Dalradian Resources Inc. a Preliminary Economic

Home , Altmore, Black Bog, Mullaghcarn, Rousky

D A L R A DI A N R ESO UR C ES IN C.

A PR E L I M IN A R Y E C O N O M I C ASSESSM E N T O F T H E C URR A G H IN A L T G O L D D EPOSI T, T Y R O N E PR OJE C T, N O R T H E RN IR E L A ND

SEPT E M B E R 6, 2012 E F F E C T I V E D A T E (R ESO UR C E): N O V E M B E R 30, 2011 E F F E C T I V E D A T E (PE A): JU L Y 25, 2012

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Table of Contents Page 1.0 SU M M A R Y ...... 1 1.1 SCOPE OF WORK ...... 1 1.2 LOCATION AND DESCRIPTION ...... 1 1.3 GEOLOGY AND MINERALIZATION ...... 3 1.4 HISTORY ...... 5 1.5 EXPLORATION ...... 6 1.6 MINERAL RESOURCES ...... 6 1.7 MINING ...... 8 1.8 METALLURGY AND PROCESSING ...... 13 1.8.1 Mineralogy and Metallurgical Testing ...... 13 1.8.2 Process Flowsheets Considered ...... 13 1.8.3 Process Description (Option A) ...... 15 1.8.4 Process Description (Option D) ...... 16 1.8.5 Process Manpower ...... 16 1.9 INFRASTRUCTURE ...... 17 1.9.1 Access and Site Roads ...... 17 1.9.2 Power Supply ...... 18 1.9.3 Water Supply ...... 18 1.9.4 Tailings Storage Facility ...... 18 1.10 ENVIRONMENTAL, PERMITTING AND SOCIAL ...... 18 1.10.1 Studies and Issues ...... 18 1.10.2 Waste and Water Management ...... 19 1.10.3 Social Management ...... 20 1.10.4 Permits ...... 20 1.10.5 Closure ...... 21 1.11 CAPITAL AND OPERATING COST ESTIMATES ...... 22 1.11.1 Capital Costs ...... 22 1.11.2 Operating Costs ...... 23 1.12 ECONOMIC ANALYSIS AND SENSITIVITY STUDIES ...... 23 1.12.1 Basis of Evaluation ...... 23 1.12.2 Macro-economic Assumptions ...... 24 1.12.3 Technical Assumptions ...... 26 1.12.4 Sensitivity Study ...... 32 1.13 CONCLUSIONS ...... 35 1.14 RECOMMENDATIONS ...... 36 2.0 IN T R O DU C T I O N ...... 39 3.0 R E L I A N C E O N O T H E R E XPE R TS ...... 41 4.0 PR OPE R T Y D ESC RIPT I O N A ND L O C A T I O N ...... 42

4.1 LOCATION ...... 42 4.2 DESCRIPTION ...... 43 4.3 LOCATION OF MINERALIZED ZONES ...... 46 4.4 PERMITS ...... 48 4.5 ENVIRONMENTAL CONSIDERATIONS ...... 48 5.0 A C C ESSIBI L I T Y, C L I M A T E, L O C A L R ESO UR C ES, IN FR AST RU C T UR E A ND PH YSI O G R APH Y ...... 49 6.0 H IST O R Y ...... 51 6.1 ACQUISITION HISTORY...... 51 6.2 EXPLORATION HISTORY ...... 52 6.3 HISTORICAL MINERAL RESOURCE ESTIMATES ...... 56 6.4 HISTORICAL PRODUCTION ...... 57 7.0 G E O L O G I C A L SE T T IN G A ND M IN E R A L I Z A T I O N ...... 58 7.1 REGIONAL GEOLOGY ...... 58 7.2 PROPERTY GEOLOGY ...... 59 7.2.1 Curraghinalt Gold Deposit Area - Licence DG1...... 59 7.2.2 Licence DG2 - Tyrone Volcanic Group ...... 67 7.2.3 DG3 and DG4 Licences ...... 70 7.3 MINERALIZATION ...... 74 7.3.1 The Curraghinalt deposit ...... 74 7.3.2 The Tyrone Volcanic Group ...... 78 8.0 D EPOSI T T YPES ...... 82 8.1 THE CURRAGHINALT DEPOSIT ...... 82 8.2 THE TYRONE IGNEOUS COMPLEX ...... 83 9.0 E XPL O R A T I O N ...... 85 9.1 NICKELODEON/NAVIGATOR/STRONGBOW 1997 - 2003 ...... 85 9.2 TOURNIGAN EXPLORATION, 2003 - JANUARY, 2005 ...... 86 9.3 TOURNIGAN EXPLORATION FROM JANUARY, 2005 ...... 87 9.4 TOURNIGAN EXPLORATION 2007 TO 2009 ...... 88 9.5 DALRADIAN RESOURCES EXPLORATION ...... 90 10.0 DRI L L IN G ...... 97 10.1 TOURNIGAN DRILL PROGRAMS ...... 97 10.2 DALRADIAN RESOURCES DRILLING ...... 99 10.3 CORE LOGGING ...... 106 11.0 SA MPL E PR EPA R A T I O N, A N A L YSES A ND SE C URI T Y ...... 109 11.1 SAMPLING ...... 109 11.2 SAMPLE ANALYSES ...... 110

