus sp Dwarf Caiman CITES Appendix II/III common

Eunectes murinus Anaconda CITES Appendix II/III common

Epipedobates trivittatus Poison Frog CITES Appendix II/III Uncommon

Epipedobates sp Poison Frog CITES Appendix II/III Uncommon

Panthera onca JaguarCITES Appendix I, IUCNLower Risk- Near Threatened Species

Uncommon

Tapirus terrestris Lowland (Brazilian) tapir IUCN Lower Risk- Near Threatened SpeciesUncommon

Cebus olivaceus Wedged-capped Monkey CITES Appendix II/III Common

Saimiri scuries Squirrel Monkey CITES Appendix II/III Common

Agouti paca Labba CITES Appendix II/III Common

Oryzoborus angolensis Lesser seed Finch CITES Appendix I Uncommon

Amazona amazonica Orange-winged parrot CITES Appendix II/III Uncommon

Amazona farinosa Mealy Parrot CITES Appendix II/III Fairly Common

Amazona ochrophala Yellow-crowned Parrot CITES Appendix II/III Fairly Common

Ara chloropterus Red & Green Macaw CITES Appendix II/III Common

Brotogeris chrysoptera Golden Winged ParakeetsCITES Appendix II/III Common

Deroptyus accipitrinus Red-fan Parrot CITES Appendix II/III Common

Piontes melanocephalusBlack-headed Parrot CITES Appendix II/III Common

Pionus fuscus Dusky Parrot CITES Appendix II/III Common

Pionus menstruus Blue head Parrot CITES Appendix II/III Common

Pyrrhura picta Painted Parakeets CITES Appendix II/III Common

Pteroglossus aracari Black-necked Aracari CITES Appendix II/III Common

Source: ETK, 2012.CITES = Convention on International Trade in Endangered Species of Wild Fauna and Flora (2013). Classification includesthree CITES appendices:

Appendix I includes species threatened with extinction. Appendix II includes species not necessarily threatened with extinction, but in which trade must be controlled in order

to avoid utilization incompatible with their survival. Appendix III contains species that are protected in at least one country.

20.1.2 Results of Geochemical Studies of Tailings, Waste Rock and Low Grade Ore

The exposure of tailings, waste rock, open pit walls, the Low Grade Ore (LGO) stockpile, and tailings

to air and water may result in the generation of Acid Rock Drainage and Metal Leaching (ARD/ML).

Geochemical characterization studies were conducted by KCB from 2011 to 2013 (KCB, 2011, 2012,

2012a, and 2013) on the dominant bedrock lithologies representing waste rock and LGO, and

metallurgical tailings representing the three main ore types [that is, saprolite, Average Copper Ore

(ACO) and Low Copper Ore (LCO)], for ARD/ML potential. The geochemical characterization studies





Mineralogy by Optical Petrography and X-ray Diffraction with Rietveld-refinement;

Net Acid Generation (NAG) tests; and

Shake Flask Extraction tests.

Optical petrography was completed by Mineral Services Inc. located in North Vancouver, Canada.

The X-ray Diffraction with Rietveld-refinement analysis was carried out by the Department of Earth

and Ocean Sciences, University of British Columbia, Vancouver, Canada. All other analysis and test

work was carried out by Maxxam Analytics located in Burnaby, British Columbia, Canada.

Metal Leaching Risk of Waste Rock and LGO

Results of the solid-phase elemental analysis indicated that the lithologies included high

concentrations of silver, arsenic, cobalt, chromium, copper, nickel, molybdenum, sulfur and selenium

in comparison to average crustal abundance of high-calcium granite. There was a wide variation

between the different lithologic units. The LGO samples indicated high solid-phase concentrations of

silver, copper, nickel, selenium and sulfur. Therefore the elevated concentrations of these elements in

the solid-phase may be at risk of leaching under site-specific field conditions.

The results of Shake Flask Extraction tests indicated elevated leachate concentrations of aluminumand selenium, relative to water quality guidelines, from non-saprolite waste rock lithologies and LGO

samples. The more aggressive NAG tests also indicated elevated chromium and copper in leachate

from one or more samples, relative to water quality guidelines. For the saprolite waste rock samples,

phosphorous, relative to water quality guidelines, was leached and readily soluble from Shake Flask

Extraction tests. For the more aggressive NAG test, chromium and silver were also leached and

readily soluble. With the exception of phosphorous, these leachate extraction test results are

consistent with elevated solid-phase concentrations of silver and chromium.

The NAG extraction test results reported concentrations of phosphorus, aluminum, chromium, copper,

selenium and silver were elevated above reference guidance using Canadian Council of Ministers of

the Environment water quality guidelines for the protection of aquatic life. The pH values for wasterock samples were above the guidance values, except for one saprolite sample that had a pH below

the guidance value of 6.5. The Shake Flask Extraction results indicated that phosphorus, aluminum

and selenium concentrations were elevated and that one sample had a low pH.

There are elevated concentrations of silver, arsenic, cobalt, copper, chromium, molybdenum, nickel,

selenium and sulfur compared to crustal rocks. The short-term leaching tests reported leaching of

aluminum, selenium, chromium and copper from non-saprolite waste rock. The saprolite waste rock

was observed to leach phosphorus, chromium and silver at concentrations above the aquatic life

guidelines.

Acid Rock Drainage Risk of Waste Rock and LGO

The paste pH results indicated that the major lithologies are alkaline with the exception of the saprolite

and the Transition Zone samples. The saprolite samples were slightly acidic to neutral while the

transition zone samples were neutral to alkaline. These results indicate that no acidity was released

from any of the samples except from the saprolite samples. The alkaline results indicate effective

carbonate buffering. The LGO samples are alkaline, which indicates a potential buffering capacity.





The waste rock sample results were low in total-sulfur and sulfide-sulfur content, and the associated

calculated sulfide-based Acid Potential (AP) values were also low. The LGO total-sulfur and sulfide-

sulfur contents and sulfide-based AP were also low.

The waste rock lithologies and LGO samples contained low to moderate Neutralization Potential, (NP)

with the exception of the saprolite and Transition Zone samples, which contained negligible NP. With

the exception of the saprolite samples, the Net Potential Ratio (NPR), the ratio of NP to AP, calculated

for the waste rock lithologies indicated that the waste rock was classified as not-Potentially Acid

Generating (N-PAG), and therefore have a very low potential to generate ARD. The Transition Zone

samples also had a low ARD potential. For the saprolite, most of the samples were Acid-Generating

(AG), although some were classified as Potentially Acid Generating (PAG) and N-PAG. The LGO

samples were N-PAG.

The NAG pH results confirmed the not-PAG ARD risk of waste rock and LGO samples. The saprolite

samples had the lowest pH results (5.9 and 6.8 pH units) whereas the other samples had NAG pH

results between 11.0 and 11.5 pH units. A NAG pH of 4.5 or less is indicative of PAG material. Based

on the results of the NAG tests, most of the samples had very low sulfide content and an abundance

of neutralizing minerals. The saprolite samples had mixed results regarding its potential to produceacid, but it was concluded to be most likely N-PAG due to its low sulfide-sulfur content (0.07% wt.).

Humidity cell testing was recommended to be completed to further assess metal leaching of waste

rock, LGO and open pit walls under alkaline conditions.

Tailings Geochemistry Testing and Results

Static testing was completed on six tailings samples and four supernatant samples by KCB (2013).

Three ore types (saprolite, ACO and LCO) were subjected to gravity separation, flotation and cyanide

leaching to create tailings of each ore type. The six tailings samples in the geochemical testing

included the following:

Two samples of ACO cleaner detoxified tailing pulp combined with rougher tailings;

One LCO rougher and cleaner composite tailings sample; and

Three saprolite samples (cleaner detoxified slurry, coarse cyanide leach slurry and coarse

flotation slurry).

The samples of these three metallurgical tailings streams (saprolite, ACO and LCO) were submitted

for geochemical characterization. The tailings samples were submitted for the following static and

leachate extraction tests:

Mineralogical analysis;

Solid-phase elemental analysis

ABA; and Supernatant aging tests.

All analysis and test work was carried out by Maxxam Analytics located in Burnaby, British Columbia,

Canada.

Metal Leaching Risk of Tailings

A screening level comparison of the elemental analysis of the tailings solids to three times average

crustal abundances indicated that silver, arsenic, bismuth, cadmium, cobalt, chromium, copper,





molybdenum, lead, antimony, tin, tungsten zinc were elevated and may be at risk of leaching under

site-specific field conditions.

A total of four tailings slurry samples, including two ACO, one LCO and one saprolite tailings slurry

sample following detoxification, were selected for an aging test of the supernatant to provide an

indication of how the quality of the TMA ponded water may vary over time. The aging tests were

conducted for a period of 90 days, with supernatant sampling analysis at 1, 7, 14, 21, 20, 60 and 90

days.

The tailings supernatant aging test results indicated that fluoride, nitrite, ammonium, CNWAD,

aluminum, arsenic, chromium, cobalt, copper, iron, molybdenum, and selenium concentrations were

above the Canadian and British Columbia water quality guidelines and therefore may be parameters

of environmental concern (KCB, 2013). Additionally, iron and CNWAD were elevated in LCO and/or

ACO tailings supernatant.

The TMA will receive a combination of precipitation, water treatment plant brine, and supernatant from

the tailings slurry. Although the metallurgical tailings leachate extraction test results indicated elevated

concentrations that may be soluble and mobile under laboratory test conditions, the results do not

imply that they will be elevated to levels above these guidelines under site-specific field conditions,

rather they identify elements that are prone to leaching. The TMA, will receive a combination of

precipitation, water treatment plant brine, and supernatant from the tailings slurry. The TMA water

quality will be influenced by contributions from all these sources. The TMA design assumes that the

natural low permeability of the surficial soils, and the lower concentrations of elements in the TMA

pond due to attenuation from natural degradation, settling, and mixing with precipitation, which

averages about 2.6 m annually, will reduce concentrations in any TMA discharge effluent to the

aquatic receiving environment. Additional analysis (i.e., predictive water quality modeling) will be

needed in a later phase to verify this assumption.

Additionally, the TMA management strategy of subaqueous tailings deposition combined with a

cyanide destruction goal of 0.5 mg/L may be sufficient to mitigate potential environmental impacts to

the aquatic receiving environment from effluent discharge from the TMA (KCB, 2013). Kinetic testing

was recommended to evaluate the behavior of the tailings under flooded conditions.

Acid Rock Drainage Risk of Tail ings

The ABA results indicated that all samples have a neutral to alkaline paste pH and an acid-buffering

capacity. However, the neutral to alkaline paste pH values are expected for metallurgical testing with

lime addition.

The total sulfur ranged from below the detection limit to 0.08% wt. Sulfur speciation analyses indicated

very low to negligible sulfide-sulfur concentrations with only one sample result reported above the

detection limit of 0.03% wt. The sulfide-based AP of all the tailings samples varied between 0.16 kgand 0.94 kg CaCO3/t, with a median value of 0.23 kg CaCO3/t, indicating a low sulfide reservoir to

oxidize and generate acidity. The inorganic carbon measured as CO2 varied between 0.010% and

2.49%. The corresponding Inorganic Carbon Neutralization Potential (Inorg-CaNP) ranged from 8.33

kg to 207.5 kg CaCO3/t, with a median value of 82.5 kg CaCO3/t.

The Sobek Neutralization Potential (NP) ranged from 10 kg to 134 kg CaCO3/t, with a median value of

56.9 kg CaCO3/t. The bulk of the Neutralization Potential appears to be from reactive carbonates

and/or the addition of lime during the metallurgical testing. The saprolite tailings samples contained





moderate NP (7 kg to 10 kg CaCO3/t) and the two copper ore tailings contained high NP (104 kg to

134 kg CaCO3/t).

The tailings samples were classified as not-Potentially Acid Generating (N-PAG) based on the sulfide-

sulfur values and are therefore considered to have negligible risk of ARD (KCB, 2013).

20.2 Environmental Issues Although additional studies are recommended to further develop mining waste management

strategies, there are no known environmental issues that could materially impact Sandspring’s ability

to extract the mineral resources or reserves at the site. Preliminary mitigation strategies have been

developed to reduce environmental impacts to meet regulatory requirements and the specifications of

the environmental permit.

20.3 Operating and Post Closure Requirements and Plans

The overall environmental management objective of the Toroparu Project is to use Best Available

Techniques (BATs), Best Management Practices (BMPs) and modern, proven technology to operate a

gold and copper mine, process plant, and supporting infrastructure consistent with the social,

economic and environmental requirements of the Government of Guyana and, to the extent that they

represent recognized international BMPs and World Bank/IFC/Equator Principles policies and

guidelines.

Sandspring will establish and maintain a documented, comprehensive Environmental and Social

Management System (ESMS) over the construction, operation and closure phases of the Project. The

ESMS will be based on current World Bank Group/International Finance Corporation guidelines.

The environmental permit requires a number of operating plans, including the Environmental

Management Plan components listed in the ESIA:

Open Pit Management

Overburden Management

Water Management

Tailings Pond Management

Hazardous Materials Management

Explosives Management

Cyanide Management

Waste Management

Spill Contingency Plan

Catchment Area Management

Social Management Plan Erosion and Sediment Control Plan

Land Reclamation and the Road Management Plan

The permit requires that a Health Safety and Environmental (HSE) officer be employed and be

responsible for the implementation of the Environmental Management Plan. In addition to the plans

listed above, the permit contained requirements for biodiversity protection and air quality

management.





Table 20.5.1: Summary of Socio-cul tural Impacts and Mitigation Strategies





20.6 Mine Closure

The overall intent of the closure plan is to achieve Project objectives for restoring the site and aquatic

environment to a high ecological value. The objectives of the closure plan are to:

1. Prevent, reduce or mitigate the adverse environmental effects associated with the Project;

2. Provide for the reclamation of all affected sites and landscapes to a stable and safe condition;3. Provide for the return of all affected ecosystems to healthy and sustainable functioning;

4. Reduce the need for long-term monitoring and maintenance by designing for closure and

instituting progressive reclamation;

5. Provide for long-term monitoring and maintenance of the sites affected by the Project as

required;

6. Provide for mine closure using the most current available proven technologies in a manner

consistent with sustainable development.

Closure will result in the establishment of conditions that support public safety through physical

stability (Physical Stability); it will encourage productive end land use by promotion of revegetation

and promote conditions for biological stability (Biological Stability); and will ensure that mechanismsare in place to protect water resources and the receiving environment, thereby providing chemical

stability (Chemical Stability). Performance standards related to physical, biological and chemical

stability would function as measures of accomplishment of the closure objectives. The Project

performance standards are as follows:

Physical Stability – Preservation of protective safety measures (in a state in which the measures

can be effective) throughout the post-closure monitoring period, once there has been no external

human influence.

Biological Stability – Effective revegetation and restoration evidenced by vegetative proliferation

on 70% of the site areas intended for revegetation by the end of the post-closure monitoring

period.

Chemical Stability – Water quality similar or improved when compared with historic data at the

end of the post-closure monitoring period, once there has been no human or related influence.

The conceptual closure activities incorporate strategies to protect surface water and groundwater;

prevent erosion and control discharge from reclaimed mine facilities; and protect wildlife, as outlined

below.

The storm-water management practices will provide systems that minimize environmental damage by:

Maximizing retention time within the system by use of detention ponds

Minimizing increases in surface runoff flow and volume

Channeling and diverting runoff Use of bench terraces

Soil conservation techniques to be employed to ensure closure success include

Modifying the soil slope

Maintaining and establishing natural vegetative cover

Securing favorable soil conditions to facilitate vegetative growth

Wildlife conservation techniques will include:





Providing food resources through establishment of vegetation

Providing habitats for wildlife through encouragement of vegetative proliferation, as well as

creation of aquatic habitats

Providing surface water and soils of good quality for wildlife consumption and use.

The approved environmental permit requires the submittal of a detailed closure plan for agency review

and approval two years in advance of closure.

20.7 Reclamation Measures during Operations and Project Closure

The facilities will be progressively closed over the duration of the mine site operations. Progressive

closure will reduce the costs of reclamation since closure will be integrated with the production

operations. In addition progressive closure will result in the development of expertise on the most

appropriate reclamation methods. Progressive closure will be undertaken, however without posing

impediments to day-to-day operations of the site. Final closure of the mine site will be undertaken

following completion of all mining operations, once treatment of site waters is no longer required, and

indications that further mining of the Toroparu Mine is not warranted.

Final closure of the facility will occur in two stages. The first stage will entail the following activities, if

not undertaken during progressive closure phases:

All fuel, chemicals, waste hydrocarbon products, and any potentially hazardous materials will

be removed from this site;

Water treatment will cease once runoff water no longer requires treatment.

During the second stage of the final closure all equipment, machinery, and storage tanks will be

removed for reuse or recycle. Where such uses are not practical, any remaining such materials will be

disposed of at a suitable storage site. All structures will be removed and/or be demolished. Structures

that are suitable for reuse or recycling will be salvaged. Structures not suitable for use will be

disposed. The Tailing Management Areas (TMAs) and other water management ponds will be closedand all disturbed areas will be reclaimed, with the exception of roads needed for monitoring access.

After the major closure activities are complete, a monitoring program may be implemented including

the site water quality monitoring and dams inspections.

The conceptual closure plan is intended to ensure the “return to nature” of the mine site. At the

conclusion of the closure process, no buildings or supporting infrastructure or facilities would remain

at the site. The areas will be fully replaced by a sustainable environment comprised of productive and

diverse lake and pond ecosystems. Spoil piles, stockpiles, borrow areas etc. would be vegetated with

general sustainable grass as well as emerging forest (primarily early stages in rainforest succession

are expected to dominate the period immediately following closure). The site will be monitored for

success of the closure plan. A few routes will be left for access to points of interest for the monitoringprogram. These routes will be closed after successful reclamation.

20.7.1 Tailings Management Area

The conceptual closure method for the tailings will include progressive closure of Cells 1 and 2. Any

exposed beaches will be reclaimed by placing a saprolite cover 1.0 m thick on the top of the tailings

beach and an additional layer 0.3 m thick of topsoil above the saprolite. The topsoil will be amended

by seeds and fertilizer. Rock armor will be placed on the side slopes. Diversion channels will be





redirected to flow back into the TMA. In the case of a precipitation event that exceeds design criteria,

discharge will be routed down spillways into the three collection ponds that will remain after closure.

The tailings facility design is based on a final slope that will be stable, thus no additional contouring of

the slope is anticipated at closure.