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11.3 SAMPLE SECURITY ...... 111 11.4 QUALITY CONTROL AND QUALITY ASSURANCE ...... 111 11.4.1 Early QA/QC Programs ...... 111 11.4.2 Dalradian Resources QA/QC Program ...... 113 11.4.3 Blank Sample Results ...... 114 11.4.4 Reference Standard Sample Results ...... 115 11.4.5 Duplicate Sample Results ...... 127 11.4.6 Dalradian Resources QA/QC Since October, 2011 ...... 128 11.5 SPECIFIC GRAVITY DETERMINATIONS ...... 129 12.0 D A T A V E RI FI C A T I O N ...... 132 12.1 DALRADIAN RESOURCES DATA VERIFICATION ...... 132 12.2 MICON DATA VERIFICATION ...... 132 13.0 M IN E R A L PR O C ESSIN G A ND M E T A L L UR G I C A L T EST IN G ...... 134 13.1 SAMPLE CHARACTERISATION - MINERALOGY ...... 134 13.2 METALLURGICAL TESTWORK RESULTS ...... 135 13.2.1 Comminution ...... 135 13.2.2 Gravity Concentration ...... 135 13.2.3 Heavy Media Separation (HMS) ...... 135 13.2.4 Cyanidation ...... 136 13.2.5 Flotation ...... 136 13.3 PROPOSED FUTURE TESTWORK ...... 137 14.0 M IN E R A L R ESO UR C E EST I M A T ES ...... 138 14.1 MICON MINERAL RESOURCE ESTIMATE...... 138 14.2 GEOLOGICAL MODEL ...... 138 14.2.1 Basic Statistics ...... 140 14.2.2 Raw Assay Intervals ...... 146 14.2.3 Compositing ...... 153 14.3 GRADE INTERPOLATION METHOD ...... 153 14.3.1 Estimation Parameters and Search Distances ...... 153 14.4 BLOCK MODEL ...... 155 14.5 BLOCK MODEL VALIDATION ...... 156 14.5.1 Visual Inspection ...... 156 14.5.2 Comparison of individual block grades with de-clustered sample grade ...... 156 14.5.3 Alternate Estimation Method ...... 157 14.6 MINERAL RESOURCE CLASSIFICATION ...... 161 14.7 MINERAL RESOURCE ESTIMATE ...... 166 15.0 M IN E R A L R ESE R V E EST I M A T ES ...... 169 16.0 M ININ G M E T H O DS ...... 170 16.1 INTRODUCTION ...... 170 16.2 MINE DESIGN PARAMETERS ...... 170

16.2.1 Geotechnical Parameters ...... 171 16.3 CUT-OFF GRADE DETERMINATION ...... 172 16.4 MINERAL RESOURCES CONSIDERED IN THE MINE PLAN ...... 173 16.4.1 Mining Recovery and Dilution...... 173 16.4.2 Classification of Mineral Resources Considered in the Mine Plan...... 174 16.5 MINING METHOD ...... 176 16.6 UNDERGROUND MINE PLAN AND DEVELOPMENT ...... 177 16.7 MINE VENTILATION REQUIREMENT ...... 181 16.8 DRILLING AND BLASTING ...... 182 16.9 HAULAGE OF WASTE AND MINERALIZED MATERIAL ...... 183 16.10 BACKFILL SYSTEM - PASTEFILL ...... 184 16.11 MINE PRE-PRODUCTION AND PRODUCTION SCHEDULE ...... 184 16.12 DILUTION CONTROL ...... 187 16.13 EQUIPMENT FLEET ...... 188 16.13.1 Mine Services ...... 188 16.13.2 Underground Mobile Equipment ...... 188 16.14 MINING MANPOWER ...... 189 17.0 R E C O V E R Y M E T H O DS ...... 192 17.1 DESIGN BASIS ...... 194 17.2 PROCESS DESCRIPTION ...... 196 17.2.1 Crushing (Options A and D): ...... 196 17.2.2 Grinding (Options A and D): ...... 197 17.2.3 Copper Flotation (Option D): ...... 197 17.2.4 Copper Concentrate Dewatering (Option D) ...... 197 17.2.5 Bulk Sulphide Flotation (Option D) ...... 197 17.2.6 Pre-treatment and Cyanidation (Options A and D) ...... 198 17.2.7 Carbon Stripping/Carbon Reactivation/Gold Room (Options A and D) ...... 198 17.2.8 Cyanide Destruction (Options A and D) ...... 198 17.3 MANPOWER ...... 198 17.4 PROCESSING CONCLUSIONS ...... 199 18.0 PR OJE C T IN FR AST RU C T UR E ...... 201 18.1 SITE ACCESS ROAD ...... 201 18.2 SITE ROADS ...... 201 18.3 POWER SUPPLY ...... 201 18.4 WATER SUPPLY ...... 202 18.5 BUILDINGS ...... 202 18.6 ANCILLARY FACILITIES ...... 202 18.7 CONSTRUCTION FILL MATERIALS ...... 203 18.8 TAILINGS STORAGE FACILITY ...... 203