20.7.2 Open PitsPit banks must be structurally stable and barriers will be constructed around the parameter of the pit

for safety. A perimeter berm and fence will be placed around the open pit, and the earthen berm will

be revegetated. The berm will be to prevent vehicles from entering the pit except at a designated,

locked access point. The fence will be used to prevent unauthorized access. It is assumed that any

future pit lake will not require water quality management activities; however additional assessment will

be required to predict future pit lake conditions and the need for post-closure treatment or

management.

20.7.3 Waste Rock Storage Areas

The waste rock facilities will be covered with about 0.3 m of saprolite. The waste rock areas will beconstructed based on final slopes that will be stable, hence no additional contouring of the slopes is

anticipated.

20.7.4 Plant Site and Facilities

All plant and related facilities will be dismantled or demolished. Concrete slabs and footings will be

broken up to allow for infiltration, or will be placed into open facilities such as ponds or onto the waste

rock facilities. Power and water lines will be disconnected and will be removed from the site. Useful

major equipment and material will be salvaged and sold to third parties. Foundations will be removed

and excavated areas will be filled with native topsoil to restore naturally-sloping topography, where

feasible. Facility sites will be graded to blend in with existing topography, and compacted areas will be

ripped and the whole area will be covered with topsoil and revegetated. All topsoil areas will be

regraded and revegetated. Covers will be installed using saprolite available at site.

Several roads will remain to provide access to the property for closure and post-closure monitoring.

Internal roads will be leveled and graded to facilitate vegetation growth and re-establish drainage. All

exploration roads at the mine site will be reclaimed in a similar manner to haul and access roads. This

will include all areas outside the active mining area.

20.8 Closure Monitoring

Surface water and groundwater quality will be monitored after closure for evidence of environmental

impacts. Water samples will be collected annually to establish water quality trends. Physical

inspections will be conducted to monitor the physical stability of remaining facilities and the condition

of the closure covers and revegetation. It is anticipated that physical inspections will take place

quarterly and after significant storm events. Environmental monitoring is assumed to continue for ten

years, or until non-hazardous conditions are achieved for any discharge from the remaining facilities

and the groundwater and surface water quality meets applicable regulatory standards. Monitoring

records will be maintained by the mine operator.





20.9 Reclamation and Closure Cost Estimate

An allowance of US$16,162,501 for the final cost of reclamation and closure of the property has been

included in the cash flow projection.




Guyana

Figure 20.1.1.1

Location of Surface Water SamplesSource: ETK, 2013




Guyana

Figure 20.1.1.2

Location of Well SitesSource: ETK, 2013





21 Capital and Operating Costs [All]

21.1 Summary

The total estimated initial cost to design and build the Toroparu Project identified in this report is

US$464 million. A summary overview of the estimate by area is presented in Table 21.1.1.

Table 21.1.1: Summary of Capital Costs by Area

PFS Capital Cost Estimates(US$ Millions)

ScopeInitial Capital

(Pre-Prod)Expansion and

Sustaining CapitalLoM Capital

Fresh Rock Pre-Stripping SRK $24 $0 $24Mine Site Preparation / Roads SRK $2 $0 $2Mining Equipment SRK $69 $168 $237Milling Circuit Tt $75 $0 $75Leaching Circuit Tt $36 $0 $36Flotation Circuit Tt $24 $0 $24Process Plant Infrastructure Tt $6 $0 $6Plant Expansion Tt $0 $50 $50

Tailings Storage Facility Tt/KCB $16 $63 $80On-Site Infrastructure Tt $11 $11 $22Power Generation Tt $27 $0 $27Water Management Tt $9 $0 $9Camp and Ancillary Buildings Tt $25 $0 $25Port and Logistics FMG $9 $0 $9

Access Road Upgrades FMG $33 $0 $33Construction Indirects (incl. EPCM) Tt $79 $0 $79Owner's Costs (Incl. Closure) Sandspring $20 $15 $35

Sub-Total Project Capital Costs $464 $307 $771Mining Contingencies (Site Prep + Equip) SRK $4 $8 $12Process and Infrastructure Contingencies Tt/KCB/FMG $32 $0 $32Owner’s Costs Contingencies Sandspring $2 $4 $6

Total Contingencies All $37 $13 $50

Total Capital Requirement All $501 $319 $821Contribution from Saprolite Au Ore Margin All ($37) $0 ($37)

Total Project Costs w/ Contingencies All $464 $319 $784

The aggregate capital estimate is considered to be within a +30% / -25% weighted average accuracy

of actual costs. Base pricing will be in Q1 2013 United States dollars, with no allowances for inflation

or escalation beyond that time.

The contingency cost is based on the total direct and/or indirect costs and are included to account for

unanticipated costs within the scope of the estimate. The contingency percentage allowances vary

and are individually assessed based on the accuracy of the quantity measurement, type and scope of

work, and price information for the capital cost estimate.

The estimate is based on the cost of new equipment supported by budget quotes from vendors which

do not reflect discounts for negotiated prices, bulk purchasing, or used equipment purchases where

appropriate, any of which could lead to reductions in actual capital costs relative to the prices used in

the capital estimate.





Each contractor used their standard estimating system to calculate its respective capital costs for the

Project. The resultant information of the direct costs, together with the indirect costs is presented in

this section.

21.2 Capital Cost Estimate

21.2.1 Mining Capital Cost

Mining capital includes US$24 million of pre-stripping. Major mining capital costs were categorized

into mine equipment (estimated by SRK Consulting) and mine support facilities (estimated by Tt). The

mining equipment requirements were based on the mine production schedule, and estimates for

scheduled production time, mechanical availabilities, equipment utilization, and operating efficiencies.

Estimates of annual operating hours for each type of equipment were made, and equipment units

were utilized in the mining operations until a unit reached its planned equipment life, after which a

replacement unit was added to the fleet, if necessary. Additional equipment units and replacement

units were calculated as sustaining mining equipment capital costs. Major mining equipment rebuild

(overhaul) costs were not included in the mining equipment capital cost estimates, and were includedin the mining operating costs.

Capital cost estimates for major mining equipment (drills, loading equipment, haul trucks, dozers,

graders, etc.) were based on quotes from equipment manufacturers. Capital cost estimates for mining

support equipment were based on the current Infomine (mining) cost reference guide.

With the introduction of mining equipment in the second and third preproduction years (Years -2 and -

1), a cost estimate was made for the purchase of new spare parts, which was equivalent to 5% of the

total cost of the mining equipment purchased in those two years. A cost estimate was made for mine

shop tools, which was US$1.643 million.

The mining equipment capital cost estimate was based on the following:

All mining units are based on new equipment purchases;

Freight cost for major mining equipment was generally estimated between 4% and 7%;

No import duties were deemed to be applicable;

Allowances were made for on-site equipment erection costs for particular units;

Allowances were included for customization of the equipment (fire suppression, A/C, etc.);

Mining equipment rebuilds (overhauls) were included in mining operating costs;

The mining equipment capital cost estimate includes US$3.207 million for initial spare parts;

The mining equipment capital cost estimate includes US$1.643 million for shop tools;

The mining equipment capital cost estimate includes US$0.475 million for shop tools;

The total mining equipment capital cost estimate has an added 5% contingency; and The mining equipment capital cost estimate shown in this section does not include the mining

support facilities (infrastructure).

Table 21.2.1.1 shows the mining equipment capital cost estimate for Years -2 and -1, for the initial

capital costs.





Table 21.2.1.1: Initial Mining Equipment Capital Cost Estimate (US$000s)

Unit -2 -1 -2 -1 Total Initial

Equipmen t Units Make Model SizeCost

US$000sUnits Units

CostUS$000s

CostUS$000s

CostUS$000s

DrillingBlasthole drill - new Atlas Copco PV-235-D 165mm 2,000 2 3,999 3,999BlastingBlasting flatbed truck Scania G360CB 4X4 214 1 214 214

ANFO/Emulsion truck Scania P360CB 6X4 13 tonne 421 1 421 421Blasters crew truck Manufacturer 4x4 54 1 54 54Blasthole stem truck Scania P360CB 6X4 291 1 291 291LoadingFront end loader Caterpillar 988H 6.4 cu m 913 2 1,827 1,827Front end loader Caterpillar 993K 12.2 cu m 2,147Hydraulic excavator Caterpillar 390DL 5.0 cu m 1,117 3 4 3,351 4,468 7,819Hydraulic exc/shovel Caterpillar 6018EX/FS 10.0 cu m 3,284 2 6,568 6,568Hydraulic exc/shovel Caterpillar 6040EX/FS 22.0 cu m 7,495HaulingHaul truck - new Caterpillar 740B 40 tonne 889 4 3,554 3,554Haul truck - new Scania G460CB 10X4 50 tonne 390 9 34 3,512 13,267 16,779Haul truck - new Caterpillar 785D 133 tonne 2,478Other Mine EquipmentCrush/Screen Plant Manufacturer Jaw/Cone 335 kW 1,320 1 1,320 1,320Track dozer - new Caterpillar D9T 306 kW 997 4 3,988 3,988Wheel dozer - new Caterpillar 844H 468 kW 1,351 1 1 1,351 1,351 2,701Motor grader - new Caterpillar 16M 221 kW 1,134 2 1 2,268 1,134 3,402Backhoe loader Caterpillar 450E 102 kW 155 1 155 155Water truck - new Scania P410CB 8X4 30,000L 276 1 1 276 276 552Excavator - new Caterpillar 374DL 355 kW 786 1 1 786 786 1,572Compactor - new Caterpillar CS/CP 54 97 kW 166 2 332 332Support EquipmentTransport/mover Manufacturer Model 360 tonne 1,350 1 1,350 1,350Truck crane Manufacturer Model 120t crane 1,134 1 1,134 1,134Recovery truck Scania G460CB 8X8 360 kW 840 1 840 840Control blasting drill Atlas Copco SROC D65 110 mm 1,452 1 1,452 1,452Secondary blast drill Manufacturer 75 kW 64 mm 144 1 144 144Fuel/lube truck Scania P410CB 8X4 30,000L 482 1 482 482HD mechanic's truck Scania P360CB 6X4 310 1 310 310Flatbed truck Scania P360CB 6X4 19t crane 263 1 263 263Welding truck Manufacturer Model 75 1 75 75Tire service truck Scania P360CB 6X4 460 1 460 460Forklift Manufacturer Model 193 1 193 193Pit pumps/generators Flygt/Gen BS2290-434 82 kW 70 4 278 278Pit pumps & engines Godwin/Cat HL260M/C18 430 kW 138 2 277 277Personnel van/bus Manufacturer Model 64 5 321 321Service pickup Manufacturer 4x4 48 5 10 241 482 722Light plant Manufacturer Portable 8 kW 24 10 235 235Total Initial Mobile Equip 70 58 41,063 23,019 64,082Major equip initial spare parts 3,207 3,207Mine shops tools 1,643 1,643Tech services equipment 475 475Total Other Equipmen t 5,325 5,325Total Initial Mining Equip 46,388 23,019 69,407Contingency (5%) 2,319 1,151 3,470Total Initial Capital W/Cont 48,707 24,170 72,877

Pre-stripping

Mining capital includes US$24 million of pre-stripping. This cost is not mining equipment and is not

included in Table 21.2.1.1

Pre-Production Clearing

This cost is not mining equipment and is not included in Table 21.2.1.1. The mining capital estimate

includes US$0.25 million allocated to the clearing of the Toroparu Pit area, so that mining equipment

will be able to access the area to construct haul roads and commence mining of saprolite waste.





Initial Haul Roads

This cost is not mining equipment and is not included in Table 21.2.1.1. The mining capital estimate

includes US$1.50 million allocated to the construction of initial haul roads around the mine site area.

Approximately 5 km of haul roads will be required, mainly for the 50 t capacity haul trucks used in the

earlier years of production.

The mining equipment capital cost includes a jaw and cone crushing and screening plant, which will

provide crushed rock for road surfacing.

Subsequent (ongoing) development of haul roads, and upgrading of the roads later in the mine life for

the 133 t capacity haul trucks is included in the mining operating costs.

Mining Support Facilities

The following items are included in the mining support facilities and are part of the on-site

infrastructure costs:

Mine equipment maintenance shops;

Mine equipment wash pad; Mine equipment fuel storage and station; and

Explosives magazines.

The mining equipment capital estimate includes US$1.643 million allocated to maintenance shop tools

and equipment and is included in Table 21.2.1.1.

Engineering Equipment, Hardware & Software

The technical services department (geology, mine engineering and surveying) will require certain

equipment, computer hardware and software. The mining equipment capital estimate includes

US$0.475 million allocated to this.

21.2.2 Process and On-Site Infrastructure Capital Cost

Tt prepared a capital cost estimate to include the costs to design, construct, install, and commission

the mine and process support facilities, process plant, TMA facilities and associated on-site

infrastructure. The total costs for Tt’s design is estimated to be US$334.4 Million (Table 21.2.2.1). This

amount includes the direct field costs of executing the Project plus indirect costs associated with

design, construction, commissioning and contingency. Base pricing will be in Q1 2013 United States

dollars, with no allowances for inflation or escalation beyond that time.





Table 21.2.2.1: Initial Capital Cost Estimate – Process and On-site Summary

Area Total (000’s) Direct Costs

Overall Site $10,983Process Plant $140,419Power Supply $26,525

Water Management & Treatment $6,063Dewatering Flood Protection $3,037Tailings $16,422

Ancillary Buildings $24,772Direct Costs Subtotal $228,221

Indirect Costs Project Indirects $78,545Indirect Costs Subtotal $78,545

Contingency

Contingency Subtotal $27,677Total $334,443

Estimate Base, Date and ValidityThis estimate has been prepared in accordance with the Class 4 Feasibility Cost Estimate standards

of the Association for the Advancement of Cost Engineering International (AACE). The accuracy of

the estimate is +35%/-25%.

The quotes used in this study estimate are budgetary, sufficient to support a preliminary feasibility

study, and were obtained in Q4 2012. Quotations are non-binding. The estimate was prepared with

United States dollars as the base currency. Foreign exchange rates were not used.

Cost Basis

The direct cost estimate performed by Tt encompasses the costs associated with the overall site

infrastructure, the processing plant facility, power supply, the water management, dewatering andflood protection systems, and the tailings management area.

Tt worked in cooperation with Sandspring to identify key locations of site infrastructure, water

management placements, and the process design. Sandspring provided the baseline and input data

necessary for Tt to perform supporting calculations in the form of material quantities, for each of the

above items.

The processing plant consists of the comminution circuit, flotation and regrind circuit, pre-leach

thickening and CIP circuit, carbon elution and electrowinning, reagents and utilities.

The support infrastructure facilities provided include a change house, administration building,

laboratory, heavy and light equipment shops, warehouse facilities; truck wash facility, fuel dispensing,

ready line and explosive storage area.

The additional on-site infrastructure includes site development, roads, fresh water supply pond, yard

services, power generation and distribution, and fuel storage depot.

The tailings management area (TMA) consists of a tailings disposal pipelines, dam, water reclaim

pipelines, pump system and secondary containment facilities.

Environmental facilities consist of sewage treatment, water management dams and channels,

stormwater collection ponds and pumping stations or pass through gates.





The following entities have the responsibility for the above areas:

Tt: Complete ore processing facility with the exception of the comminution circuit. Tt is also

responsible for the on-site infrastructure water management protections and tailings and

water reclaim pipe lines;

Jacobs Engineering: Preliminary development and selection of the comminution circuit

inclusive of the primary crushing, ore storage and reclaim, secondary crushing, HPGR circuit.

The costs include all earthworks, foundations structural steel, and bulk commodities

consistent with a turn-key package. Tt adapted the HPGR circuit into its overall process

design; and

KCB: Complete TMA facility inclusive of all site development (cell and pond configuration),

dam structures, service roads and bridges, and internal water and tailings management

pipelines. The KCB developed quantities were provided to and priced by Tt.

Exclusions

The following items are excluded from Tt’s capital cost estimate:

working or deferred capital;

financing costs;

refundable taxes and duties;

land acquisition;

currency fluctuations;

lost time due to severe weather conditions;

lost time due to force majeure;

additional costs for accelerated or decelerated deliveries of equipment, materials, or services

resultant from a change in Project schedule;

warehouse inventories, other than those supplied in initial fills;

any Project sunk costs (studies, exploration programs, etc.); mine reclamation costs;

mine closure costs;

escalation costs; and

community relations.

21.2.3 Off-Site Infrastructure Capital Cost

Port Facilities

The initial capital costs for the port facility at Pine Tree Landing were developed by FMG, Inc. and port

sub-consultant Cargo Velocity, LLC, and are summarized in Table 21.2.3.1.

The port facility study considered three wharf concepts to observe the cost implications of differing

barge berthing options. All three wharf concepts facilitate the handling of the anticipated cargo types

and volume and apply the same backland infrastructure concept. For its lowest expected capital and

operating costs and lack of tidal range restrictions, a concrete ramp wharf concept is preferred and

considered in the port capital cost estimate.

The port capital costs, including contingencies are estimated to be US$10.4 million and include:

Backland infrastructure;





Wharf; and

Mooring facilities.

Excluded from the capital costs are:

Cargo container inventory;

Barge costs; Container chassis; and

Reefer mobile generators.

Access Road

The capital and operating costs to upgrade and operate the Toroparu Mine Road were developed by

FMG, Inc. through consultation with the Sandspring road maintenance crew supervisor. Sandspring

constructed the portion of the Toroparu Mine Road from Camp 4 to Toroparu in the mid-2000’s and

has been maintaining and upgrading portions of the roadway from Itaballi to Toroparu continuously

since 2011. Currently, Sandspring maintains a man camp near the Kumung River from which it bases

localized road maintenance operations which include: grading, placing aggregate surfacing, repairing

bridges, stabilizing soft sections and installing culverts. Thus, Sandspring has accumulated a direct

knowledge of the capital costs of roadwork along the Toroparu Mine Road.

The following components formed the basis of the capital cost estimate:

Equipment/Infrastructure;

Brush Back;

Culverts;

Single-Span Modular Steel Bridges;

Multi-Span Modular Steel Bridge;

Long-Span Modular Steel Bridges;

General Earthwork; Slope Stabilization;

Aggregate Surfacing;

Signage; and

Road Camp.

The initial capital requirement for the upgrade of the Toroparu Mine Road from the proposed port at

Pine Tree to Toroparu, including contingencies, is estimated at US$35.5 million. A contingency of 5%

was assigned to all direct costs for which Sandspring has demonstrated cost experience through their

current maintenance efforts. A contingency of 20% was assigned to the Furnishing and Placement of

Aggregate Surfacing. Although Sandspring has performed reconnaissance for laterite and fresh rock

occurrences and identified sufficient volumes, some variability is perceived in the processing andtransport costs for this item.

The Off-site Infrastructure Capital Cost Estimate is summarized in Table 21.2.3.1.