18.9 SCHEMATIC SITE LAYOUT ...... 203 19.0 M A R K E T ST UDI ES AND C O N T R A C TS ...... 205 20.0 E N V IR O N M E N T A L ST UDI ES, PE R M I T T IN G A ND SO C I A L O R C O M M UNI T Y I MPA C T ...... 206 20.1 ENVIRONMENTAL STUDIES AND ISSUES ...... 206 20.1.1 Introduction ...... 206 20.2 WASTE AND WATER MANAGEMENT ...... 209 20.2.1 Waste Characterization and Waste Management ...... 209 20.2.2 Water Management ...... 209 20.2.3 Social and Environmental Management ...... 210 20.3 PERMITTING REQUIREMENTS ...... 211 20.3.1 Permits for Exploration ...... 211 20.3.2 Permit Requirements for Development ...... 212 20.4 SOCIAL AND COMMUNITY ASPECTS, STAKEHOLDER CONSULTATION ...... 213 20.5 MINE CLOSURE REQUIREMENTS ...... 214 21.0 C API T A L A ND OPE R A T IN G C OSTS ...... 216 21.1 CAPITAL COSTS...... 216 21.1.1 Mine Development ...... 216 21.1.2 Mining Equipment ...... 217 21.1.3 Processing Plant ...... 217 21.1.4 Infrastructure ...... 217 21.1.5 ,QGLUHFW&DSLWDO2ZQHU¶V&RVWDQG&RQWLQJHQF\ ...... 218 21.2 OPERATING COSTS ...... 219 21.2.1 Mine Operating Costs ...... 219 21.2.2 Processing Operating Costs ...... 219 21.2.3 General and Administration Costs ...... 219 22.0 E C O N O M I C A N A L YSIS ...... 221 22.1 BASIS OF EVALUATION ...... 221 22.2 MACRO-ECONOMIC ASSUMPTIONS ...... 221 22.2.1 Exchange Rate and Inflation ...... 221 22.2.2 Weighted Average Cost of Capital ...... 221 22.2.3 Expected Metal Prices ...... 222 22.2.4 Taxation Regime ...... 223 22.2.5 Royalty ...... 224 22.2.6 Selling Expenses ...... 224 22.3 TECHNICAL ASSUMPTIONS ...... 224 22.3.1 Mine Production Schedule ...... 224 22.3.2 Operating Costs ...... 225 22.3.3 Capital Costs ...... 226 22.3.4 Base Case Cash Flow ...... 227 22.3.5 Base Case Evaluation ...... 230

22.4 SENSITIVITY STUDY ...... 230 22.4.1 Capital, Operating Costs and Revenue Sensitivity ...... 230 22.4.2 Metal Price Sensitivity ...... 231 22.4.3 Sensitivity to Process Flowsheet ...... 232 22.5 CONCLUSION ...... 233 23.0 A DJA C E N T PR OPE R T I ES ...... 234 24.0 O T H E R R E L E V A N T D A T A A ND IN F O R M A T I O N ...... 236 25.0 IN T E RPR E T A T I O N A ND C O N C L USI O NS ...... 237 25.1 MINERAL RESOURCES ...... 237 25.2 MINING ...... 239 25.3 MINERAL PROCESSING ...... 239 25.4 INFRASTRUCTURE AND CAPITAL COSTS ...... 240 26.0 R E C O M M E ND A T I O NS ...... 241 26.1 PROPOSED EXPLORATION AND DEVELOPMENT PROGRAM ...... 241 26.1.1 Curraghinalt Resource Drilling and Update ...... 241 26.1.2 Curraghinalt Development ...... 241 26.1.3 Budget ...... 243 27.0 R E F E R E N C ES ...... 246 28.0 C E R T I FI C A T ES ...... 249 29.0 APPE NDI X 1 ...... 256 30.0 APPE NDI X 2 ...... 263 31.0 APPE NDI X 3 ...... 285

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List of Tables Page

Table 1.1 Classified Mineral Resource Estimate at 5 g/t Gold Cut-off Grade and 0.1 m Minimum Width, Diluted to a Minimum Width of 1.0 m ...... 7 Table 1.2 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper ...... 7 Table 1.3 Measured, Indicated and Inferred Mineral Resources Considered in the Mine Plan ...... 9 Table 1.4 Total Underground Development for the Curraghinalt Project ...... 11 Table 1.5 Underground Mobile Equipment Fleet ...... 11 Table 1.6 Estimate Mine Labour Requirement ...... 12 Table 1.7 General and Administrative Manpower ...... 12 Table 1.8 Production Summary for the Dalradian Resources Tyrone Project Milling Plant ...... 14 Table 1.9 Summary of Estimated Processing Plant Operations Personnel ...... 17 Table 1.10 Key Applicable Environmental Legislation ...... 21 Table 1.11: Capital Cost Summary ...... 22 Table 1.12 Summary of Life-of-Mine Operating Costs ...... 23 Table 1.13 Estimated Cost of Equity ...... 24 Table 1.14 Metal Price Averages ...... 25 Table 1.15 Life-of-Mine Cash Flow Summary ...... 29 Table 1.16 Base Case Life of Mine Annual Cash Flow ...... 30 Table 1.17 Base Case Cash Flow Evaluation ...... 32 Table 1.18 Sensitivity to Metal Prices ...... 34 Table 1.19 Exploration and Development Program Summary Budget ...... 37 Table 4.1 Licence Details ...... 44 Table 4.2 Licence Expenditures ...... 45 Table 6.1 Summary of Polygonal Mineral Resource Estimate, 1997 ...... 57 Table 7.1 Correlation of Dalradian Rocks in Scotland and North of Ireland...... 62 Table 7.2 Definitions for Geometry of the Vein System ...... 65 Table 9.1 Selected Results from Trench T10, Glenlark, Sampled Widths ...... 85 Table 9.2 Summary of 2011 Prospecting Samples ...... 90 Table 9.3 Tyrone Volcanic Group Exploration Targets ...... 93 Table 9.4 Dalradian Supergroup Exploration Targets ...... 94 Table 10.1 Summary of Drilling by Year ...... 97 Table 10.2 Summary of Drilling 2005-2007 ...... 98 Table 10.3 Summary of Drilling 2008 ...... 98

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Table 10.4 Summary of Dalradian Resources Drilling From Hole CT-58 ...... 103 Table 10.5 Summary of Abandoned Holes Not Used in Resource Estimation...... 104 Table 10.6 Summary of Dalradian Drill Holes Completed Since the 2011 Resource Update ...... 105 Table 10.7 Drill Set-up and Core Logging Controls ...... 108 Table 11.1 Drill Set-up, Sampling Procedures and Controls ...... 110 Table 11.2 OMAC Laboratory Assay Procedures ...... 111 Table 11.3 'HWDLOVRI7RXUQLJDQ¶V4$4&6DPSOHV...... 112 Table 11.4 'HWDLOVRI'DOUDGLDQ5HVRXUFHV¶4$4&6DPSOHV...... 114 Table 11.5 Dalradian Resources Blank Samples Designated as Outliers ...... 115 Table 11.6 Certified Value of CDN Resource Standard Samples...... 116 Table 11.7 Analyzed Value of CDN Resource Standard Samples ...... 116 Table 11.8 Certified & Analyzed Value of Rocklabs Standard Samples ...... 121 Table 11.9 Samples Discarded From QA/QC Analysis Due to Likely Data Entry/Recording Error ...... 121 Table 11.10 Specific Gravity Determinations ...... 129 Table 11.11 Mean Specific Gravity and Standard Deviation for Major Veins ...... 131 Table 12.1 Micon Check Samples ...... 132 Table 14.1 Previous and Present Codes Used for Mineralized Veins ...... 138 Table 14.2 Mineralized Veins and Codes ...... 139 Table 14.3 Classical Statistics for the Mineralized Veins for Raw Assays for Gold from Underground Channel Samples ...... 140 Table 14.4 Classical Statistics for the Mineralized Veins for Raw Assays for Gold from Drill Core ...... 141 Table 14.5 Classical Statistics for the Mineralized Veins, Capped Assays for Gold from Underground Channel Samples ...... 141 Table 14.6 Classical Statistics for the Mineralized Veins for Capped Assays for Gold from Drill Core ...... 142 Table 14.7 Classical Statistics for the Mineralized Veins for Composited Assays for Gold ...... 143 Table 14.8 Classical Statistics for the Mineralized Veins for Composited Assays for Linear Metal Accumulation ...... 144 Table 14.9 Classical Statistics for the Mineralized Veins for Estimated Blocks for Gold ...... 145 Table 14.10 Classical Statistics for the Mineralized Veins for Estimated Blocks for Linear Metal Accumulation ...... 146 Table 14.11 Grade Capping Including Number of Samples Affected and Percentile at Which Capping is Applied ...... 147 Table 14.12 Summary of Search Volume Reference Number, Estimation Method and Vein ...... 154

Table 14.13 Summary of Search Parameters ...... 154 Table 14.14 Summary of Search Volume Reference Number, Direction and Angle of Rotation ...... 155 Table 14.15 Block Model Parameters ...... 156 Table 14.16 Mineral Resource Estimate at Different Grade and Thickness Cut-offs...... 165 Table 14.17 Mineral Resource Estimate at 5.0 g/t Au Cut-off Grade With Different Vein Thickness ...... 166 Table 14.18 Classified Mineral Resource Estimate at 5 g/t Gold Cut-off Grade and 0.1 m Minimum Width, Diluted to a Minimum Width of 1.0 m ...... 167 Table 14.19 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper ...... 167 Table 16.1 RQD and Joint Data by Lithology ...... 172 Table 16.2 Percentage Dilution ...... 174 Table 16.3 Measured, Indicated and Inferred Mineral Resources Considered in the Mine Plan ...... 175 Table 16.4 Total Underground Development for the Curraghinalt Project ...... 181 Table 16.5 Estimated Ventilation Requirement ...... 182 Table 16.6 Curraghinalt Project Annual Production Plan ...... 185 Table 16.7 Estimated Underground Service Equipment ...... 188 Table 16.8 Estimated Underground Service Equipment ...... 189 Table 16.9 Underground Mobile Equipment Fleet ...... 189 Table 16.10 Estimate Mine Labour Requirement ...... 190 Table 16.11 General and Administrative Manpower ...... 191 Table 17.1 Production Summary for the Dalradian Resources Tyrone Project Milling Plant ...... 194 Table 17.2 Process Design Criteria ...... 195 Table 17.3 Summary of Estimated Processing Plant Operations Personnel ...... 199 Table 20.1 Key Applicable Environmental Legislation ...... 212 Table 21.1: Capital Cost Summary ...... 216 Table 21.2 Capital Cost Summary - Mine Development ...... 216 Table 21.3 Capital Estimate - Mining Equipment ...... 217 Table 21.4 Process Plant Capital Estimate ...... 217 Table 21.5 Infrastructure Capital Estimate ...... 218 Table 21.6 Indirect Capital Cost Estimate ...... 218 Table 21.7 Summary of Life-of-Mine Operating Costs ...... 220 Table 22.1 Estimated Cost of Equity ...... 222 Table 22.2 Metal Price Averages ...... 223 Table 22.3 Life-of-Mine Cash Flow Summary ...... 227

Table 22.4 Base Case Life of Mine Annual Cash Flow ...... 229 Table 22.5 Base Case Cash Flow Evaluation ...... 230 Table 22.6 Sensitivity to Metal Prices ...... 232 Table 23.1 Galantas Mineral Resource Estimate ...... 234 Table 25.1 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper ...... 238 Table 25.2 Curraghinalt Initial Capital Expenditures ...... 240 Table 26.1 Exploration and Development Program Summary Budget ...... 244