Table 21.2.3.1: Off-site Infrastructure Capital Cost Estimate (US$ mill ions)

Capital Cost Port Facility Access RoadDirect 7.4 31.5Indirect 1.5 1.3Contingency 1.6 2.7Total Capital Costs 10.4 35.5

Source: FMG, Inc., Cargo Velocity, Inc., Sandspring 2013

21.2.4 Owner’s Cost

Owner’s costs have been developed by Sandspring and the summary of these costs during the

construction period are shown in Table 21.2.4.1. The cost is broken down for the various core activity

areas and for the additional activities that are specific to the Toroparu Project.

Table 21.2.4.1: Major Components of Owner’s Costs ($US mill ion)

Item $US (milli on)Material Transport Equipment 4,350

Owner’s Development Team 4,705Construction InsuranceOwner’s G&A

7,2703,616

Contingency 1,912Total Owner’s Cost: 21,853

*All figures in million US$

Material Transport Equipment

The initial capital requirement for the fleet required to transport equipment and construction materials

to site from the proposed port at Pine Tree to Toroparu is estimated at US$4.35 million. A contingency

of 8% was assigned to the equipment purchase costs, for a total of US$4.7 million.

Transportation equipment requirements were based on an estimate of the number of trips required per

day for the transportation of fuel, containerized supplies (reagents, consumables, and materials),

during full production and acquiring these units in the first year of pre-production to provide

transportation for construction equipment, materials, and supplies to site.

Owner’s Development Team

This is a temporary department, during only the development phase of the Project. It is composed of

the Owner’s team, during the EPCM phase of the construction. Areas of responsibility include Project

controls management, construction management, engineering management, cost control, and

contracts. The role of this team ceases after operations start-up. The cost for this function is

US$4,705 million. An additional 10% contingency was added.

Construction Insurance

Insurance costs during construction are estimated at US$7,270,000 over the three year construction

period. The basis was a quotation received from a major North American construction brokerage firm.





Owner’s G&A

The majority of the Owner’s G&A Cost is due to Personnel in Administration functions, such as

management, finance, accounting, IT, supply chain, human resources, environmental, health, safety

and security and camp cost for all these functions during the construction period.

21.2.5 Sustaining Capital Costs

Mining Sustaining Capital

The LoM mining sustaining capital primarily consists of additional and replacement mining equipment.

The LoM mining sustaining capital estimate is as follows:

LoM additional and replacement (sustaining) mining mobile equipment - US$163.835 million;

LoM mining other equipment (large equip. spares parts, dispatch system) - US$4.218 million;

LoM sustaining mining equipment contingency (5%) - US$8.403 million; and

LoM total sustaining mining equipment including contingency - US$176.455 million.

These estimates are detailed further in this section. Mining sustaining capital costs occur between

Year 1 and 14. There are no mining sustaining capital costs in Year 15 and after.

A mine dispatch system capital cost estimate was included as sustaining capital in Year 2 of full

production, when the mining operations are planned to have expanded. With the introduction of larger

mining equipment in Year 6, a sustaining capital cost estimate was made for the purchase of

additional spare parts for this equipment, which was US$2.385 million.

Other mining related sustaining capital cost includes in Year 6 a third maintenance shop facility

planned to be constructed for the 133 t haul trucks, which will consist of four bays. This is included in

the Infrastructure sustaining capital cost.

As previously mentioned, the mining equipment rebuilds (overhauls) were included in mining

operating costs. Table 21.2.5.1 shows the LoM mining equipment sustaining capital cost estimate.





Process Systems Sustaining Capital Costs

The process systems sustaining capital is the costs associated with the expansion to the process

circuit and are inclusive of the following:

7,500 t/d gravity feed circuit – US$213,000;

CIP circuit area expansion – US$17,594,000; HPGR comminution circuit additional – US$30,943,000; and

Tailings pump expansion – US$793,000.

The total sustaining cost for process systems is estimated to be US$49,543,000. These sustaining

capital costs occur between production Years 3 and 5. There is no anticipated process system

sustaining capital costs after production Year 5.

TMA Facilities Sustaining Capital Costs

The TMA facilities sustaining capital is the costs associated with the expansion to the TMA cells and

containment ponds and are inclusive of the following:

Internal access road cell 1 – US$203,000; Internal access road cell 2 – US$1,997,000;

Internal access road cell 3 – US$415,000;

Cell 1 development – US$1,571,000;



Collection pond - cell 2 – US$2,736,000;

Collection pond - cell 3 – US$509,000;

Hydraulic deposition of soil cover – US$6,651,000; and

Pipelines & reclaim barges – US$14,466,000.

The total sustaining cost for TMA facilities is estimated to be US$63,356,000. These sustaining capital

costs occur between production Years 2 and 14. Table 21.2.5.2 shows the distribution of these costs

over this time frame. There are no anticipated TMA facilities sustaining capital costs after production

year 14.

Onsite Infrastructure Facilities Sustaining Capital Costs

The onsite infrastructure facilities sustaining capital is the costs associated with the expansion to the

main and satellite pits and mining operations. These costs are inclusive of the following:

Dewatering Pumps - Main Pit – US$4,651,000;

Dewatering Pumps - Satellite Pit – US$978,000;

Mine Shop Expansion – US$770,000; and Power Station Expansion – US$4,854,000.

The total sustaining cost for the onsite infrastructure facilities is estimated to be US$11,253,000.

These sustaining capital costs occur between production Years 1 and 16. Table 21.2.5.3 shows the

distribution of these costs over this time frame. There are no anticipated onsite infrastructure facilities

sustaining capital costs after production Year 16.




PC/MLM Toroparu_NI43-101_TechnicalReport_349800.020_044_ML

Table 21.2.5.2: TMA Facilities Sustaining Capital Cost Estimate

TMA FacilitiesSustaini ng Capital by Operating Year ($000s)

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

Area 4,021 13,758 4,498 4,930 18,481 6,472 6,298

Internal Access Road Cell 1 203Internal Access Road Cell 2 1,733 264

Internal Access Road Cell 3 318 Cell 1 Development 1,571Cell 2 Development 6,988 2,365Cell 3 Development 17,168 4,892 Collection Pond - Cell 2 2,736Collection Pond - Cell 3 509Hydraulic Deposition of Soil Cover 3,384 3,267Pipelines & Reclaim Barges 2,247 2,301 1,114 2,301 486 3,205 1,406

Values are rounded to the nearest thousand dollars

Table 21.2.5.3: Onsite Infrastructure Sustaining Capital Cost Estimate

Onsite Infrastructure Facilities

Sustaining Capital by Operating Year ($000s)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Area 5,506 280 770 296 326 815 3

Dewatering Pumps - Main Pit 280 296 815 3Dewatering Pumps - Satellite Pit 652 326Mine Shop Expansion 770Power Station Expansion 4,854

Values are rounded to the nearest thousand dollars





21.3 Operating Cost Estimates

21.3.1 Summary

The PFS estimate is based on independent reputable vendor quotations and where not available, first

principal calculations completed by the Consultants. Each contractor used their standard estimatingsystem to calculate its respective operating costs for the Project.

Operating costs have been prepared in Q4 2012 US dollars and exclude:

Contingency;

Escalation;

Taxes (VAT); and

Import Duties.

Imported equipment, materials, and operating supplies are not subject to Taxes (VAT), import or other

duties as per the Mineral Development Agreement.

The operating cost estimates have been assembled by area and component, based upon estimatedstaffing levels, consumables and expenditures according to the mine and process design. Life-of mine

operating costs are shown in Table 21.3.1.1, and annual operating costs in Table 21.3.1.2 (rounded to

nearest US$1,000).

Table 21.3.1.1: Operating Cost Life –of Mine, US$ x 1,000

Area Labor ($000s) Expenses ($000s) US$/t-Mined US$/t-Mi ll

Mine 159,656 946,003 1.86 8.70

Processing 22,238 1,313,509 n/a 10.51

G&A 92,515 81,306 n/a 1.37

Total Operating 274,408 2,340,818 n/a 20.57

n/a=not applicable

Table 21.3.1.2: Annual Operating Cost, US$ x 1,000

Year Mining Processing G&A Total $/t-milled

1 64,209 89,615 25,268 179,092 21.332 62,516 66,626 10,639 139,781 23.213 65,722 66,370 10,643 142,735 23.824 71,999 86,342 10,694 169,035 20.425 67,618 86,342 10,460 164,420 19.876 83,639 86,342 10,459 180,440 21.807 94,336 86,342 10,538 191,216 23.108 104,999 86,342 10,605 201,946 24.409 93,772 86,342 10,168 190,282 22.99

10 91,530 86,342 10,396 188,269 22.7511 90,175 86,342 10,195 186,713 22.5612 97,362 86,342 10,214 193,918 23.4313 67,749 86,342 10,130 164,221 19.8414 36,635 86,342 8,392 131,369 15.8715 7,052 86,342 7,395 100,788 12.1816 6,347 77,031 7,624 91,003 12.32Total 1,105,659 1,335,747 173,821 2,615,227 20.57

1 Year 1 mining costs also includes preproduction2. Includes stockpile re-handle





21.3.2 Mining Operating Costs

Mine operating costs were developed by SRK and based on the mine plan, equipment requirements,

and manpower requirements, parts of which have been presented in previous sections. The basis of

the operating costs is an owner operated mine. The mine operating costs include all the supplies,

parts, and labor costs associated with mine supervision, operation, and equipment maintenance

including rebuilds. Some pre-production development costs were included as part of the operating

costs. These costs may be treated as a capital cost during financial evaluation. Table 21.3.2.1

summarizes the total mine operating costs on a cost per tonne mined (from the pits), and total cost

per year basis.





Basis of Estimate

SRK estimated the required mining equipment fleets, required production operating hours, and

manpower to arrive at an estimate of the mining costs that the mining operations would incur. The

mining costs were developed from first principles. The mining operating costs are presented in the

following categories:

Production Drilling;

Production Blasting;

Production Loading;

Production Hauling;

Other Mine Operations (dozing, grading, dust suppression, other road maintenance

operations, etc.);

Support Equipment Operations (equipment fueling and maintenance, pit lighting, etc.);

Miscellaneous Operations (explosives storage operations, fleet dispatch operations, etc.);

Pit dewatering operations (pumps and piping);

Mine Engineering (mine technical personnel including technical consulting);

Mine Administration and Supervision (mine and maintenance supervision, etc.);

Mining Personnel Camp Costs; and

Freight (for equipment supplies and parts, excluding freight for fuel and explosives).

A maintenance cost was allocated to each category that requires equipment maintenance. The

operating costs are defined as starting in Year -2, and exclude any “pre-production operations”.

The mining costs are referenced as per tonne mined (waste and ore tonnes mined basis), and as per

ore tonne mined, (note the latter is not necessarily the same as per ore tonne processed in the same

year due to stockpile ore re-handling). By “per tonne mined” is meant as excavated from the open pits,

and does not include re-handled stockpile ore. This can lead to some distortion (increase) of the

“operating cost per tonne mined” (cost estimate) when significant amounts of stockpile material re-handling are carried out in a year, and relatively small amounts of pit mining occur in the same year.

Employee classifications, wages and burden benefits are based on information provided by

Sandspring. The costs for maintenance supplies and materials were based on estimates presented in

the current Infomine mining cost service publications.

It was assumed that the Toroparu Mine will not incur duties on imported equipment and supplies. No

account has been made for value added taxes (VAT), as these are normally accounted for in cash

flow modeling.

The mining operating cost estimates include the following parameters:

Diesel fuel cost of US$3.785/US gallon, or US$1.00/L (delivered to site); Average mining bench height of 10 m with ore mined in two 5 m flitches (lifts);

Average drilling penetration rate of 0.65 m/minute (instantaneous rate, no delays);

Blasting required for 88% of in-situ tonnage mined from the pit (12% free-digging);

Blasting powder factor of 0.233 kg/t (kg explosives per tonne of rock);

100% use of bulk emulsion explosives for blasting;

Averaged bulk emulsion cost of US$880/t (at site);

Average density for saprolite of 1.84 t/m3 and fresh rock 2.76 t/m3;





Average moisture content for saprolite of 20% and fresh rock 6%;

Average swell factor of mined saprolite 20% and fresh rock 40%;

Typical mining operations support equipment utilization of 3,600 to 4,600 operating hours per

year (for track dozers, wheel dozers graders etc.);

2% of plant ore feed re-handled in primary crusher stockpile;

Estimated average tire lives of:Wheel loaders – 3,500 operating hours

Haul trucks – 4,500 operating hours

Motor graders – 3,000 operating hours, and

Other major mining equipment – 3,500 operating hours .

7% freight cost on mining operating and maintenance supplies;

Mining equipment rebuilds (overhauls) were included in the mining operating costs. These

were estimated based on a total of 75% of the original cost of the equipment unit over the

operating life of the machine, and scheduled as three overhauls during the operating life; and

No contingency is included in the mining operating cost estimates.

The mine operating cost estimate includes the following:

All mine labor, salaried and hourly;

Blasting supplies and loading of explosives;

Consumables, fuel, parts, tires, etc.;

An allowance for mine related overheads;

An allowance for general operating expenses for the mine offices and explosives magazines;

and

All mine functions to deliver material to the dumps, stockpile, and primary crusher.

Excluded from the mine operating cost estimate are the following:

Initial clearing of the pit area and initial mining haul roads (in mining pre-production capital); Post mining reclamation costs;

Process related or crushing costs;

General overheads outside of the mine; and

Taxes and property holding costs etc.

Mine Labor

The mine department will have salaried staff for mine administration, supervision of mine operations,

supervision of mine equipment maintenance, and for technical services (geology and mining

departments). Many of these positions will be a permanent day shift. Hourly employees will fill mining

production, mining support functions, and mining equipment maintenance positions.

The maximum mine administration and operations supervision staff will total 16 positions, and the

technical services staff total 13 positions. Total salaried staff was planned to reach a maximum of 38

positions. Salaried staff requirements were estimated for both expatriate and national employees.

The operations, mine equipment maintenance, and technical services will include:

Mine administration will include the mine manager and secretary. The mine manager will be

an expatriate position up to Year 7;





Mine operations will include the mine general foreman, shift foremen, drill and blast foreman,

mine infrastructure foreman, mine supervisors, dispatch operator, cost controller, and training

supervisor;

Mine maintenance includes the maintenance foreman, senior maintenance supervisors, shift

foremen, and supervisors;

Technical services will include the technical services manager, secretary, and chief surveyor.The technical services manager will oversee the Mining and Geology departments, and will be

an expatriate position up to Year 7;

Mine geology includes the chief geologist, geologist, grade control engineer, and a

geotechnical engineer. The mine geologist will handle pit mapping, development drilling, and

other resource duties (such as local resource estimation and reconciliations). The grade

control engineer will supervise ore grade control in the mine. The geotechnical engineer will

be responsible for monitoring slope stability in the pits and waste dumps, as well as

monitoring material compaction and embankment stability at the TMA. The chief geologist will

be an expatriate position up to Year 7; and

Mine engineering includes the senior mining engineer, short- and long-term planners,

dispatch engineer (supervising mining equipment deployment), and drafts technicians.

Three mine production and maintenance hourly crews will be necessary (to be rotated on the two shift

system). Equipment operator labor positions were based on the number of mining equipment units

required, and on the assumption that some of the operators will be cross-trained. When some of the

operators are not required to be on one type of heavy equipment unit they will be able to operate

another. To maintain this situation it is planned for the mine department to have an equipment trainer

permanently on staff.

Operator positions were estimated for each year of operation. As mentioned, the number of

equipment operator labor positions was based on the number of mining equipment units required.

Required drilling, loading and hauling fleet equipment numbers were each rounded up to the nearestwhole unit required for a year, and each equipment unit was allocated three operators. This operator

estimate was adjusted up to allow for a 15% factor for vacation, sickness and absence (VSA). The

resulting operator estimate was then adjusted down (by 15%) to target meeting an average 85%

mechanical availability (MA) for the mining equipment units.

A mining equipment maintenance department will be staffed with mechanics, electricians, welders and

other maintenance personnel. Hourly maintenance man-hours were estimated as 70% of major

mining equipment man-hours required.

Mine total hourly labor requirements, including VSA and MA adjustments, are shown in Table

21.3.2.2. The hourly labor is divided into mine operations and mine maintenance. The peak number of

personnel occurs in Year 6, and the lowest number of personnel occurs in Years 15 and 16 when onlystockpile re-handling is taking place.

Annual salaries and annual (hourly paid) wages include burdens for the national staff personnel, and

the few expatriate staff planned.





Table 21.3.2.2: Mine Hour ly Labour Requi rements

Mine Hourly Labor - 2

- 1

1

2

3

4

5

6

7

8

9

1 0

1 1

1 2

1 3

1 4

1 5

1 6

BlastingBlaster 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Blasting laborer 6 6 6 6 6 6 6 6 6 6 6 6 6 6 6

Sub-Total Blasting 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7

Equipment Ops

Drill operators 3 9 9 9 12 9 12 15 15 12 12 12 15 6 3

Loading operators 9 12 35 38 35 38 38 38 29 29 26 18 18 21 12 6 3 3

Truck drivers 12 38 135 152 167 182 179 199 176 188 147 135 135 138 73 41 6 6

Other mine equipment 32 50 50 50 50 50 50 59 59 59 59 59 59 59 59 47 18 18

Support equipment 7 10 10 10 10 10 10 10 10 10 10 10 10 10 10 10 3 3

Sub-Total Equip Ops 60 113 239 260 271 292 286 318 289 301 254 233 233 242 160 107 30 30

General Mine Ops

General mine Ops 14 14 14 14 14 14 14 14 14 14 14 14 14 14 14 14 3 3

Grade Control Tech 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Surveyor 1 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 1

Rodman 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 1

Sub-Total Eq Ops 18 18 19 19 19 19 19 19 19 19 19 19 19 19 19 19 6 6

Total Mine Ops 77 138 265 285 297 318 312 344 315 326 280 259 259 268 186 133 36 36

Maintenance

Truck fleet mechanics 26 44 89 96 100 107 105 116 106 110 94 86 86 90 61 42 12 12

Load/spprt fleet mech 13 22 44 48 50 53 52 58 53 55 47 43 43 45 30 21 6 6

Field maint mech 13 22 44 48 50 53 52 58 53 55 47 43 43 45 30 21 6 6

Sub-Total Maint 51 89 177 191 200 214 210 232 212 220 187 173 173 179 122 85 23 23

Total Hourly 129 227 442 477 497 532 522 577 527 547 467 432 432 447 307 218 60 60

Source: SRKIncluding VSA and MA adjustments

21.3.3 Process Operating Costs

Process operating costs are estimated at US$1,336 million, or US$10.51/t-milled over the LoM, as

summarized in Table 21.3.3.1.

Table 21.3.3.1: Plant Operating Costs (LoM)

Cost ItemLoM Cost

($000s)Unit Cost$/t-milled

Labor $74,486 0.59Electric Power $583,725 4.59Reagents $323,727 2.55Consumables $231,626 1.82

Maintenance $101,818 0.80Tailings $20,364 0.16CIL Operating Cost $1,335,747 10.51

Source: Tt, 2013 Addition differences due to rounding





Labor

A small contingent of expatriate managers supplemented with trained nationals at supervisory

positions will be required for production and maintenance at the process facility and operation of the

tailings management facility. The operating and maintenance workforce will be Guyanese personnel.

Average all-in cost estimates for Guyanese labor were provided by Sandspring. During Pre-

Production, the process labor is estimated to be 91 personnel with an estimated annual cost of

US$2.30/t-milled. For Phase 1, the personnel requirement climbs to 136 persons with an estimated

annual labor cost of US$0.73/t-milled. For Phase 2, an additional 22 persons are added for a total of

158 persons with an estimated annual labor cost of US$0.87/t-milled.

Electric Power

Electric power will be generated on site to service mining, processing and all site service power

requirements using reciprocating engine generators burning intermediate fuel oil (IFO 180).

Power costs are estimated to be US$0.1769/kWh. The estimate is based upon a quoted fuel oil cost

of US$0.80/L ($0.1474/kWh) delivered and a 20% O&M cost of US$0.0294/kWh for operating the

generators.

Process power requirements are summarized in Table 21.3.3.2 for Pre-production, Phase 1 and

Phase 2 operations.

Table 21.3.3.2: Plant Power Requirements

Facility AreaPre-Production

@ 3.25k t/d(kWh/t)

Phase 1@ 15k t/d

(kWh/t)

Phase 2@ 22.5k t/d

(kWh/t)

Crushing 0.00 0.868 0.631HPGR/Milling 4.37 13.715 15.592Thickening 0.00 0.271 0.197Flotation 0.00 8.110 2.947CIP 8.42 1.658 2.411Elutions 1.11 0.218 0.301Tailings 3.83 1.652 2.406Reagents 0.08 0.124 0.090Water 0.42 0.553 0.402

Power Requirement 18.22 27.168 24.977

Source: Tt, 2013 Addition differences due to rounding

Reagents

Reagent consumption levels are based upon metallurgical test work conducted at SGS laboratories.

Unit costs for the major reagents consumed are based upon budgetary prices from suppliers or in-

house data. Consumption rates and estimated costs are presented by Phase in Tables 21.3.3.3

through 21.3.3.5.





Table 21.3.3.3: Reagent Consumption Rates and Cost – Preproduct ion

Reagent

Pre-ProductionSAP Leach

Cons.Rate

unit ReagentCost

Freigh t uni t Cost(US$/t)

1 Lime 1.33 kg/t $180 $19 per tonne $0.239

2 Sodium Cyanide 0.38 kg/t $3,000 $19 per tonne $1.1303 Sodium Hydroxide 0.10 kg/t $690 $19 per tonne $0.069

4 Flocculant 0.0 g/t $3,894 $19 per tonne $0.0005 Sodium Metabisulfite 1.06 kg/t $930 $219 per tonne $0.9866 Copper Sulfate 0.06 kg/t $3,880 $430 per tonne $0.2137 Carbon 40.0 g/t $3,810 $19 per tonne $0.152

8 Borax 0.15 kg/kg Cons $1,100 $410 per tonne $0.00029 Silica 0.15 kg/kg Cons $500 $174 per tonne $0.000110 Soda Ash 0.10 kg/kg Cons $550 $174 per tonne $0.000111 Potassium Nitrate 0.03 kg/kg Cons $1,650 $410 per tonne $0.0000

12 Diesel for Refinery/Carbon Kiln 182 L (000s) $0.82 per Liter $0.1261Total - - - - $2.790





Table 21.3.3.4: Reagent Consumption Rates and Cost – Phase 1

Reagent

Phase 1

15k t/d Flotation w/Tailings Leach

Cons.Rate

unitReagent

Costunit

Cost(US$/t)

Cons.Rate

unitR

1 A208 30.00 g/t $2,820 $19 per tonne $0.085

2 MIBC 50.00 g/t $3,910 $19 per tonne $0.1963 PAX 50.00 g/t $2,760 $19 per tonne $0.138

4 CMC 20.00 g/t $4,750 $219 per tonne $0.095

5 Lime 0.50 kg/t $180 $19 per tonne $0.089 1.29 kg/t

6 Sodium Cyanide 0.23 kg/t $3,000 $19 per tonne $0.702 0.34 kg/t

7 Sodium Hydroxide 0.02 kg/t $690 $19 per tonne $0.012 0.10 kg/t

8 Flocculant 40.0 g/t $3,894 $19 per tonne $0.156 40.00 g/t

9 Sodium Metabisulfite 0.10 kg/t $930 $219 per tonne $0.092 0.92 kg/t

10 Copper Sulfate 0.00 kg/t $3,880 $430 per tonne $0.000 0.05 kg/t

11 Carbon 6.68 g/t $3,810 $19 per tonne $0.025 40.00 g/t

12 Nitric Acid 0.02 kg/t $410 $19 per tonne $0.007 0.10 kg/t

13 Borax 0.03 kg/kg Cons $1,100 $410 per tonne $0.0000 0.15 kg/kg Cons

14 Silica 0.03 kg/kg Cons $500 $174 per tonne $0.0000 0.15 kg/kg Cons

15 Soda Ash 0.02 kg/kg Cons $550 $174 per tonne $0.0000 0.10 kg/kg Cons

16 Potassium Nitrate 0.01 kg/kg Cons $1,650 $410 per tonne $0.0000 0.03 kg/kg Cons 17 Diesel for Refinery/Carbon Kiln 114 L (000s) $0.82 $0 per Liter $0.0171 68 L (000s)

Total - - - - $2.790 - -





Table 21.3.3.5: Reagent Consumption Rates and Cost – Phase 2

Reagent

Phase 2

7.5k t/d Flotation w/Tailings Leach SAP Leach

Cons.Rate

unitReagent

Costunit

Cost(US/t)

Cons.Rate

unitReagent

Cost unit

Cost(US/t)

Cons.Rate

u

1 A208 30.00 g/t $2,820 $18.50 per tonne $0.085

2 MIBC 50.00 g/t $3,910 $18.50 per tonne $0.1963 PAX 50.00 g/t $2,760 $18.50 per tonne $0.138

4 CMC 20.00 g/t $4,750 $218.50 per tonne $0.095

5 Lime 0.50 kg/t $180 $18.50 per tonne $0.089 1.29 kg/t $180 $18.50 per tonne $0.232 0.84 k

6 Sodium Cyanide 0.23 kg/t $3,000 $18.50 per tonne $0.692 0.34 kg/t $3,000 $18.50 per tonne $1.025 0.55 k

7 Sodium Hydroxide 0.02 kg/t $690 $18.50 per tonne $0.012 0.10 kg/t $690 $18.50 per tonne $0.069 0.10 k

8 Flocculant 40.00 g/t $3,894 $18.50 per tonne $0.156 40.0 g/t $3,894 $18.50 per tonne $0.156 40.0 g

9 Sodium Metabisulfite 0.31 kg/t $930 $218.50 per tonne $0.288 0.92 kg/t $930 $218.50 per tonne $0.856 0.48 k

10 Copper Sulfate 0.00 kg/t $3,880 $429.50 per tonne $0.000 0.05 kg/t $3,880 $429.50 per tonne $0.186 0.03 k

11 Carbon 6.68 g/t $3,810 $18.50 per tonne $0.025 40.00 g/t $3,810 $18.50 per tonne $0.152 40.0 g

12 Nitric Acid 0.02 kg/t $410 $18.50 per tonne $0.007 0.10 kg/t $410 $18.50 per tonne $0.041 0.1 k

13 Borax 0.15 kg/kg Cons $1,100 $409.50 per tonne $0.0002 0.15 k

14 Silica 0.15 kg/kg Cons $500 $173.50 per tonne $0.0001 0.15 k

15 Soda Ash 0.10 kg/kg Cons $550 $173.50 per tonne $0.0001 0.10 k

16 Potassium Nitrate 0.03 kg/kg Cons $1,650 $409.50 per tonne $0.0000 0.03 k

17Diesel for Refinery/Carbon Kiln

13.88 L/t $0.82 $0.00 per Liter $0.0042 2.22 L/t $0.82 $0.00 per Liter $0.0249 166.28 L

Total - - - - $1.786 - - - - $2.742 - -





Consumables

Consumption estimates for crusher liners, ball mill liners, and ball mill balls are based upon:

Abrasion indices

Power consumption

Metallurgical test data.

The cost estimate for consumables is presented by Phase in Tables 21.3.3.6 through 21.3.3.8.

Table 21.3.3.6: Consumables Cost Estimate – Pre-production

Consumables

Pre-Production

SAP LeachCons.

Rateunit

Cons.Cost

Freight unitCost

(US$/t)

1 Crusher Liners 0.000 kg/t $5,500 per tonne $0.0002 HPGR Rolls 0.000 kg/t $5,000 per tonne $0.000

3 Ball Mill Liner 1.00 allowance $100,000 per set $0.0844 Ball Mill Balls 0.276 kg/t $1,256 $174 per tonne $0.3945 Filtering Consumables 0.00 allowance $1,000 per tonne $0.000

Total - - - - $0.478





Table 21.3.3.7: Consumables Cost Estimate – Phase 1

Phase 1

15k t/d Flotation w/Tailings Leach SA

Consumables Cons.Rate

unit Cons.Cost

Freight unit Cost(US$/t)

Cons.Rate

unit Cons.Cost

1 Crusher Liners - fixed perJacobs

kg/t fixed perJacobs

per tonne $0.027 0.000 kg/t $5,500

2 HPGR Rolls - fixed perJacobs


15% per tonne $0.725 0.000 kg/t $5,000

3 Ball Mill Liner - fixed perJacobs

set/yr fixed perJacobs

15% per tonne $0.314 1.00 set/yr $100,000

4 Ball Mill Balls - fixed perJacobs


15% per tonne $0.651 0.276 kg/t $1,256

5 Filtering Consumables - 0.10 allowance $1,000 per tonne $0.100 0.00 kg/t $1,000

Total - - - - - $1.818 - - -

Table 21.3.3.8: Consumables Cost Estimate – Phase 2

Consumables

Phase 2

7.5k t/d Flotation w/Tailings Leach SAP Leach

Cons.Rate

unit Cons.Cost

Freight unit Cost(US/t)

Cons.Rate

unit Cons.Cost

Freight unit Cost(US/t)

Cons.Rate

1 Crusher Liners fixed perJacobs


per tonne $0.037 0.000 kg/t $5,500 per tonne $0.000 fixed peJacobs

2 HPGR Rolls fixed perJacobs


15% per tonne $0.725 0.000 kg/t $5,000 per tonne $0.000 fixed peJacobs

3 Ball Mill Liner fixed perJacobs

set/yr fixed perJacobs

15% per set $0.514 1.00 set/yr $100,000 per set $1.370 fixed peJacobs

4 Ball Mill Balls fixed perJacobs


15% per tonne $0.651 0.276 kg/t $1,256 $174 per tonne $0.394 fixed peJacobs

5 FilteringConsumables

0.10 allowance $1,000 per tonne $0.100 0.00 kg/t $1,000 per tonne $0.000 0.00

Total - - - - $2.027 - - - - $1.764 -





21.3.4 Off-Site Infrastructure Operating Cost

Port Facilities

The port facility operating costs were developed by FMG, Inc. and port sub-consultant Cargo

Velocity, LLC with consultation from Sandspring for local labor rates. Considered in the operatingcost estimate were office administrative staff, costs of administrative and office buildings, facility

security. Not considered in the operating cost estimate were annual debt payments for capital

expenditures, re-grading of aggregate yard surfacing and barge operating costs.

The port facility study considered three wharf concepts to observe the cost implications of differing

barge berthing options. All three wharf concepts facilitate the handling of the anticipated cargo types

and volume and apply the same backland infrastructure concept. For its lowest expected capital and

operating costs and lack of tidal range restrictions, a concrete ramp wharf concept is preferred and

considered in the port operating cost estimate.

The annual operating costs for the Pine Tree Landing port are anticipated to be US$1.25 million, and

include labor, equipment, fuel and camp costs. The port facility annual operating costs aresummarized in Table 21.3.4.1.

Table 21.3.4.1: Port Facility Operating Cost Estimate

Cost Item US$’000Labor 542EquipmentFuel

78487

Camp 145Total Annual Operating Costs 1,252Source: FMG, Inc., Cargo Velocity, Inc., Sandspring 2013

Access Road

Although Sandspring provides localized maintenance of the roadway from a man camp near the

Kumung River, overall operation and maintenance of the roadway is currently administered by the

GoG via a contract with ETK and a local contractor. After upgrade of the roadway, the GoG will bear

the full operating costs associated with the Toroparu Mine Road.

21.3.5 General & Admin istrative Costs

The G&A for the Toroparu Project was estimated for a typical year of operation. The total cost varies

between US$7 to US$11 million per year or US$1.37/t processed on average over LoM. A large

portion of the G&A cost is the labor component. The basis for the G&A personnel structure is the

experience gained by Sandspring’s management from other operating mines.

The management department consists of the general manager and the operations manager. The

finance/accounting department includes tax and cost accounting and IT services for the entire

organization. The supply chain department includes purchasing, contracts and warehouse functions.

The EHSS department is designed to have four functions: safety, occupational health,

safety/emergency response and security for the gold room and the main entrances to the site.

Because of its remote location, Toroparu will have on site an elaborate first aid/trauma center with

doctors and paramedics. The environmental department forms an integral part of the operation

although structurally belonging to administration.





A portion of the G&A cost is the indirect cost component. It includes items such as safety supplies,

winter clothes, medical supplies, life insurance, general training, recreation and office supplies.





22 Economic Analysis

22.1 Method of Evaluation

A discounted cash flow model was created to evaluate the Toroparu Project assuming the Project is100% equity financed. All revenues and costs are expressed in US dollars.

Mining cost estimates were provided by SRK and process costs were provided by Tt. Offsite

infrastructure costs were provided by FMG and Owner’s cost by Sandspring. Additional costs such

as refining, royalties and administrative costs are also subtracted from the revenue to calculate an

estimated cash operating margin. The evaluation considers the following terms for the calculation of

doré and copper concentrate net smelter return.

Doré

Payable Gold: 99.9%;

Dore Transportation & Insurance: US$2.45/Au-oz;

Refining Charge: US$0.65/Au-oz.

Copper Concentrate

The copper concentrate of the Toroparu project will have a copper grade of 21% and will yield

significant quantities of gold, which could result into a scenario that gold is the major value

contributor of these concentrates. It will also present some content of bismuth and selenium

contents, what will result in some penalties. The cash flow model assumes concentrates will be bulk

shipped in containers to Europe, on a weekly or bi-monthly basis. The following are the assumed net

smelter return terms.

Payable Copper: 96.5% of the agreed analytical copper content, subject to a minimum

deduction of 1 unit; Payable Gold: 97% of the agreed analytical gold content subject to a minimum deduction of

1 g/dmt shall be paid for at the mean of the morning/afternoon London gold fixings in USD$

Concentrate Treatment Charges: US$95/t-concentrate;

Copper Refining Charges: US$0.065/lb-Cu;

Predicted Penalties: US$5.44/lb-Cu, based on the following:

o Selenium: 200ppm free; excess to be penalized at US$ 1.50/100ppm Se;

o Bismuth: 200ppm free; excess to be penalized at US$ 2.75/100ppm Bi and dmt of

material;

Copper Insurance: 0.167% of gross revenue minus deducts;

Gold Refining Charges: US$6.50/Au-oz; Gold Insurance: 0.167% of gross revenue minus deducts

Additional handling charge of US$25.00/dmt of material due to the characteristics of the material for

special treatment, handling, storage etc. (for container shipments). This is in addition to the ocean

freight charges of US$100/dmt of concentrate.

An income tax rate of 30% is assumed based on estimates provided by Sandspring. The resulting

cash flow in each year of the Project life is discounted back to January of Year -3 to determine the





estimated discounted cash flow at a 5%, 8% and 10% discount rate. Using this same data, the

estimated internal rate of return and the undiscounted cash flow were also determined.

The Prefeasibility makes use of Proven and Probable Reserves only.

Results of the base case analysis indicate that the Toroparu Project has a potential after-tax internal

rate of return of 23.1% and a present value of approximately US$690.1 million, based on a 5%

discount rate.


of the production period (from the start of Year 1). Figure 22.1.1 presents the behavior of the

accumulated free cash flow.

22.2 Input Parameters

The proposed Project including the open pits, processing facility and on-site and off-site

infrastructure would be developed by Sandspring with assistance from EPCM contractors and

suppliers. The contractors would assist Sandspring in port development and the construction of the

camp, processing, HFO power generation facility, tailings management area and other infrastructure.

The open pits would be developed by Sandspring using its own labor force and equipment. The

open pits, processing facility and on-site and off-site infrastructure, logistics including concentrate

transportation to the port, port operation and barge loading would be operated and maintained by

Sandspring using its own labor force and equipment with the assistance of equipment maintenance

specialists; geotechnical consultants, an explosive supplier; and other specialists. The key criteria,

principal assumptions and input parameters used in the Base Case are shown in Table 22.2.1.

The major input parameters to the model include gold and copper prices, sustaining capital,

operating costs, mining rates, and estimated taxes and royalties. Additionally, several minor

assumptions throughout the model such as working capital, environmental accruals and depreciation

rates affect the estimated Project economics to a lesser degree.

SRK and Sandspring prepared a detailed financial model (the Financial Model) estimating cash flows

by year for the forecast mine life.





Table 22.2.1: Key Criteria, Princ ipal Assumpt ions and Input Parameters Used in the BaseCase

Item Key Criterion / Principal Assumptions

Open pit operationOpen pit mining

method

Conventional open pit mining

Saprolite is excavated, loaded and hauledFresh rock is drilled, blasted, loaded and hauled

Total tonnes mined 596 MtTotal mill feed minedand processed

127 Mt (total tonnage of saprolite & fresh rock mill feed)

Total waste mined 469 MtSaprolite miningand processingTotal tonnes ofsaprolite mill feed

A total of 5 Mt saprolite mill feed are mined

When mined Saprolite mining operations are completed in Year 11Saprolite mining rate 2.4 Mt of saprolite mined and processed in Years 1 and 2

The mining rate varies in Years 1 to 11 (up to 0.48 Mt/y)When processed Saprolite processed in Years -2 to 16,

Saprolite processingrate

A maximum of 3,250 t of saprolite processed per day during the pre-productionperiod.