List of Figures Page

Figure 1.1 General Location Map...... 2 Figure 1.2 Isometric View of Curraghinalt Project Mine Layout ...... 10 Figure 1.3 Option A Whole Ore Leach ...... 14 Figure 1.4 Option D Copper Flotation and Pyrite Flotation Concentrate Leach ...... 15 Figure 1.5 Monthly Average Gold and Silver Prices since July 2006 ...... 25 Figure 1.6 Annual Mining Schedule...... 26 Figure 1.7 Annual Processing Schedule ...... 26 Figure 1.8 Annual Production Schedule ...... 27 Figure 1.9 Direct Operating Costs ...... 27 Figure 1.10 Capital Expenditures ...... 28 Figure 1.11 Life-of-Mine Cash Flows ...... 29 Figure 1.12 Distribution of Gold and Silver Revenues ...... 31 Figure 1.13 Sensitivity Diagram ...... 33 Figure 1.14 Sensitivity to Metal Prices ...... 33 Figure 1.15 Sensitivity to Process Flowsheet ...... 34 Figure 4.1 General Location Map...... 42 Figure 4.2 Curraghinalt Gold Deposit Location ...... 47 Figure 4.3 Locations of Principal Veins and Adit on DG1 ...... 47 Figure 5.1 Tyrone Project Licence Map, Major Road Access and Drainages ...... 49 Figure 5.2 General View of the Curraghinalt Deposit Area Looking South ...... 50 Figure 6.1 Ennex Drilling at Curraghinalt ...... 54 Figure 6.2 Ennex Trenching at Curraghinalt ...... 55 Figure 7.1 Regional Geology of the North of Ireland ...... 60 Figure 7.2 Local Geology in the Area of Curraghinalt ...... 61 Figure 7.3 Plan of Curraghinalt Vein System on Level 170 ...... 65

Figure 7.4 Appalachian Correlations ...... 69 Figure 7.5 Simplified Geological Map of the Project Area ...... 75 Figure 7.6 Tyrone Volcanic Group Mineralized Locations ...... 79 Figure 8.1 Mineralization Models and Target Types on DG1 ...... 84 Figure 9.1 Location of Holes Drilled by Tournigan ...... 88 Figure 9.2 Location of Prospecting Samples, 2011 ...... 91 Figure 9.3 Exploration Targets ...... 92 Figure 9.4 Airborne Geophysical Survey Coverage ...... 95 Figure 9.5 Area to be Soil Sampled in Relation to Existing Grid ...... 96 Figure 10.1 North-South Schematic Section at 258555 E, Looking West ...... 101 Figure 10.2 Curraghinalt Project, Drill Hole Locations ...... 102 Figure 10.3 Location of Drill Holes Drilled After the 2011 Resource Estimate Cut-off Date ...... 104 Figure 10.4 Quartz-Sulphide Vein in Drill Core ...... 107 Figure 11.1 Results for Blanks Analyzed by Ennex ...... 114 Figure 11.2 Results for Dalradian Resources Blanks Analyzed by OMAC ...... 115 Figure 11.3 Error Plot for the Standard Samples Used by OMAC ...... 117 Figure 11.4 Repeatability by OMAC on Reference Material (3.4 g/t Au) ...... 117 Figure 11.5 Repeatability by OMAC on Reference Material (5.1 g/t Au) ...... 118 Figure 11.6 Repeatability by OMAC on Reference Material (7.47 g/t Au) ...... 118 Figure 11.7 Repeatability by OMAC on Reference Material (9.98 g/t Au) ...... 119 Figure 11.8 Repeatability by OMAC on Reference Material (15.31 g/t Au) ...... 119 Figure 11.9 Repeatability by OMAC on Reference Material (20.6 g/t Au) ...... 120 Figure 11.10 Error Plot for Standard Samples Used by OMAC ...... 121 Figure 11.11 Curraghinalt Drilling Standard SH35 X-Chart (OMAC) ...... 122 Figure 11.12 Curraghinalt Drilling Standard SH41 X-Chart (OMAC) ...... 122 Figure 11.13 Curraghinalt Drilling Standard SJ39 X-Chart (OMAC) ...... 123 Figure 11.14 Curraghinalt Drilling Standard SJ53 X-Chart (OMAC) ...... 123 Figure 11.15 Curraghinalt Drilling Standard SL46 X-Chart (OMAC) ...... 124 Figure 11.16 Curraghinalt Drilling Standard SL51 X-Chart (OMAC) ...... 124 Figure 11.17 Curraghinalt Drilling Standard SN38 X-Chart (OMAC) ...... 125 Figure 11.18 Curraghinalt Drilling Standard SN50 X-Chart (OMAC) ...... 125 Figure 11.19 Curraghinalt Drilling Standard SP37 X-Chart (OMAC) ...... 126 Figure 11.20 Curraghinalt Drilling Standard SQ36 X-Chart (OMAC) ...... 126 Figure 11.21 Scatter Plot between Original and Re-submitted Samples ...... 127 Figure 11.22 Error Plot for Gold Assays ...... 128 Figure 11.23 Plot of Gold Grade against Specific Gravity...... 130 Figure 14.1 Typical Geological Section for the Curraghinalt Area ...... 139

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Figure 14.2 Schematic 3D Disposition of Different Mineralized Veins for the Curraghinalt Area ...... 140 Figure 14.3 Probability Plot for Vein 10 of Curraghinalt Deposit ...... 148 Figure 14.4 Probability Plot for Vein 20 of Curraghinalt Deposit ...... 148 Figure 14.5 Probability Plot for Vein 30 of Curraghinalt Deposit ...... 149 Figure 14.6 Probability Plot for Vein 40 of Curraghinalt Deposit ...... 149 Figure 14.7 Probability Plot for Vein 50 of Curraghinalt Deposit ...... 150 Figure 14.8 Probability Plot for Vein 60 of Curraghinalt Deposit ...... 150 Figure 14.9 Probability Plot for Vein 70 of Curraghinalt Deposit ...... 151 Figure 14.10 Probability Plot for Vein 80 of Curraghinalt Deposit ...... 151 Figure 14.11 Probability Plot for Vein 90 of Curraghinalt Deposit ...... 152 Figure 14.12 Probability Plot for Vein D of Curraghinalt Deposit ...... 152 Figure 14.13 Probability Plot for Vein G of Curraghinalt Deposit ...... 153 Figure 14.14 North-south Cross-section Showing Block Model Grades for Mineralized Veins for Curraghinalt Deposit ...... 157 Figure 14.15 Plan View at 170 m RL of the Block Model Categorized into Different Grades for Curraghinalt Deposit ...... 158 Figure 14.16 Schematic 3D Rendering of Block Model Categorized into Different Grades for Curraghinalt Deposit ...... 158 Figure 14.17 Comparison of De-clustered Drill Hole Data with Resource Model Block Grade for Curraghinalt Deposit ...... 159 Figure 14.18 Comparison of De-clustered Drill Hole Data with Resource Model Block Grade for Curraghinalt Deposit ...... 159 Figure 14.19 Omni-directional Semi-variogram Model Using Data from Vein 60 ...... 160 Figure 14.20 Comparison of Two Methods of Estimation ...... 160 Figure 14.21 Vertical Longitudinal Projection of Vein 60 ...... 162 Figure 14.22 North-South Cross-Section Showing Resource Classifications for Curraghinalt Deposit ...... 162 Figure 14.23 Plan View at 170 m RL Showing Resource Classifications for the Curraghinalt Deposit ...... 163 Figure 14.24 Schematic 3D View of the Block Model for the Curraghinalt Deposit ...... 163 Figure 14.25 Grade-Tonnage Distribution of the Measured and Indicated Mineral Resource at Different Thickness Cut-off ...... 164 Figure 14.26 Grade-Tonnage Distribution of the Inferred Mineral Resource at Different Thickness Cut-off ...... 164 Figure 16.1 Typical Schematic of Longitudinal Sublevel Retreat with Multiple Mining Levels ...... 177 Figure 16.2 Isometric View of Curraghinalt Project Mine Layout ...... 178 Figure 16.3 Isometric View of the Mine Layout with Mineralized Veins ...... 179

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Figure 16.4 Typical Plan View of the Mine Layout (at 180 m Elevation) ...... 180 Figure 16.5 Production Drilling in Narrow Vein ...... 183 Figure 16.6 Annual Production Tonnage ...... 186 Figure 16.7 Annual Gold Grades ...... 186 Figure 16.8 Annual Silver and Copper Grades...... 187 Figure 17.1 Option A Whole Ore Leach ...... 192 Figure 17.2 Option B Bulk Flotation Concentrate Leach Whole Ore Leach ...... 193 Figure 17.3 Option C Bulk Flotation Concentrate - No Leach ...... 193 Figure 17.4 Option D Copper Flotation and Pyrite Flotation Concentrate Leach ...... 193 Figure 18.1 Schematic Surface Layout...... 204 Figure 20.1 Ecological and Landscape Designations ...... 206 Figure 20.2 Environmental Monitoring Locations ...... 207 Figure 20.3 Landscape of the Project Area ...... 208 Figure 20.4 Cultural Heritage Sites ...... 213 Figure 22.1 Real Return on Canadian Long Bonds ...... 222 Figure 22.2 Monthly Average Gold and Silver Prices since July 2006 ...... 223 Figure 22.3 Annual Mining Schedule...... 224 Figure 22.4 Annual Processing Schedule ...... 225 Figure 22.5 Annual Production Schedule ...... 225 Figure 22.6 Direct Operating Costs ...... 226 Figure 22.7 Capital Expenditures ...... 226 Figure 22.8 Life-of-Mine Cash Flows ...... 227 Figure 22.9 Distribution of Gold and Silver Revenues ...... 228 Figure 22.10 Sensitivity Diagram ...... 231 Figure 22.11 Sensitivity to Metal Prices ...... 231 Figure 22.12 Sensitivity to Process Flowsheet ...... 233

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List of Abbreviations

Abbreviation Unit or Term ' minutes of longitude or latitude ~ approximately % percent < less than > greater than º degrees of longitude, latitude, compass bearing or gradient ºC degrees Celsius 2D two-dimensional 3D three-dimensional µm microns, micrometres Ag silver As arsenic Au gold CEC Crown Estate Commissioners CDN$ Canadian dollar(s) CIM Canadian Institute of Mining, Metallurgy and Petroleum cm centimetre(s) Co. County Cu Copper d Day DETI Department of Enterprise, Trade and Investment E east et al. and others EM electromagnetic, usually in reference to and EM geophysical survey EPCM Engineering, Procurement and Construction Management ES Environmental Statement FA fire assay ft foot, feet g/t grams per tonne g/t Au grams per tonne of gold GBP Great Britain Pound (Sterling) GPS global positioning system h, h/d hour(s), hours/day ha hectare(s) HMS Heavy Media Separation HP Horsepower HQ H-sized core, Longyear Q-series drilling system ICP inductively coupled plasma ICP-AES inductively coupled plasma-atomic emission spectrometry ID Inverse distance grade interpolation in inch(es) IP induced polarization (geophysical survey) kg kilogram(s) km, km2 kilometre(s), square kilometre(s) kW, kWh, kWh/t Kilowatt, kilowatt hours, kilowatt hours per tonne L litre(s) lb pound(s) LOM Life of Mine m metre(s) m3 cubic metre(s) m/s metres per second M million(s) Ma million years masl metres above sea level