Fresh rock mi ningand processing

122 Mt of fresh rock mill feed consisting of 67 Mt of ROM fresh rock mill feed and 55Mt of low grade stockpile material

Total tonnes of freshrock mill feed

Fresh rock mined in Years -2 to 14.

When mined Up to 15 kt/d combined mill feed and low grade stockpile material mined in Years 1 to3. Mill feed and low grade stockpile material mined at rates ranging from 19.5 kt/d to22.5 kt/d in Years 4 to 16. These rates do not include waste rock mining.

Mining rate Fresh rock ROM material would be processed in Years 1 to 16.Fresh rockprocessing rate

15,000 t/d processed in Years 1 to 3.

Fresh rockprocessing rate

22,500 t/d processed in Years 4 to 16.

Gold revenueGold price $1,400/ troy oz AuGold reporting todoréGold recovery fromsaprolite

Saprolite gold recovery ranges from 88.2% to 98%. This recovered gold would reportto the doré bar product.

Gold recovery fromFresh rock

Fresh rock fed to the flotation circuit will report 43.5% of gold to doré, throughleaching of rougher tails and leaching of a gravity concentrate, while the leachingcircuit recovers 95.9% of gold in the fresh rock.

Projected goldpayable

99.9% payable.

Projected doré bartransportation,insurance andrefining costs

$3.10/oz Au

Gold reporting tocopper concentrate

It is estimated that 42% of the gold recovered from the rougher flotation stage wouldreport to the copper concentrate.

Gold payable 97% payableProjected goldtransportation andsmelter

The costs are assigned to the copper concentrate

Projected goldrefining charge

$6.5/ troy oz Au





Item Key Criterion / Principal Assumptions

Copper CreditProjected copper $3.25/lb CuProjected flotationCu recovery 91% CuProjected copper con

grade

21% Cu

Moisture content 8% moisturePayable deduction 3.5% of analytical copper contentProjected contransportation cost

$137.15/wet tonne of concentrate

Projected smeltertreatment charge

$95/dry tonne of concentrate

Projected copperrefining charge

$0.065/lb Cu

Royalties 8% gold royalty1.5% copper royalty

Operating costsMine saprolite $1.86/t-mined, including costs with rehandling and rebuilds.Processing $10.51/t processed (for processing, mill power and tailings disposal costs)

General andadministration cost $1.37/t processed

Escalation Costs and gold and copper prices are presented on real terms, based on Q1 2013

Depreciation method Straight line depreciation over five years

Projected corporatetax rate in Guyana

30%

The financial model assumes a three stage construction approach: the construction of small initial

startup for mining and processing of saprolite in Year -3 and the main grinding and floatation plant for

the fresh rock construction during Year -2 and -1. Commencement of processing is assumed at

January 1, Year -2 for the saprolite operation and at January 1, Year 1 for the fresh rock operation.

The processing rates for the stages described above are presented in Table 22.2.2.

Table 22.2.2: Project Stages

Description Value Units

Pre-Product ion Processing Rates

Saprolite Leach Daily Capacity 3,250 t/day

Fresh Rock Flotation Daily Capacity 0 t/day

Fresh Rock Leach Daily Capacity 0 t/day

Phase 1 Process ing Rates

Saprolite Leach Daily Capacity 1,500 t/day

Fresh Rock Flotation Daily Capacity 15,000 t/day

Fresh Rock Leach Daily Capacity 0 t/day

Phase 2 Process ing RatesSaprolite Leach Daily Capacity 175 t/day

Fresh Rock Flotation Daily Capacity 7,500 t/day

Fresh Rock Leach Daily Capacity 15,000 t/day

The model was also based on the following Project basic schedule:

Permitting and approvals: 1 year;

Construction period: 2 years;





Production period: 16 years.

The model revenue stream is supported by the following production parameters.

Table 22.2.3 Production Parameters

Description Value uni tsWaste Mined 468,875 kt

RoM Material Mined/Milled 127,111 kt

Saprolite 5,022 kt

Fresh Rock Flotation 51,780 kt

Fresh Rock Leach 70,309 kt

RoM Average Au Grades

Saprolite 0.91 g/t

Fresh Rock Flotation 1.17 g/t

Fresh Rock Leach 0.89 g/t

RoM Average Cu Grades

Saprolite - %

Fresh Rock Flotation 0.18% %Fresh Rock Leach - %

Plant Metallurgical Recoveries

Au Recoveries

Saprolite Circuit 96.1% %

Fresh Rock Flotation 85.5% %

Fresh Rock Leach 95.9% %

Cu Recoveries

Saprolite Circuit 0.0% %

Fresh Rock Flotation 91.0% %

Fresh Rock Leach 0.0% %

Doré Gold Production

Au Metal Content 2,914 koz

Copper Concentrate Production 414 kt

Au Metal Content 820 koz

Cu Metal Content 191,781 klb

Total Production

Au Metal Content 3,735 koz

Cu Metal Content 191,781 klb

Payable Gold 3,707 koz

Payable Copper 182,649 klb

Market Assumption

Gold Price US$1,400 $/Au-ozCopper Price US$3.25 $/Cu-lb

Depreciation – Depreciation of US$820 million during the life of the operation includes initial capital

of US$501 million, sustaining capital of US$319 million, and previously invested capital of US$105

million. The inclusion of sunk capital is important as it affects taxes in the determination of cash flow

during the life of the mine.





Start-up – For the purpose of the model, the plant is estimated to commence the processing of

saprolite ore on January 1st of Year -2. Fresh rock starts to be processed on January 1st of Year 1.

Working capital – Working capital was included in the model. This estimate was considered as 20%

of all operating costs for each period.

Taxation – A 30%corporate tax rate was applied over the life of the Project.

Escalation – The components of the economic model were based on the following:

Base capital pricing for the Project is in Q1 2013 United States dollars, with no allowances

for inflation or escalation beyond that time;

Equipment quotes from vendors were obtained in Q4 2012; and

Operating costs were prepared in Q4 2012 terms.

All financial results are based in Q4 2012 and Q1 2013 dollars and no escalation has been assumed

for the metal prices or cost inputs.

22.3 Cashf low Forecasts and Annual Production ForecastsBased on the parameters aforementioned, Project evaluation resulting economics present an after-

tax net present value of US$690 million, at 5% discount rate, and an internal rate of return of

23.14%. Table 3.1 presents further details of the economic results. Payback from plant start of

operations (January of Year 1) is 2.6 years.





Table 22.3.1: Project Evaluation Economic Results

Description Value UnitsMarket Prices

Gold $1,400 /Au-ozCopper $3.25 /Cu-lb

Estimate of Cash Flow (all values inUS$000s)

Gross Income $/Au-ozPayable Gold (Doré+Concentrate) $5,190,263 $1,400.00Payable Copper $593,609 $160Gross Income $5,783,872

$/Au-ozTreatment Charges ($39,353) ($10.61)Refining Charges ($18,937) ($5.11)Predicted Penalties ($2,253) ($0.61)Freight Insurance Cost ($71,641) ($19.32)Gross Revenue $5,651,687

$/Au-oz

Guyana Au Royalty ($413,937) ($111.65)Guyana Cu Royalty ($7,162) ($1.93)One Time Royalty to Surface Owner ($20,000) ($5.39)Net Revenue $5,210,588Operating Costs $/Au-ozMining Cost ($1,105,659) ($298.24)Processing Cost ($1,335,747) ($360.30)Site G&A Cost ($173,821) ($46.89)Total Operating ($2,615,227)$/t-ore ($20.57)Cash Cost ($/Au-oz) ($700)Operating Margin (EBITDA) $2,595,362Initial Capital ($501,192)

Total Capital ($820,651)Income Tax ($506,310)Cash Flow Available for Debt Service $1,268,400Pre-Tax IRR 27.19%Pre-Tax Present Value 0% $1,774,710Pre-Tax Present Value 5% $991,516Pre-Tax Present Value 8% $702,064

After-Tax IRR 23.14% After-Tax Present Value 5% $690,869 After-Tax Present Value 8% $476,171

Af ter-Tax Present Value 10% $367,345

The economic modeling resulted in a life of mine cash cost of US$700/Au-oz, as presented in thetable below.





Table 22.3.2: Summary of LoM Production and Cashflow

Type LoM ($000s) US$/Au-oz

Direct Cash Costs 2,143,915 578

Mining Cost 1,105,659 298

Processing Cost 1,335,747 360

Site G&A Cost 173,821 47Freight Cost 61,753 17

Treatment Charges 39,353 11

Refining Charges 18,937 5

Predicted Penalties 2,253 1

By-Product Credit (593,609) (160)

Indirect Cash Costs 450,986 122

Royalties 441,099 119

Freight Insurance Cost 9,887 3

Total Cash Cost 2,594,901 700

Table 22.3.3 shows annual production and revenue forecasts for the life of the project. All productionforecasts, ore grades, plant recoveries and other productivity measures were developed by

independent consultants.

Table 22.3.3 Project LoM Annual Production and Revenues

PeriodRoM

(kt)Waste

(kt)Ore Milled

(kt)Doré Gold

(koz)Cu Con. Gold

(koz)Cu Con. Copper

(klb)Revenue

(US$000s)

-3 0 0 0 0 0 0 (151,850)

-2 1,566 698 1,186 47 0 0 (164,030)

-1 5,863 7,137 1,186 36 0 0 (148,536)

1 13,623 26,375 6,023 146 129 29,519 250,614

2 8,000 31,999 6,023 113 99 25,293 165,209

3 8,164 32,132 5,992 105 91 22,164 76,875

4 7,981 31,371 8,276 263 36 8,484 151,754

5 8,962 31,907 8,276 197 60 11,051 110,645

6 5,899 41,855 8,276 172 37 9,116 38,237

7 8,194 44,385 8,276 194 60 10,574 84,655

8 19,695 37,200 8,276 290 64 7,853 176,039

9 1,600 45,490 8,276 152 30 8,503 34,918

10 11,711 33,335 8,276 232 38 8,666 131,434

11 12,725 37,428 8,276 286 49 8,198 175,775

12 2,255 48,125 8,276 152 19 8,239 31,548

13 4,499 17,301 8,276 149 36 10,149 73,972

14 6,374 2,139 8,276 189 36 8,236 126,982

15 0 0 8,276 101 18 8,217 57,668

16 0 0 7,385 89 17 7,519 28,29117 0 0 0 0 0 0 18,201

Total 127,111 468,875 127,111 2,914 820 191,781 1,268,400

22.4 Sensit ivity Analysis

A sensitivity analysis for key operating and economic parameters is shown in Tables 22.4.1 through

22.4.3 and Figures 1.14.1 and 22.4.1. The Project is most sensitive to gold price. Because the

change in NPV is greatest with gold price change, it is this parameter that most significantly affects





the economics of the Project. Additionally, sensitivity to discount rate has been simulated and is

presented in Figure 22.4.2.

Table 22.4.1: Sensitivity to Capital Costs

Capital Costs Sensitiv ity Base 80% 90% 100% 110% 120%

After-Tax NPV 5% 690,869 787,312 739,104 690,869 642,634 593,883 After-Tax NPV 8% 476,171 567,303 521,756 476,171 430,586 384,289

After-Tax NPV 10% 367,345 455,516 411,453 367,345 323,238 278,323

IRR 23.14% 29.07% 25.86% 23.14% 20.80% 18.72%

Table 22.4.2: Sensitivity to Operating Costs

Operating Costs Sensitiv ity Base 80% 90% 100% 110% 120%

After-Tax NPV 5% 690,869 922,918 806,893 690,869 574,844 458,260

After-Tax NPV 8% 476,171 658,171 567,171 476,171 385,171 293,426

After-Tax NPV 10% 367,345 523,935 445,640 367,345 289,051 209,928

IRR 23.14% 27.62% 25.43% 23.14% 20.73% 18.10%

Table 22.4.3: Sensitivi ty to Metal Prices

Revenue Sensi tivit y Base 80% 90% 100% 110% 120%

After-Tax NPV 5% 690,869 203,584 448,392 690,869 932,182 1,172,911

After-Tax NPV 8% 476,171 87,525 283,383 476,171 667,357 857,707

After-Tax NPV 10% 367,345 29,065 199,899 367,345 532,976 697,634

IRR 23.14% 11.18% 17.57% 23.14% 28.13% 32.70%



(160)

(120)

(80)

(40)

-

40

80

120

160

200

240

280

320

360

400

(560,000)

(420,000)

(280,000)

(140,000)

-

140,000

280,000

420,000

560,000

700,000

840,000

980,000

1,120,000

1,260,000

1,400,000

Cumulative Free Cash and NPVs

Cumulative FCF

Project Capital

Total Payable Gold


Guyana

Figure 22.1.1

Cumulative Cash FlowSource: SRK, 2013



0.00%

5.00%

10.00%

15.00%

20.00%

25.00%

30.00%

35.00%

80% 90% 100% 110% 120%

IRR Sensivity (%)

Capital Costs

Operating Costs

Revenue


Guyana

Figure 22.4.1

IRR Sensiti vitySource: SRK, 2013



(200,000)

0

200,000

400,000

600,000

800,000

1,000,000

1,200,000

1,400,000

0 %

1 %

2 %

3 %

4 %

5 %

6 %

7 %

8 %

9 %

1 0 %

1 1 %

1 2 %

1 3 %

1 4 %

1 5 %

1 6 %

1 7 %

1 8 %

1 9 %

2 0 %

2 1 %

2 2 %

2 3 %

2 4 %

2 5 %

2 6 %

2 7 %

2 8 %

2 9 %

3 0 %

NPV vs Discount Rate

NPV


Guyana

Figure 22.4.2

NPV sensit ivity to Discoun t RateSource: SRK, 2013





23 Adjacent PropertiesSandspring is not aware of any significant properties situated immediately adjacent to Toroparu.





24 Other Relevant Data and InformationThere is no other additional information or explanation necessary to make the technical report

understandable and not misleading.





25 Interpretation and Conclusions

25.1 Geology and Resources

The geology of the Toroparu gold-copper deposit is defined primarily from drill core. The depositgeology is a network of veinlets and fractures hosting quartz and calcite and minor amounts of pyrite

and chalcopyrite that contain the gold and copper mineralization, hosted in a WNW to E-W structural

zone that forms the deposit. The Toroparu deposit is sufficiently drilled for feasibility level study.

Drilling and sampling procedures, analytical procedures, and QA/QC programs and results establish

the drillhole database that supports mineral resource estimation as credible, verifiable, and

developed by appropriate common industry best practices.

Since the 2012 PEA resource estimation (P&E, 2012), the drillhole database has been augmented

with an additional 214 drillholes for 56,259 m, or an increase of 38% in total meters of drilling. A total

of 48 of the holes, for 12,163 m, in both the Main Zone and the Southeast Zone, were for targeted

infill drilling designed to both increase confidence in the estimation and to provide a better definitionof mineralized zones. The majority of the remaining 166 holes were for delineation drilling. The

additional sampling was effective in upgrading confidence in the resource estimate. Additional

information has allowed the development of internal controls delineating mineralized versus non-

mineralized areas of the overall deposit.

Mineral resources have been updated for this prefeasibility level technical report. The Toroparu

deposit contains approximately 240.2 Mt of Measured and Indicated mineral resources grading 0.89

g/t Au and 0.084% Cu for 6.894 Moz of contained gold and 444 Mlb of contained copper, at a 0.30

g/t Au cut-off grade, in all mineralized zones. In addition there is an additional 129.5 Mt of Inferred

mineral resources grading 0.74 g/t Au and 0.042% Cu, for 3.090 Moz of contained gold and 120 Mlb

of contained copper, at the same 0.30 g/t Au cut-off grade.Within the prefeasibility design pit shape, there is 97% Measured and Indicated resources.

Therefore, additional targeted drilling to convert Inferred resource’s to Measured and Indicated is

possible, but in SRK’s opinion not necessary to advance the Project to full feasibility level study. The

resource estimate, therefore, is deemed sufficient for feasibility level.

Additional step-out definition drilling can be done to explore the eastern extension of mineralization in

the main pit area; a region of shallow placer working that is low in topography and partially flooded.

The potential exits in this area for expanding the pit to the east with incremental additions of low

grade material. The Southeast Zone is open at depth and to the southeast, and is a target for

exploration drilling – access is difficult due to the proximity of the river. These areas of additional

exploration drilling are warranted, at some point in time, but are not in SRK’s opinion necessary toadvance the Project to full feasibility level study.

25.2 Mining and Reserves

Pit Slope Geotechnical

The PFS pit slope design for the Toroparu Project was based on the currently available geotechnical

data and the geological model. This design mainly focuses on the proposed open pit in the main

deposit area. The stability analyses confirm that the recommended pit slope angles are reasonable





and appropriate. This design has a number of operational constraints including requirements for

careful blasting and for effective management of surface run-off water and seepage. Extensive

monitoring and ongoing commitment to data collection are also recommended throughout the

operational life of the mine.

The 2010 field program did not cover the SE Pit area. Additional data collection and further studieswill be required to carry the Project to a feasibility level of design, however, the geologic conditions

are similar and this pit is a likely proxy for the Toroparu main deposit. The SE Pit contributes 6.8% to

the Proven and Probable reserves and it is commissioned after the 4th year of full production.

Mining

The PFS mining studies were based on multiple trade-off studies to arrive at an optimum production

schedule for the Project. Some refinements with respect to the pit designs, mine production

schedule, and mine equipment selections can be made at a feasibility level study. The current PFS

mining studies exceed the requirements of a prefeasibility level study.

25.3 Metallurgy, Processing and RecoveriesSandspring conducted several metallurgical testwork programs between 2009 and 2013 which

provided the basis for process design to prefeasibility level.

Sandspring has provided a list of drillholes and intervals of samples collected for metallurgical

testwork performed between 2011 and 2013, along with a 3-D plot showing the location of the

samples within the deposit. Assays of drill core intervals show that low, medium and high grade

portions of the ore body are represented in the sample set. The samples appear to be spread

geographically across the ore body. Although more sampling is recommended to cover variability

metallurgical testwork to support a feasibility study, the samples appear to be representative and

sufficient for the Prefeasibility study.

Testwork showed that multiple processes are warranted to provide economic benefit to different ore

variabilities in the deposit. Metallurgical testwork and financial analysis tradeoffs were performed to

show that processing the deposit with both flotation and cyanide leaching, depending on Cu content,

would provide economic benefit due to the recovery of a marketable Cu concentrate.