Abbreviation Unit or Term mg milligram mm millimetre(s) mL millilitre(s) MIBC Methyl Isobutyl Carbinol (frother) Mn manganese Mt million tonnes Mt/y million tonnes per year MW Megawatt N north n.a. not applicable, not available Na sodium NaCN Sodium cyanide NI 43-101 Canadian National Instrument 43-101 NPV, NPVn Net Present Value, Net Present Value(annual discount rate) NQ N-sized core, Longyear Q-series drilling system NSR Net smelter return (royalty) OK ordinary kriging grade interpolation oz troy ounce(s) oz/ton troy ounces per short ton PAG Potentially Acid Generating (waste rock) Pb lead pH concentration of hydrogen ion (-log10 of) ppb parts per billion ppm parts per million, equal to grams per tonne (g/t) QA/QC quality assurance/quality control QEMSCAN Quantitative Evaluation of Minerals by SCANning electron microscopy QP qualified person RC reverse circulation ROM Run of Mine RQD rock quality designation (data) s second S south Sb antimony SD standard deviation SG specific gravity SI International System of Units t tonne(s) (metric) t/h tonnes per hour t/d tonnes per day t/m3 tonnes per cubic metre t/y tonnes per year ton, T short ton (2,000 lbs) UG Underground UK United Kingdom of Great Britain and Northern Ireland US United States of America US$ United States dollar(s) US$/oz United States dollars per ounce US$/t United States dollars per tonne VLF-EM very low frequency - electromagnetic geophysical surveys W west or Watt wt % percent by weight y year yd, yd3 yard, cubic yard Zn zinc

xvi

7KHFRQFOXVLRQVDQGUHFRPPHQGDWLRQVLQWKLVUHSRUWUHIOHFWWKHDXWKRUV¶EHVWMXGJPHQWLQOLJKW of the information available to them at the time of writing. The authors and Micon International Limited (Micon) reserve the right, but will not be obliged, to revise this report and conclusions if additional information becomes known to them subsequent to the date of this report. Use of this report acknowledges acceptance of the foregoing conditions.

This report is intended to be used by Dalradian Resources Inc. (Dalradian Resources) subject to the terms and conditions of its agreement with Micon. That agreement permits Dalradian Resources to file this report as a National Instrument 43-101 Technical Report with the Canadian Securities Regulatory Authorities pursuant to provincial securities legislation. Except for the purposes legislated under provincial securities laws, any other use of this UHSRUWE\DQ\WKLUGSDUW\LVDWWKDWSDUW\¶VVROHULVN

xvii

1.0 SU M M A R Y

1.1 SCOPE OF WORK

At the request of Dalradian Resources Inc. (Dalradian Resources), Micon International Limited (Micon) has prepared a Preliminary Economic Assessment (PEA) of the Curraghinalt gold deposit ZLWKLQ WKDW FRPSDQ\¶V Tyrone project in County Tyrone and County Londonderry, Northern Ireland. This Technical Report, prepared in accordance with the reporting standards and definitions required under Canadian National Instrument (NI) 43- 101, summarizes the results of that study. This PEA is based upon an estimate of the Curraghinalt mineral resources originally prepared by Micon in November, 2011.

The Tyrone project is indirectly held by Dalradian Resources through a wholly-owned subsidiary, Dalradian Gold Limited (Dalradian Gold), in whose name the licences and option agreements are registered. Where the distinction is considered minor, the two companies are referred to herein simply as Dalradian.

The scope of the proposed project includes the development of an underground mine to extract the mineralized vein material that comprises the resource, treatment of that material in a conventional crushing, milling and carbon-in-leach plant to extract the gold and silver content, and disposal of the leached residue underground using paste backfill to the extent possible, with excess material stored in a purpose-built tailings storage facility near the mine site. At the proposed scale of operation, the mine has a projected life of about 15 years.

1.2 LOCATION AND DESCRIPTION

The Tyrone project is located in County Tyrone and County Londonderry, Northern Ireland. The Curraghinalt gold deposit, located near the centre of the property, in County Tyrone, is approximately 127 km west of Belfast by road and 15 km northeast of the town of Omagh. Access to the Curraghinalt deposit is via a number of paved highways and local roads. These include the B48 which runs from Omagh to Gortin and the B46 which runs from Gortin to Greencastle. Private roads and farm tracks provide access within the property.

The relevant property is the subject of four Mineral Prospecting Licences and four Option Agreements. The four Mineral Prospecting Licences are issued by the Department of Enterprise, Trade and Investment (DETI) and the four Option Agreements are issued by the Crown Estate Commissioners (CEC).

In Northern Ireland, DETI is responsible for the granting of prospecting licences and mining licences for the exploration and development of minerals. Dalradian Resources, through its subsidiary Dalradian Gold, presently has the benefit of the following DETI prospecting licences: DG1/08, DG2/08, DG3/11 and DG4/11.

The CEC are responsible for the licensing of precious metal exploration. Prior to 2011, the CEC issued prospecting licences similar to DETI. In 2011, the CEC switched to a system of

option agreements, whereby the option holder can exercise the option to convert the property into a lease. Dalradian Resources, through its subsidiary Dalradian Gold, also has the benefit of the following CEC option agreements: DG1, DG2, DG3 and DG4. (It should be noted that in this document any reference to ³TG-3´ and ³TG-4´ rather than to ³DG3´ and ³DG4´ is a carryover from the previous ownership period by Tournigan Gold (hence ³TG´) and has no practical bearing or significance to the prospecting licences or option agreements.)