Ores containing a higher Cu content responded more economically to gravity concentration followed

by Cu flotation and cyanide leaching of the cleaner flotation tailings. Overall recoveries of this

process are expected to be 91% for Cu and 88% for Au.

Ores with low Cu content responded more economically to gravity concentration followed by cyanide

leaching of the gravity tailings. Overall recoveries of this process are expected to be 96% for Au.

Gold bearing saprolite ores showed no benefit from gravity concentration but did respond best to

whole ore cyanide leaching. Recoveries of this process ranged between 88% to 98% depending on

the retention time in the leach circuit, which varies throughout the LoM.

Process facilities were designed to achieve the stated recoveries based on test results and standard

engineering design practices to a prefeasibility level. Process facilities include comminution circuits

consisting of primary/secondary crushing circuit feeding into one of two HPGR/ball mill grinding

circuits. Each circuit is followed by Cu flotation and/or cyanide leaching. Tailings from the process





facility will be treated through a cyanide destruction circuit prior to discharge into the TMA facility.

Products from the process facility include gold bearing copper concentrate and gold doré.

25.4 Infrastructure


On-site infrastructure was identified and designed sufficient to support the PFS report. Internal

service roads, bridge crossings, and operational support facilities were developed to a point sufficient

for costing and use in the economic model. Power is generated to support the operation. Water is

sourced from the Puruni River and Wyanmu River. An extensive water management plan was

performed to assure year round operation.

Off-Site Infrastructure

Considering the relatively remote location of the Toroparu Mine, off-site infrastructure will play a key

role in the operation and success of the Project. The presence of an existing roadway from the

Mazaruni River to the mine is advantageous, and upgrades to the roadway will ensure safe, reliable

transport of cargo to and from the mine. Overall operation and maintenance of the roadway is

currently administered by the GoG via a contract with a local contractor. After upgrade of the

roadway, the GoG will bear the full operating costs associated with the Toroparu Mine Road.

Construction of the port facility at Pine Tree Landing will create a secure shipping point for mine

equipment, bulk mine supplies and concentrate, and will serve as a base camp for mine supply chain

transport vehicles.

25.5 Project Implementation

The Project schedule would be an estimated duration of 36 months from the award of the EPCM

contract for detailed engineering, to the mechanical completion of the processing plant, through

commissioning activities and into production.

The schedule critical path, which determines the overall Project duration as determined by the

shortest logical chain of related events, must be followed to ensure completion of the Project by the

scheduled date. For the process plant construction of Toroparu Gold Project, it is estimated that the

critical path will run through the substation installation, electrical procurement, and electrical design.

Critical milestones included Sandspring’s ability to obtain Project financing, award of EPCM contract,

as well as the associated pre-development work in the open pit and tailings area.

Project approval and the beginning of pre-development construction activities are scheduled to occur

right away in order to maximize the usable construction seasons. It is imperative that road access

into the various parts of the site be completed early in order to facilitate the transport of constructionmaterials and equipment to their work areas.

Three engineering disciplines must be started right away. The first is geotechnical in order to confirm

site conditions and provide the civil engineers with necessary information for the early detailed

engineering. The second and third are the process and mechanical in order to facilitate plant design

for the early ordering of long lead time equipment.





There are a number of pieces of equipment in this Project which typically have long delivery times

associated with them, which will need to be addressed as priorities. The longest delivery items

should get priority for placing purchase orders to support the construction schedule.

This Project schedule is driven by construction, and construction in turn is driven by site conditions.

There are three key activities which must start on schedule and which will dictate the success of

bringing this Project in on time. These activities are:

Primary power;

Process plant design;

Site infrastructure; and

Roads.

25.6 Environmental Studies and Permitting

The Project area has been historically impacted by mining activities, logging, and hunting.

With only a few exceptions, species classified as rare, threatened or endangered have not beenobserved in the Project area.

There are no formal or established communities in the immediate vicinity of the site. The Project is

not expected to generate many direct socio-economic impacts. A Social Management Plan has been

proposed to mitigate the socio-cultural impacts identified in the ESIA.

No indigenous hunting activity or cultural resources were identified within the proposed mining area.

Results of the geochemical testing of the waste rock showed that the waste rock lithologies and LGO

samples contained very low sulfide-sulfur concentrations, indicating low risk of PAG, except for the

saprolite. The saprolite and transition zone samples contained very low NP, whereas the waste rock

and LGO had NP related to reactive carbonate minerals. The saprolite samples were classifiedprimarily as acid-generating and PAG, whereas the other waste rock and LGO samples were

classified as non-PAG.

The tailings samples contained low to negligible sulfide-sulfur concentrations and were classified as

non-PAG. The majority of the NP of the tailings was associated with the reactive carbonate minerals

and/or lime added during the metallurgical testing. The saprolite tailings contained little to no reactive

carbonate minerals, and thus the NP present in the saprolite tailings was related to the lime added

during the metallurgical process.

Leachate testing indicated that the waste rock may develop alkaline drainage with the possibility of

elevated concentrations of aluminum, selenium, chromium and, to a lesser extent, copper and

phosphorus. The tailings could develop alkaline drainage with the possibility of elevatedconcentrations of aluminum, selenium, chromium, arsenic, cobalt, copper, iron, molybdenum, WAD

cyanide and sulfate. The TMA design assumes that the natural low permeability of the surficial soils,

and the lower concentrations of elements in the TMA pond due to attenuation from natural

degradation, settling, and mixing with precipitation, which averages about 2.6 m annually, will reduce

concentrations in any TMA discharge effluent to the aquatic receiving environment. Additional

analysis (i.e., predictive water quality modeling) will be needed in a later phase to verify this

assumption.





The Project will develop and implement an Environmental Management Plan.

An ESIA was prepared and submitted to the Guyana EPA, which subsequently issued an

environmental permit for mining and processing. The final mining permit will be required prior to

commencing full-scale operations.

25.7 Economic Analysis

A discounted cash flow model was created to evaluate the Toroparu Project assuming the Project is


An income tax rate of 30% has assumed. The resulting cash flow in each year of the Project life was

discounted back to January of Year -3 to determine the estimated discounted cash flow. Using this

same data, the estimated internal rate of return and the undiscounted cash flow were also

determined.

Using a gold price of US$1,400/oz and a copper price of US$3.25/lb, results of the base case

analysis indicate that the Toroparu Project has a potential after-tax internal rate of return of 23% and

a present value of approximately US$690 million, based on a 5% discount rate.


of the production period (from the start of Year 1).

These positive results indicate that the Project should be advanced to a Feasibility Study.

25.8 Risks and Uncertainties

Mining

The following potential risk aspects should be further assessed during a feasibility study for the

Project:

Determination of peak water inflows to the pit at various stages of mine development to

ensure adequate pit dewatering capacity exists to achieve close to continuous operations;

Assessment of appropriate pit development methods (including drainage) and mining fleet

equipment selection for efficient mining of the significant thicknesses of saprolite material (up

to approximately 75 m); and

Detailed mine production scheduling to ensure continuous ore exposure in the pits,

particularly during periods of low ore stockpile inventories.

Mineral Processing

A full risk assessment of the transportation of reagents and consumables to site and concentrates to

the port facility should be conducted to determine any logistics issues given the plant site location.

Whilst there is no shortage of water on-site, a clean source of fresh water has yet to be finalized, and

should be identified early in the next phase of the Project.

On-site Infrastructure Risks

The site locations for plant facilities and man camp do not have condemnation drilling. There is a

potential that resource could be found under these locations that could potentially necessitate

relocation of this infrastructure.





The cost estimate has been developed based on maximizing the use of local borrow sources and

reduced construction material haul distances to help lower the construction costs. Increase of

material haul distances will negatively impact the construction costs.

Tailings Management Area

The following potential risk aspects exist for the TMA:

The key risk at the tailings storage area is the absence of ground information. The designs

presented above are only conceptual until it is shown the ground can support the proposed

dam structures. It has been assumed that the site presents a thick horizon of low

permeability residual soils, which is key to minimize seepage into the groundwater regime

and to eliminate the need of a geomembrane liner along the pond basin;

The climatic information at the site is sparse so the sizing of diversion ditches, spillways, and

freeboard are rough estimates at this time. Climatic data needs to be collected now and

throughout the life of the Project; and

A key assumption is that the surface soils within the tailings pond will form an effective

barrier to contaminants exiting the impoundment area as was experienced at the second

Omai tailings impoundment. This cannot be confirmed without a ground investigation

program at the tailings area.

Off-site Infrastructure

The use of trained, reliable local labor has been assumed in developing the roadway construction

costs and estimated schedule of work. Therefore, the unavailability of trained, reliable labor would

negatively impact construction costs and may cause delays in completion of the upgrades.

Also, the estimated roadway upgrade cost was based on accessible local aggregate surfacing

sources which have been preliminarily identified by Sandspring. The logistics to access, process and

transport these materials may vary somewhat from those assumed in the development of the costestimate. To hedge against this variability, a 20% contingency was assigned to this item in the

upgrade capital costs.

Overall operation and maintenance of the roadway is currently administered by the Government of

Guyana (GoG) via a contract with a location contractor, and the GoG will bear the full operating costs

associated with the Toroparu Mine Road after upgrade of the roadway. If the GoG should

discontinue its participation in the roadway, Sandspring will be required to facilitate some

maintenance of the portion of the roadway necessary for operation of the Toroparu Mine.





26 Recommendations

26.1 Recommended Work Programs and Costs

Phase I is the fieldwork and testwork necessary in order to complete a feasibility study, with Phase IIbeing the actual feasibility study having all fieldwork and testwork available.

26.1.1 Phase I

Open Pit Geotechnical/Hydrogeological Programs

The PFS pit slope design for the Toroparu Project was based on the currently available geotechnical

data and the geological model, which mainly focused on the proposed open pit in the main deposit

area. The 2010 field program did not cover the SE Pit area, and additional data collection is

recommended. Recommendations for future pit geotechnical/hydrogeological programs to carry the

Project to feasibility are below:

Complete additional geotechnical investigation at the proposed satellite pit; and

Evaluate the hydrogeological characteristics and pit dewatering requirements (shallow well

pumping tests for dewatering design).

The costs for the above pit geotechnical/hydrogeological programs are estimated below:

Oriented core drilling and testing at South-East deposit (3 holes, 200 m deep), with

mobilization/demobilization, special equipment, materials and testing: US$170,000; and

Engineering time and travel charges for field supervision and senior supervision site visit:

US$100,000.

The total cost for pit geotechnical/hydrogeological programs is in the order of US$270,000.

Drilling/Geology/Resources:

There is no need for additional in-fill or step-out drilling for Phase I recommended work to advance

the Project to completion of feasibility level study.

Mining

No mining related fieldwork or testwork will be required in advance of a feasibility study.

Condemnation Drilling

Sandspring intends to perform a planned condemnation drilling program for the Toroparu Project

site, to include sites for the waste dumps, stockpile, process plant, site infrastructure, tailings

management area, mine village, etc. The two main site locations would be the mine facilities area(dumps, stockpile, plant, etc.) and the tailings management area (approximately 8 km to the

northeast). It is proposed to dill a total of 16,000 m of drilling consisting of drillhole lines spaced

either 500 m or 1,000 m apart, with drillholes placed at 100 m intervals along these lines. Drillholes

would be drilled at 45 degree inclinations to a depth of 70 m. An allowance for an additional 2,000 m

of drilling has been included to further investigate areas where drill results indicate mineralization.

The total drilling program may take seven months.

The estimated costs for the condemnation drilling program are shown in Table 26.1.1.1.





Table 26.1.1.1: Condemnation Drilling Program

Condemnation Drilling RAB RigDrillingLength

(m)No. of

Samples

Cost per Unit

($/unit)Total Costs

(US$s)

RAB Drill Costs 16,000 $30/m $480,000 Laboratory Costs (RAB Samples) 900 $24/sample $138,240 Drill Fuel 96,000 Road and Pad Preparation

Fuel, Oil, Lubricants 80,00 Maintenance and Spare Parts 179,200

Technical Personnel 89,600 Local Team Wages (4 RAB) 118,400 Sub-totals 16,000 900 $1,181,440

Additional Investigative Drilling 2,000 150 $148,560Totals 18,000 1,050 $1,330,000

Source: Sandspring, 2013Costs exclude existing camp operating costs, and transport costs to and from site of personnel and materials.

Metallurgy and Processing

Testwork on the ACO and LCO composites has shown that a significant amount of copper is loaded

onto the carbon following cyanide leaching. To limit the amount of copper loading on the carbon, the

initial cyanide concentration may need to be increased. The higher cyanide concentration will likely

increase the extraction of copper and we recommend further optimization testing to further define the

design parameters. It is also recommended that a Gravity Recoverable Gold (GRG) test be

conducted to further define the effect of grind size on the recovery of gold. Flotation testwork

conducted during this program has focused on evaluating the effect of grind size on recovery, but

additional study will be required to finalize a flotation circuit flowsheet. The results from the locked

cycle test of the ACO Master Composite indicated that there is a build-up of non-sulfide gangue in

the cleaning circuit. It is recommended that further flotation testwork be conducted to betterunderstand the effect of cleaner circuit flowsheet and use of gangue depressant, CMC, and other

reagents and flotation flowsheets to potentially improve the process.

We recommend additional variability testing with regard to gravity separation, copper flotation and

cyanide leaching on the LCO composite. The testing should identify the gold response in variation to

head grade. .

A pilot scale HPGR test, in which 2 tonnes of material is tested, is recommended to properly size the

HPGR unit, as the relevant testwork to date only provides preliminary sizing information on a scoping

level. This should be performed in conjunction with variability comminution testing to provide better

selection of ball mills.

Due to the number of thickeners present in the proposed flowsheet, additional thickening tests are

recommended to properly size the thickeners. These tests should evaluate the following:

ACO rougher flotation tailings;

A mixture of the CIP tailings comprised of saprolite and ACO cleaner flotation tailings;

LCO whole ore material; and

A mixture of the CIP tailings comprised of LCO material and ACO cleaner flotation tailings.





In the event that variability gravity/float/and leach testing of the ACO or LCO material indicates an

alternative process flowsheet, then thickening tests should be expanded to reflect such a change.

This is especially relevant if an LCO cleaner flotation process is considered to better define the

impact of finer particles due to the regrind circuit which will impact on settling behavior.

If sufficient sample is produced, thickening and filtration testing of the copper concentrate should beperformed as selection and sizing of concentrate dewatering equipment is currently assumed due to

the absence of any related testwork.

Additional CND testing should be performed to confirm earlier findings. In addition, should any of the

variability testing of the ACO or LCO material yield a flowsheet significantly different than what is

currently proposed then CND testing should be expanded to included such changes.

Along with the recommend testwork necessary for the feasibility study, Sandspring may consider

investigating alternative processes methods such as SART (Sulfidization Acidification Recycle

Thickening).

It is estimated that the costs of this testwork is US$225,000. This excludes the costs of engineering

oversight and the costs of procurement of the samples themselves.

On-site Infrastructure

Tt recommends the following work to progress the Project to the next level:

Perform condemnation drilling and testing at the proposed plant site and man camp site;

Perform a geotechnical site investigation that includes test pits, drillholes and seismic

refraction surveys to support foundation design and engineering;

Perform a feasibility study design on the on-site infrastructure;

Confirm water supply source, quality and quantity requirements; and

Confirm wastewater treatment, discharge point, quality and quantity requirements.

Geotechnical Site Investigation

Tt recommends a geotechnical field investigation at the Toroparu Gold Project to supplement

previous investigation testing programs conducted at the site and to provide data to support a

feasibility study. The investigation program should include drilling of eight (8) boreholes and

excavating six test pits within the footprints of the proposed diversion structures, saddle dikes, man

camp and process facilities. The drilling and sampling program is aimed at characterizing the

subsurface conditions at the site as part of the feasibility design of the facilities.

Boreholes

The drilling program consists of a total of eight boreholes drilled to depths up to 50 m below ground

surface. Drilling can be stopped 3 m past the surface of competent bedrock if less than 50 m.

Surface drilling to the rock should be conducted using auger techniques that allows the ability of

performing Standard Penetration Testing, and collect disturbed and undisturbed samples. Sonic

drilling techniques are likely to fracture and/or crush coarse grained formations due to the vibratory

action at the drill bit and are therefore not acceptable for obtaining representative subsurface

samples for geotechnical characterization.





Test Pits

The test pit program should consist of a series of test pits to be trenched within the proposed

excavation areas for the process facilities, man camp, diversion structures and saddle dams. Test

pits depths should range 3 to 6 m below ground surface based on the ability of the available

equipment.

Sampling Requirements

The following sampling techniques are recommended for the Project:

SPT Split Barrel Sampling (ASTM D1586);

Undisturbed Samples with Shelby Tubes; and

Disturbed Samples in 5 gallon buckets.

Laboratory and Field Tests

Tt recommends the following laboratory and field tests. All testing should be performed according to

the appropriate ASTM Standard.

Modified proctor test;

Full gradation;

-200 sieve wash;

Atterberg limits;

Moisture content/in-place density;

Unconfined compressive strength of rock;

Swell/consolidation test;

Direct shear test (3 point); and

pH, chlorides and sulfates Test.

The estimated cost for performing the above mentioned work ranges between US$75,000 toUS$100,000.


Recommended fieldwork and testwork for feasibility design of the Tailings Management Area

includes the following:

Geotechnical field investigation program including:

o Local mapping;

o Drilling, coring and sampling of foundation materials;

o In situ testing and piezometer installation;

o

Test pit program with excavator for review of shallow foundation soils and sampling;o Site wide mapping program to search for filter materials, and

o Laboratory testing.

Estimated costs (including a 25% contingency):

o Drilling contractor costs: US$ 200,000 to US$ 250,000);

o Field supervision and travel fees: US$ 215,000;

o Laboratory fess: US$ 15,000; and

o Costs for construction of accesses not included and to be assumed by Sandspring.






Logistics Phase I – US$30,000:

o Verification of road maintenance agreement with GoG;

o Further definition of fuel supply logistics. Possible Trade-off Study of turnkey supply to

mine vs. SSP transport from port bulk storage; ando Verify availability and cost of port aggregate surfacing.

Pre-Design Engineering Phase I – US$85,000

o Further proving of aggregate surfacing sources;

o Further definition of tidal fluctuations;

o Bathymetric survey of Pine Tree Landing area; and

o Bridge and culvert hydrologic review.

Site Environmental and Project Closure

The geochemistry program should be advanced to a more detailed program that will include

predictions of water quality associated with the mining wastes run-off and discharges. Water quality

management strategies are needed for the tailings pond. Further static and kinetic testing isrecommended.

It is recommended that monitoring at the weather station be re-established, and that the data

collected include evaporation information as well as precipitation data recorded at intervals less than

1 hour to understand the severity of storms. Long-term and detailed metrological data are needed as

input for the design of the tailings impoundment facility and surface water control facilities.

The water quality baseline sampling has not included specific sampling events to establish a

baseline characterization trend with seasonal variability. SRK recommends that quarterly sampling

be re-established to coincide with the variation in the wet and dry seasons. The groundwater quality

results exhibited abnormally high concentrations of total suspended solids. In SRK’s experiencethese levels can be due to improper monitor well construction, development or sampling techniques.

SRK recommends that the sampling methodology and water construction and development

procedures be further reviewed to see if the well filter pack is appropriate, the well development was

adequate and the sampling technique is acceptable to international standards.

Consultation with the community should be continued. The Social Management Plan proposed in the

ESIA should be prepared and aspects implemented.

A conceptual closure plan should be developed based on the results of the additional information

generated during the Phase I.

The estimated cost for performing the environmental and closure tasks mentioned above ranges

between US$150,000 to US$250,000.





26.1.2 Phase II

Phase II work will be completing a feasibility study (having all fieldwork and testwork available from

Phase I).

Drilling/Geology/Resources:

The mineral resource model developed for the prefeasibility study will be quite adequate for a

feasibility study. No additional resource estimation work will be required, unless further definition of

the saprolite resource is requested in order to optimize the Project front-end saprolite processing

rate.

Open Pit Geotechnical/Hydrogeological Programs

Recommendations for future pit geotechnical/hydrogeological studies to complete a feasibility study

are below:

Incorporate the exploration drillhole data into the existing geotechnical database to refine the

rock mass model;

Update the geological and structural models; and

Conduct additional slope stability analyses (including numerical stress modeling).

The costs for the above pit geotechnical/hydrogeological feasibility level pit slope design and

dewatering plan are estimated to be US$80,000.

Mining

Additional mining studies required at the feasibility level of design for the Project include:

Detailed schedule for pre-production earthworks;

Incorporate additional geotechnical design data into the SE Pit design;

Feasibility level pit designs including dewatering structures;

Improved estimates of groundwater in-flow from local structures into the pit;

Assessment of a condemnation drilling program to confirm the locations of the low grade ore

stockpile, primary crusher and waste dumps;

Develop a feasibility level mine production schedule including monthly periods to start, and

completing the LoM schedule in quarterly periods to determine continuous ore exposure;

Assessment of an expanded articulated dump truck (ADT) fleet to mine part of the saprolite

waste and ore throughout most of the LoM mine production schedule;

Continued discussion with vendors for equipment quotes and detailed fuel usage;

Assessment of the adoption of a Maintenance and Repair Contract (MARC) as being an

optimum approach for mining equipment maintenance for the Toroparu Project;

Development of operational guidelines for treatment of any ARD waste rock; and

Further assessment of the low grade ore stockpiling approach with respect to water run-off

water quality management and oxidation of the ore. Given the competent nature of the fresh

rock (average UCS 150 MPa) it may be expected that oxidation will not penetrate RoM low

grade ore significantly.

The estimated cost for mining and mining related studies for a feasibility study are estimated to be

approximately US$250,000.





Saprolite Stockpile

For feasibility level the saprolite stockpile and ore re-handling will need to be defined. The stockpile

will need to be designed, and determination made whether re-handling will be by truck hauling or

pumping. The estimated cost for this portion of work is US$60,000.

Metallurgy and Processing

After the metallurgical results from the recommended testwork program are analyzed, the design

criteria will be reassessed and confirmed or adjusted. Additional trade-off studies should be

investigated and performed. Tt recommends Sandspring Resources perform a feasibility level study

on this criteria, process design, and mass balance revisions. As part of this study, Sandspring should

consider the incorporation of silver within the resource as a payable metal. This would be part of an

overall feasibility study. The estimated cost for this portion of the work ranges between US$500,000

to US$750,000,


Tt recommends Sandspring Resources Ltd. perform a feasibility study design on the on-siteinfrastructure. This would be part of an overall feasibility study effort. The estimated cost for this

portion of the work is US$250,000 to US$350,000.

Water Supply

Tt recommends Sandspring Resources Ltd. confirm the water supply source, quality and quantity

requirements. The estimated cost for this portion of the work is US$30,000.

Wastewater Treatment

Tt recommends Sandspring Resources Ltd. confirm the wastewater treatment, discharge point,

quality and quantity requirements. The estimated cost for this portion of the work is US$30,000.


Recommended work for feasibility design of the Tailings Management Area includes the following:

TMA Feasibility Design including the following main tasks:

o Site characterization: field program results compilation, design basis update and

hydrology parameters review;

o Feasibility level tailings characterization;

o Tailings facility layout, staging optimization, deposition modeling and analysis;

o Water management and water balance;

o Construction schedule and quantities; and

o Estimated cost (including a 25% contingency): US$300,000.


Logistics Phase II = US$7,000

o Availability and costs of desired barge service;

o Further verification of the navigability of the Cuyuni River for desired barge traffic;

Pre-Design Engineering Phase II – US$20,000

o Detailed boundary and topographical survey of Pine Tree Landing and road section from

Pine Tree Landing to Toll House. US$300,000.





Feasibility Study Project Management and Other Disciplines

Additional feasibility study Project management and other discipline studies required at the feasibility

level of design for the Project include:

Feasibility study Project management;

Review and risk analysis;

Economic modeling;

Other as required; and

Report compilation.

The estimated costs for these items for a feasibility study are estimated to range between

US$250,000 and US$300,000.

26.1.3 Summary of Recommended Work Program Costs

A summary of the estimated costs for the Phase I Programs and Phase II Studies to complete a

feasibility study are shown in Table 26.1.3.1.

Table 26.1.3.1: Cost Summary for Recommended Work for FS Completion

Recommended Work ProgramsBase Estimate

(US$)Upper Limit Estimate

(US$)Phase I Programs – Fieldwork and TestworkPit Geotechnical/Hydrogeological Program 270,000 270,000Condemnation Drilling Program 1,330,000 1,330,000Metallurgy and Processing Testwork 225,000 225,000On-Site Infrastructure Geotechnical Site Investigation 75,000 100,000Tailings Management Area Program 430,000 480,000Off-Site Infrastructure Program 115,000 115,000Site Environmental Program 150,000 250,000Phase I Totals $2,595,000 $2,770,000

Phase II Studies – Feasibility Study Completion Pit Geotechnical/Hydrogeological Studies 80,000 80,000Mining Studies 250,000 250,000Saprolite Stockpile Design 60,000 60,000Metallurgy and Processing Studies 500,000 750,000On-Site Infrastructure Studies 310,000 410,000Tailings Management Area Studies 300,000 300,000Off-Site Infrastructure Studies 27,000 27,000Site Environmental Studies/Report 50,000 50,000Feasibility Study Project Mgmt and Other Disciplines 250,000 300,000Phase II Totals $1,827,000 $2,227,000

Phase I and Phase II Totals $4,422,000 $4,997,000

Source: SRK based on other consultants’ submissions.Costs excluded existing camp operating costs.Costs excluded for construction of accesses.

The estimated costs for completing Phase I range between US$2,595,000 and US$2,770,000. The

estimated costs for completing Phase II range between US$1,777,000 and US$2,177,000.





27 References American Water Works Association, (1984). Design and Construction of Small Water Systems: A

Guide for Managers,

British Columbia Ministry of Environment (2012). Approved and Working Water Quality Guidelinesfor the Protection of Freshwater Aquatic Life. British Columbia Ministry of Environment,

Environmental Protection Division (http://www.env.gov.bc.ca/wat/wq/wq_guidelines.html).

Caterpillar Inc., (2012). Caterpillar Performance Handbook Edition 42, Caterpillar Publications,

Peoria, Illinois, 27 Sections.

Canadian Council of Ministers of the Environment (2012). Water Quality Guidelines for the Protection

of Aquatic Life (http://ceqg-rcqe.ccme.ca/).

Dardenne M.A. and Schobbenhaus C., (2003). Metallogeny of the Guiana Shield. Géologie de

France 2-3-4, p. 291-319.

Delor C., Roever E.W.F., Lafon J.M., Lahondère D., Rossi P., Cocherie A., Guerrot C. and Potrel A., (2003). The Bakhuis ultrahigh-temperature granulite belt (Surinam): Implications fro late

Transamazonian crustal stretching in a revisited Guiana Shield framework (2è Ed.):

Géologie de France, vol. 2-4, p. 207-230.

Dubé B. and Gosselin P., (2007). Greenstone-Hosted Quartz-Carbonate Vein Deposits. In

Goodfellow, W.D., ed., Mineral Deposits of Canada : A Synthesis of Major Deposit-Types,

District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods:

Geological Association of Canada, Mineral Deposits Division, Special Publication No. 5, p.

49-73

ETK Inc. (2012). Final environmental and social impact assessment, Toroparu Gold and Copper

Mine, January, 2012, 335 p.

Fraga L.M., Reis N.J. and Agnol R.D., (2009). Cauarane – Coeroeni Belt - The Main Tectonic

Feature of the Central Guyana Shield, northern Amazonian Craton. International Geological

Congres , 33. Oslo, Norway. Abstract.

Gibbs A.K. and Barron C.N., (1993). The Geology of the Guiana Shield. Oxford University Press,

New York., Claredon Press. 245 p.

Guyana Environmental Protection Agency (2012). Environmental Permit No. 20050201-ETKIO.

Issued by Guyana Environmental Protection Agency for ETK Inc., June 2012, 13 p.

Hirdes W., Davis D.W., Ludtke G. and Konan G., (1996). Two generations of Birimian

(Paleoproterozoic) volcanic belts in northeastern Côte d’Ivoire (West Africa): consequences

for the “Birimian controversy”. Precambrian Research 80, p. 173-191.

Hollister V.F., (1992). On a proposed plutonic porphyry-gold deposit model. Nonrenewable

Resources 1 (9), p. 293-302.

IUCN (2010). International Union for Conservation of Nature. Web. 01 May 2013. <http://iucn.org/>.

InfoMine USA, Inc., (2012). Mine and Mill Equipment Costs, Spokane Valley, Washington.





InfoMine USA Inc., (2013). Mining Cost Service, Spokane Valley, Washington.Klein E.L., Moura

C.A.V., Krymsky R.S. and Griffin W.L., 2005. The Gurupi Belt, northern Brazil :

Lithostratigraphy, geochronology, and geodynamic evolution. Precambrian Res. 141, p. 83-

105

Inspectorate. (2013) Prefeasibility Metallurgical Testing to Recover Gold on Samples fromSandspring’s Toroparu Project, Guyana. Prepared for Sandspring Resources Ltd. Report

dated February 18, 2013.

International Finance Corporation World Bank Group (2007). Environmental, Health and Safety

Guidelines for Mining, December 10, 2007, p. 26.

International Network for Acid Prevention (2012). The Global Acid Rock Drainage (GARD) Guide.

(http://www.gardguide.com/index.php).

Jacobs Minerals Canada Inc. (2011). SAG Versus HPGR Trade-off Study Toroparu Project.

Prepared for Sandspring Resources Ltd. Report dated November 18, 2011.

Jacobs Minerals Canada Inc. (2012). 15KTPD SAG Versus HPGR Trade-Off Study – ToroparuProject; Project 1878 – DRAFT. Prepared for Sandspring Resources Ltd. Report dated

January 13, 2012.

Jacobs Minerals Canada Inc. (2013). SAG Versus HPGR Trade-off Study – Toroparu Project;

CD1878.00. Prepared for Sandspring Resources Ltd. Report dated May 1, 2013.

Jacobs Minerals Canada Inc. (2013). Trade Off Study: SAG vs HPGR – 15k with future expansion;

Project 1878 – Rev A. Prepared for Sandspring Resources Ltd. Report dated January 31,

2013.

Klohn Crippen Berger (2011). Draft Tailings Storage Facility Alternatives Selection Assessment, draft

report prepared for Sandspring Resources Limited, December 5, 2011, 269 p.

Klohn, Krippen, Berger, (2012). Toroparu Geotechnical site Characterization-DRAFT. March 8.

Klohn Crippen Berger (2012a). Toroparu Project, Baseline ARD/ML Characterization – Waste Rock

and Low Grade Ore – Static Tests: unpublished report prepared for Sandspring Resources

Limited, May 5, 2012, 78 p.

Klohn Crippen Berger (2012b). Draft Toroparu Project, Acid Rock Drainage and Metal Leaching

(ARD/ML) Characterization Program, Waste Rock Short-term Leach Test Results:

unpublished draft report prepared for Sandspring Resources Limited, December 21, 2012,

31 p.

Klohn Crippen Berger Ltd (2012c) – “Toroparu Tailings Site Selection Assessment”.

Klohn Crippen Berger Ltd (2012d) – “Toroparu Geotechnical Site Characterization-DRAFT”. March 8.

Klohn Crippen Berger (2013). Toroparu Project, Baseline Geochemical (ARD/ML) Characterization

for ARD/ML Prediction, Draft: unpublished draft report prepared for Sandspring Resources

Limited, May 8, 2013, 103 p.

Knight Piésold Ltd. (2013). Toroparu Project – Prefeasibility Pit Slope Design Report.





Lang J.R. and Baker T., 2001. Intrusion-related gold systems : the present level of understanding.

Mineralium Deposita 36, p. 477-489.

Liègeois J.P., Claessens L.W., Camara D. and Klerkx J., (1991). Short-lived Eburnian orogeny in

southern Mali. Geology, tectonics, U-Pb and Rb-Sr geochronology. Precambrian Research

50, p. 111-136.

Lompo M., (2009). Geodynamic evolution of the 2.25-2.0 Ga Paleoproterozoic magmatic rocks in

the Man-Leo Shield of the West African Craton. A model of subsidence of an oceanic

plateau. In Palaeoproterozoic Supercontinents and Global Evolution. Geological Society,

London, Special Publications, 323, p. 231-254.

McCuaig T.C., Behn M., Stein H.J., Hagemann S.G., McNaughton N., Cassidy K.F., Champion D.

and Wyborn L., (2001). The Boddington gold mine : a new style of Archean Au-Cu deposit.

AGSO – Geoscience Australia 2001, p. 453-455.

Meixner (2008). Technical Report on the Toroparu Copper-Gold Prospect, Upper Puruni River Area,

Guyana”, dated July 10, 2008, revised August 8, 2008 with a second revision dated October

15, 2008.

Milesi J.P., Ledru P., Feybess J.L., Dommanget A. and Marcoux E., (1992). Early Proterozoic ore

deposits and tectonics of the Birimian orogenic belt, West Africa. Precambrian Research 58,

p. 305-344

Mine Development Associates, (2007). NI 43-101 Technical Report, Update Feasability Study, Las

Christinas Project, Venezuela (Crystallex International Corporation).

Mine Environment Neutral Drainage (2009) Prediction Manual for Drainage Chemistry from Sulphidic

Geologic Materials MEND Report 1.20.1. NRCan.

Morgan Building Systems. 2013. Budget Proposal for Man Camp. March 7.

P&E Mining Consultants, January 06, (2009); Technical Report, Resource Estimate on the Toroparu

Gold-Copper Deposit, Upper Puruni River Area, Guyana; NI 43-101 technical report No 153

prepared by P&E Mining Consultants Inc, for Sandspring Resources Ltd., Effective date of

October 26, 2008, 105 pages.

P&E Mining Consultants, (2010). Technical Report, Updated Resource Estimate on the Toroparu

Gold-Copper Deposit, Upper Puruni River Area, Guyana; NI 43-101 technical report No 186

prepared by P&E Mining Consultants Inc, for Sandspring Resources Ltd., Effective date of

May 12, 2010, Report Date July 16, 2010, 129 pages.

P&E Mining Consultants, March 12, 2012; Technical Report, Updated Resource Estimate and

Preliminary Economic Assessment of the Toroparu Gold-Copper Deposit, Upper PuruniProperty, Upper Puruni River Area, Guyana; NI 43-101 Technical Report No 234 prepared

by P&E Mining Consultants Inc, for Sandspring Resources Ltd., Effective Date of January

30, 2012, 215 pages.

P&E Mining Consultants, May 05, 2011; Technical Report, Updated Resource Estimate and

Preliminary Economic Assessment of the Toroparu Gold-Copper Deposit, Upper Puruni

Property, Upper Puruni River Area, Guyana; NI 43-101 Technical Report No. 208 prepared





by P&E Mining Consultants Inc., for Sandspring Resources Ltd., Effective Date of April 30,

2011, 213 pages.

P&E Mining Consultants, October 13, 2010; Technical Report, Updated Resource Estimate on the

Toroparu Gold-Copper Deposit, Upper Puruni River Area, Guyana; NI 43-101 Technical

Report No 193 prepared by P&E Mining Consultants Inc, for Sandspring Resources Ltd.,Effective date of September 12, 2010, 145 pages.

Pincock, Allen & Holt, 2008. NI 43-101 Technical Report, Update Feasability Study, Las Brisas

Project, Venezuela (Gold Reserve Inc.).

Planning for an Individual Water System. American Association for Vocational Instructional Materials,

1982

Ray, Brian, March 21, 2013;, Toroparu Property Check Assay Report, Sandspring Resources

internal Memorandum, 24 p.

Resource Development Inc. (2011) Gravity Test Results. Prepared for Sandspring Resources Ltd.

Report dated May 24, 2011.

Rowley D.B. and Pindell J.L., 1989. End Paleozoic-Early Mesozoic Western Pangean

Reconstruction and its Implications for the Distribution of Precambrian and Paleozoic Rocks

around Meso-America. Precambrian Research 42, p. 411-444.

SGS Canada Inc. (2013) A Prefeasibility Study of Gold Ore Containing Low and Average Copper

Grade from the Toroparu Deposit; Project 12520-002 – Final Report. Prepared for

Sandspring Resources Ltd. Report dated January 9, 2013.

SGS Canada Inc. (2011) A Deportment Study of Gold Deportment Study for a Sample from the

Toroparu Cu-Au Porphyry Property; Project 125020-001 – Final Report. Prepared for

Sandspring Resources Ltd. Report dated May 18, 2011.

SGS Canada Inc. (2012) A Metallurgical Evaluation of the Master Composite and Variability Samples

from the Toroparu Deposit; Project 12520-001 – Report 2. Prepared for Sandspring

Resources Ltd. Report dated November 19, 2012.

SGS Canada Inc. (2012) A Preliminary Metallurgical Evaluation of a Master Composite from the

Toroparu Deposit; Project 12520-001 – Report 1. Prepared for Sandspring Resources Ltd.

Report dated October 3, 2012.

SGS Lakefield Research Limited. (2009) An Investigation Into the Recovery of Gold and Copper

From the Toroparu Deposit of Sandspring Resources (Guyana); Project 12039-001 – Final

Report. Prepared for ETK. Report dated June 22, 2009.

Sidder G.B., 1995. Mineral Deposits of the Venezuelan Guayana Shield. U.S. Geological Survey

Bulletin 2124, 20 p.

Taylor G.H., Coste B., Lambert A., and Zeegers H., 1989. Geochemical signature (bedrock and

Saprolite) of gold mineralization and associated hydrothermal alteration at Dorlin, French

Guyana. Journal of Geochemical Exploration 32, p. 59-60

Tetra Tech, 2013. NI 43-101 Technical Report, Update Feasibility Study, Aurora Gold Project,

Guyana (Guyana Goldfields Ltd.)





Voicu G., Bardoux M., Jébrak, M. and Crépeau R., 1999. Structural, Mineralogical and

Geochemical Studies of the Paleaproterozoic Omai Gold Deposit, Guyana. Economic

Geology 94, p. 1277-1304

Voicu G., Bardoux M. and Stevenson R., 2001. Lithostratigraphy, geochrono-logy and gold

metallogeny in the northern Guiana Shield, South America: a review. Ore Geology Reviews18, p. 211-236.



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project Appendices


Appendices



SRK Consulting (U.S.), Inc.NI 43-101 Technical Report – Toroparu Gold Project Appendices


Appendix A: Cert ificates of Authors



SRK Consulting (US) Inc.

Suite 240, 3275 West Ina Road

Tucson, AZ 85741

T: 520 544 3688

F: 520 544 9853

[email protected]

www.srk.com

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 Office:

Guadalupe,

Zacatecas

52.492.927.8982

Canadian Offi ces:

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 AUTHOR

I, Al lan V. Moran, a Registered Geologist and a Certified Professional Geologist, do herebycertify that:

1. I am currently employed as a consulting geologist to the mining and mineral explorationindustry, as Principal Geologist with SRK Consulting (U.S.) Inc, with an office address of3275 W. Ina Rd., Tucson, Arizona, USA, 85741.

2. I graduated with a Bachelors of Science Degree in Geological Engineering from theColorado School of Mines, Golden, Colorado, USA; May 1970.

3. I am a Certified Professional Geologist through membership in the American Institute of

Professional Geologists, CPG - 09565, and have been since 1995; and I am a RegisteredGeologist in the State of Oregon, USA, # G-313, and have been since 1978.

4. I have been employed as a geologist in the mining and mineral exploration business,continuously, for the past 42 years.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association(as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to bea “qualified person” for the purposes of NI 43-101. The Technical Report is based upon mypersonal review of the information provided by the issuer. My relevant experience for thepurpose of the Technical Report is:

Manager, Exploration North America for Cameco Gold Inc., 1998-2002:

Vice President and U.S. Exploration Manager for Independence Mining Company,Reno, Nevada, 1990-1993;

Exploration Geologist for Freeport McMoRan Gold, 1980-1990;

Gold exploration experience, as an exploration geologist, from 1980 through 2002; andwork with gold deposits since 2002 as a consultant with SRK Consulting (U.S.) Inc;

Experience in the above positions working with and reviewing resource estimationmethods and models, in concert with resource estimation geologists and engineers.

As a consultant, from 2003 to 2013, I have completed several NI 43-101 Technicalreports.

6. I am responsible for all of Sections 6 through 12, and for inputs to Sections 1, 25, and 26 ofthe Technical Report titled “NI 43-101 Technical Report, Prefeasibility Study, Toroparu GoldProject, Upper Puruni River Area, Guyana.”, with an effective date of May 08, 2013 (the



SRK Consulting Page 2

“Technical Report”) relating to Sandspring Resources Ltd.’s interests in the Toroparu GoldProject . I have personally visited the Project on April 18 and 19, 2012, for 2 days.

7. I have not had prior involvement with the property that is the subject of this TechnicalReport.

8. As of May 08, 2013, the effective date of this report, to the best of my knowledge,

information and belief, the Technical Report contains all the scientific and technicalinformation that is required to be disclosed to make the Technical Report not misleading.

9. I am not aware of any material fact or material change with respect to the subject matter ofthe Technical Report that is not reflected in the Technical Report, for which the omission todisclose would make the Technical Report misleading.

10. I am independent of the issuer applying all of the tests in Item 1.5 of National Instrument43-101.

11. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report hasbeen prepared in compliance with that instrument and form.

12. I consent to the filing of the Technical Report with any stock exchange and other regulatoryauthority and any publication by them, including electronic publication in the publiccompany files on their websites accessible to the public, of the Technical Report.

Dated in Tucson, Arizona, May 17, 2013.

Allan V. Moran, R. Geol., CPG

AIPG CPG – 09565



Tetra Tech, Inc.

350 Indiana Street, Suite 500, Golden, CO 80401

Tel 303-217-5700 Fax 303-217-5705 tetratech.com

C E R T I F I C A T E O F Q U A L I F I E D P E R S O N

D. Erik Spiller

As co-author of this report entitled “NI 43-101 Technical Report, Toroparu Gold Project, Upper Puruni River Area,Guyana”, with an effective date of 8 May, 2013 (the “Technical Report”), I, D. Erik Spiller, do hereby certify that:

1. I am employed as a Vice-President - Principal Metallurgist by, and carried out this assignment for: Tetra Tech,Inc., 350 Indiana Street, Suite 500, Golden, Colorado 80401, USA tel. (303) 217-5700 email:[email protected].

2. I hold the following academic qualifications: B.Sc.in Metallurgical Engineering, Colorado School of Mines, 1970.

3. I am a Qualified Professional (QP) member of the Mining and Metallurgical Society of America (MMSA#01021QP). In addition, I am a Registered (QP) member of Society for Mining, Metallurgy, and Exploration, Inc.(SME #3051820RM).

4. I have more than 40 years of relevant experience working as a metallurgical engineer in the mineral resourceindustry. During this career I held responsible positions in process research, process development, engineering,and senior management. In addition, since 2008 I have served as an appointed Research Professor at theColorado School of Mines; this was preceded by 15 years as an Adjunct Instructor at the same institution. I

lecture in mineral beneficiation and direct graduate students conducting metallurgical research in my area ofexpertise.

5. By reason of education, experience and qualified professional registration, I fulfill the requirements of aQualified Person as defined in NI 43-101.

6. I did not personally visit the Toroparu Gold Project site. I did visit SGS Canada to inspect the on-goingmetallurgical test work. My most recent personal inspection of the laboratory was November 1, 2012 for one day.

7. I am responsible for the preparation of Sections 13, 17, 18.1.8, 21.2.2, 21.2.3, and the portions of Sections 1,25 and 26 summarized therefrom, of the Technical Report.

8. I am independent of Sandspring Resources Ltd., as defined in Section 1.5 of NI 43-101.

9. I have had no previous involvement with the property.

10. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared incompliance with the instrument.

11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of thisTechnical Report for which I am responsible contain all scientific and technical information that is required to bedisclosed to make this report not misleading.

Dated this 24th day of May 2013

SIGNED AND SEALED

Signature of Qualified Person

D. Erik Spiller

Print Name of Qualified Person

mailto:[email protected]




Tetra Tech, Inc.

3801 Automation Way, Suite 100, Fort Collins, CO 80525Tel 970-223-9600 Fax 970-223-7171 tetratech.com

C E R T I F I C A T E O F Q U A L I F I E D P E R S O N

Daniel Lloyd Evans

As co-author of this report entitled “NI 43-101 Technical Report, Toroparu Gold Project, Upper Puruni River Area,Guyana”, with an effective date of 8 May, 2013 (the “Technical Report”), I, Daniel Lloyd Evans, do hereby certifythat:

1. I am employed as a Senior Water Resources Engineer by, and carried out this assignment for: Tetra Tech, Inc.,3801 Automation Way, Suite 100, Fort Collins, CO, 80525 tel. (970) 223-9600 email: [email protected] .

2. I hold the following academic qualifications: B.A. Geophysics, Occidental College, 1988; M.S. Civil Engineering – emphasis in Hydrology, Colorado State University, 1991.

3. I am a Professional Engineer registered with the State of Colorado, USA. (Registration number 32081).

4. I have practiced my profession continuously since 1991. Since 1994, I have engineered infrastructure designsfor a variety of early and advanced base mineral projects in the Yukon, Saskatchewan, Alaska, Australia, State ofWashington, and the State of Nevada.

I have worked in the minerals industry for 11 years; my work experience includes 11 years as a water resources

engineer working in the field of surface water hydrology and hydraulics.

5. I do, by reason of education, experience and professional registration, fulfill the requirements of a QualifiedPerson as defined in NI 43-101.

6. My most recent personal inspection of the Property was October 19 – 21, 2012 for three days.

7. I am responsible for water management Sections 16.9.1 through 16.9.3, 18.1.2, 18.1.9, 18.1.10 and portions ofSections 1, 25 and 26 summarized therefrom, of the Technical Report.

8. I am independent of Sandspring Resources Ltd., as defined in Section 1.5 of NI 43-101.

9. I have had no previous involvement with the property.

10. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared incompliance with the instrument.

11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of thisTechnical Report for which I am responsible contain all scientific and technical information that is required to bedisclosed to make this report not misleading.

Dated this 24th day of May 2013

Signature of Qualified Person

Daniel Lloyd Evans

Print Name of Qualified Person







SRK Tucson

3275 W. Ina Road

Suite 240

Tucson, AZ 85741

T: 520 544 3688

F: 520 544 9853

[email protected]

www.srk.com

U.S. Offices:


Denver 303.985.1333

Elko 775.753.4151


Reno 775.828.6800

Tucson 520.544.3688

Mexico Offices:

Zacatecas 52.492.927.8982

Querétaro 52.442.218.1030

Canadian Offi ces:


Sudbury 705.682.3270

Toronto 416.601.1445



Group Offices:

Africa

Asia

Australia

Europe

North America

South America


I, Dawn H. Garcia, P.G., CPG, do hereby certify that:

1. I am a Principal Hydrogeologist of SRK Consulting (U.S.), Inc., 3275 W Ina Road, Suite 240, Tucson Arizona 85741.

2. This certificate applies to the technical report titled “NI 43-101 Technical Report, Prefeasibility Study,Toroparu Gold Project, Upper Puruni River Area, Guyana” with an Effective Date of May 8, 2013 (the“Technical Report”).

3. I graduated with a degree in Geological Sciences from Bradley University in 1982. In addition, I haveobtained a M.S., Geology, 1995, California State University Long Beach. I am a Certified ProfessionalGeologist of the American Institute of Professional Geologists (CPG-08313) and a Registered Member of

the Society of Mining Metallurgy, and Exploration, Inc. (RM-4135993).

4. I have worked as a geologist/hydrogeologist for a total of 28 years since my graduation from university.My relevant experience includes environmental compliance permitting, hydrogeological studies andgeotechnical studies at mining and processing operations.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43-101) andcertify 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 ofNI 43-101.

6. I have not visited the property and thus I have relied on the descriptions of others.

7. I am responsible for the preparation of Sections 20 (excluding geochemistry portions of 20.1 and TMAportion of 20.7.1), and portions of 1, 25 and 26 summarized therefrom, of this Technical Report.

8. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101.9. I have not had prior involvement with the property that is the subject of the Technical Report.

10. I have read NI 43-101 and Form 43-101-F1 and the sections of the Technical Report I am responsible forhave been prepared in compliance with that instrument and form.

11. As of 8 May 2013, to the best of my knowledge, information and belief, the sections of the TechnicalReport I am responsible for contains all scientific and technical information that is required to bedisclosed to make the Technical Report not misleading.

Dated this 8th Day of May, 2013, at Tucson, Arizona, USA.

________________________________ “Sealed ”

Dawn H. Garcia



SRK Denver

Suite 3000

7175 West Jefferson Avenue

Lakewood, CO 80235

T: 303.985.1333

F: 303.985.9947

[email protected]

www.srk.com

U.S. Offi ces:


Denver 303.985.1333

Elko 775.753.4151


Reno 775.828.6800

Tucson 520.544.3688

Mexico Office:

Guadalupe, Zacatecas

52.492.927.8982

Canadian Offi ces:


Sudbury 705.682.3270

Toronto 416.601.1445



Group Offices:

Africa

Asia

Australia

Europe

North America

South America

QP_Cert_Rodrigues_Toroparu_2013


I, Fernando Rodrigues, B.S. Mining, MMSA do hereby certify that:

1. I am Senior Mining Engineer of:

SRK Consulting (U.S.), Inc.7175 W. Jefferson Ave, Suite 3000Denver, CO, USA, 80235

2. I graduated with a Bachelors of Science degree in Mining Engineering from South Dakota School ofMines and Technology in 1999.

3. I am a QP member of the MMSA.

4. I have worked as a Mining Engineer for a total of 13 years since my graduation from South DakotaSchool of Mines and Technology in 1999. My relevant experience includes contributions to multiple,feasibility, pre-feasibility, preliminary assessments, due diligence and competent person reports; designand implementation, short term mine design, dump design, haulage studies, blast design, ore control,grade estimation, and database management.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43-101) andcertify 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 ofNI 43-101.

6. I am responsible for the preparation of Sections 15, 16 (except for sub-section 16.2, 16.8 and 16.9), andportions of Sections 1, 25 and 26 summarized therefrom, of the technical report titled “NI 43-101Technical Report, Prefeasibility Study, Toroparu Gold Project, Upper Puruni River Area, Guyana”, withan effective date of May 08, 2013, and dated May 24, 2013 (the “Technical Report”). I have not visitedthe property.

7. I have not had prior involvement with the property that is the subject of the Technical Report.

8. I am independent of the issuer applying all of the tests in section 1.5 of National Instrument 43-101.

9. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliancewith that instrument and form.




QP_Cert_Rodrigues_Toroparu_2013

10. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority andany publication by them for regulatory purposes, including electronic publication in the public companyfiles on their websites accessible by the public, of the Technical Report.

11. As of May 8, 2013, to the best of my knowledge, information and belief, the portion(s) of the TechnicalReport 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 24th Day of May, 2013.

“Signed” “Sealed” ________________________________Fernando Rodrigues, B.S. M.Eng, MMSA MMSA # 01405QP



SRK Denver

Suite 3000


Lakewood, CO 80235

T: 303.985.1333

F: 303.985.9947

[email protected]

www.srk.com


I, Frank Daviess, MAusIMM, Registered SME do hereby certify that:

1. I am currently employed as a consulting resource geologist to the mining and mineral explorationindustry and I am currently under contract as 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 Technical Report, Prefeasibility Study,Toroparu Gold Project, Upper Puruni River Area, Guyana”” with an Effective Date of effective date ofMay 8, 2013 and dated May 24, 2013 (the “Technical Report”).

3. I graduated from the University Of Colorado, Boulder, Colorado, USA with a B.A. in Geology in 1971 anda M.A. in Natural Resource Economics and Statistics in 1975. I am a Member of the AustralasianInstitute of Mining and Metallurgy (Registration No. 226303). I am a Registered Member of the Societyfor Mining, Metallurgy and Exploration, Inc. (Registration No. 0742250). I have been employed as ageologist in the mining and mineral exploration business, continuously, for the past 38 years, since mygraduation from university. I am qualified as a competent person for the resource estimation of manycommodities under the JORC/CIM guidelines and also have experience with Sarbanes-Oxley (SOX)compliance, due diligence auditing and risk assessment.


5. I visited the property on November 10 through 12, 2010 for three days.

6. I am responsible for the resource estimation Section 14 and portions of Sections 1, 25 and 26summarized therefrom, of this Technical Report.

7. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101. I have not had prior

involvement with the property that is the subject of the Technical Report.

8. I have read NI 43-101 and Form 43-101-F1 and the sections of the Technical Report I am responsible forhave 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, thesections of the Technical Report I am responsible for contains all scientific and technical information thatis required to be disclosed to make the Technical Report not misleading.


“signed” “sealed”

________________________________

Frank Daviess



SRK Denver

Suite 3000


Lakewood, CO 80235

T: 303.985.1333

F: 303.985.9947

[email protected]

www.srk.com

U.S. Offi ces:


Denver 303.985.1333

Elko 775.753.4151


Reno 775.828.6800

Tucson 520.544.3688

Mexico Office:

Guadalupe, Zacatecas

52.492.927.8982

Canadian Offi ces:


Sudbury 705.682.3270

Toronto 416.601.1445



Group Offices:

Africa

Asia

Australia

Europe

North America

South America

QP_Cert_Clarke_Toroparu_2013


I, Peter Clarke, B.Sc., MBA, P. Eng. do hereby certify that:

1. I am a Principal Mining Engineer of:

SRK Consulting (U.S.), Inc.7175 W. Jefferson Ave, Suite 3000Denver, CO, USA, 80235

2. I graduated with a B.Sc. degree in Mining Engineering granted by the University of Leeds in 1975 and anMBA granted by the University of Phoenix in 2002.

3. I am a registered member in good standing of the Association of Professional Engineers andGeoscientists of British Columbia since 1982.

4. I have worked as a mining engineer for a total of 28 years since my graduation from university.experience as an open-pit mining engineer in mining operations and mine engineering consulting.Experience includes mining of precious metals, copper, lead, zinc, nickel, and industrial minerals in North

America and overseas. I have an extensive background in open-pit mine design, planning, productionscheduling, equipment selection and cost estimating. Studies conducted include property evaluations,scoping studies, feasibilities, mine planning optimizations, and due diligence.


6. I am responsible for project economics, mining equipment and costs and other information in Sections 2through 5, 16.8, 16.9.4, 19, 21 (except for 21.2.2, 21.2.3, 21.3.3 and 21.3.4), 22 through 24, and portionsof Sections 1, 25 and 26 summarized therefrom of the technical report titled “NI 43-101 TechnicalReport, Prefeasibility Study, Toroparu Gold Project, Upper Puruni River Area, Guyana”, with an effectivedate of May 8, 2013, and dated May 24, 2013 (the “Technical Report. I did not visit the property.

7. I have not had prior involvement with the property that is the subject of the Technical Report.

8. I am not independent of the issuer applying all of the tests in section 1.5 of National Instrument 43-101.

9. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliancewith that instrument and form.




10. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority andany publication by them for regulatory purposes, including electronic publication in the public companyfiles on their websites accessible by the public, of the Technical Report.

11. As of May 08, 2013, the effective date of this report, to the best of my knowledge, information and belief,the Technical Report contains all the scientific and technical information that is required to be disclosed

to make the Technical Report not misleading.


“Signed” “Sealed”

________________________________Peter Clarke, B.Sc., MBA, P. Eng. P.Eng Registration No.: 13473

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