The 4 DETI licences correspond to the 4 CEC option agreements thereby granting base and precious metal mineral exploration rights to four contiguous pieces of ground. The licences cover a total area of some 844 km2 (see Figure 1.1).

Figure 1.1 General Location Map

Source: Dalradian Resources, 2011

The four pieces of property are often referred to simply as DG1, DG2, DG3 and DG4 or DG- 1, DG-2, DG-3 and DG-4 (with or without hyphens). In this report they shall be referred to as DG1, DG2, DG3 and DG4 although some older figures provided herein show them as DG-1, DG-2, DG-3 and DG-4.

The approximate centre of the two most important licence areas, DG1 and DG2, is located at latitude 54o ¶ ´QRUWKDQGORQJLWXGHo ¶ ´ZHVW

The mineral exploration rights are owned by a former Tournigan subsidiary, Dalradian Gold, which was acquired by Dalradian Resources in a transaction dated September 28, 2009. They are renewable, subject to certain requirements, for two-year periods from the date of issue and must have biennial reports submitted to the licensing authorities. At the end of six years (provided that it has endured that long), the licence is not continued; rather, a fresh licence must be issued. Dalradian Gold had previously provided the renewal reports to DETI in support of the licences.

Licences DG1 and DG2 (and the corresponding CEC option agreements) were reissued in January, 2008. The licences were issued for a term which may total up to six years (subject to the conditions described in Section 4 of this report). A first renewal of these licences was granted in January, 2010 and a second renewal was granted in November, 2011 with an effective date of January, 2012. Reissue of these licences is scheduled for January 1, 2014.

Licences DG3 and DG4 (and the corresponding CEC option agreements) were due for their first extension in April, 2011. Due to timing problems in the renewal process, the licences were reissued as new licences in April, 2011. The licences were issued for a term which may total up to six years (subject to the conditions described in Section 4 of this report) with reissue scheduled for April 24, 2017.

1.3 GEOLOGY AND MINERALIZATION

The Tyrone project contains mesothermal gold mineralization, with gold disseminated in quartz-sulphide veins, at the Curraghinalt deposit on licence DG1. The general strike direction of the quartz-sulphide veins is approximately west-northwest and the dip is approximately 60º to 70º to the north. Additionally, on licence DG2, there are known occurrences of gold mineralization in silicified rhyolite breccias, and porphyry-style copper mineralization. It has also been postulated that the DG2 licence is also a potential host for volcanogenic massive sulphide (VMS) style mineralization due to stratigraphic correlations with mineral belts in Newfoundland and Scandinavia.

The geological history of the project and the immediate surrounding area is related to the closing of the Iapetus Ocean in Ordovician time.

The Curraghinalt deposit is located 3 km to the north of the northeast-southwest striking Omagh Thrust Fault. This major fault has thrust Dalradian Supergroup rocks from the

northwest over the Ordovician-aged Tyrone Volcanic Group (TVG) rocks, part of the Tyrone Igneous Complex, located to the south. The Dalradian metasediments on the northern side of the thrust strike northeast-southwest and dip to the northwest. The lithologies are interpreted to lie on the lower limb of a gently to moderately northwest dipping, southeast upward- facing, major recumbent isoclinal fold, referred to as the Sperrin overfold or the Sperrin Nappe.

The Dalradian stratigraphy in the DG1 licence area is inverted, occupying the lower limb of the overturned Sperrin Nappe. The Dalradian formations present in the deposit area are, from the southeast (at the Omagh Thrust) to the northwest and, from the youngest to the oldest, the Mullaghcarn, Glengawna and Glenelly Formations. The Mullaghcarn Formation, which hosts the Curraghinalt deposit, consists of mixed semi-pelites, quartz semi-pelites and psammites. The Glengawna Formation comprises a mixed package of graphitic pelites, graphitic semi-pelites, chloritic semi-pelites and minor psammites. Furthest north on licence DG1 is the Glenelly Formation, which is comprised of pelites, semi-pelites and quartz semi- pelites. Further to the north on licences DG3 and DG4 lie more clastic sediments and limestones of the Dalradian along with a small area of unconformable Carboniferous sandstone.

Mineralized vein structures at the Curraghinalt deposit represent composite veins that have been developed over time by four distinct periods of quartz injection and sulphide-rich hydrothermal fluids. Mineralized veins are best developed in the relatively brittle semi- pelites and psammites, rather than in the more ductile pelitic horizons. Barren quartz veins represent, for the most part, the first stage of quartz injection that was essentially barren in gold. These veins are generally quite narrow, generally occur in pelite units, and fill the available shear, thus effectively sealing the shear for later stage gold-bearing quartz injection. Available evidence suggests that the hosting structures for these veins, or parallel ones, may strike onto the DG3 and DG4 licences.

The upper part of the TVG is comprised of bimodal sequences of basaltic and andesitic to rhyolitic submarine and subaerial lavas, volcaniclastics with chert horizons and intercalations of graptolitic shales. These lithologies are preceded by submarine basaltic andesite pillow lavas and associated intrusives. A series of porphyry bodies and a series of calc-alkaline granitic intrusions intrude the volcano-sedimentary sequence.

Work by Earls (1987) and others suggests the presence of three island arc volcanic cycles favourable for VMS deposits, adjacent to the back arc basal ophiolites, within the TVG on licence DG2. These cycles are designated from top to bottom as Cycle 3, the Broughderg Formation, Cycle 2, the Bonnety Bush-Glencurry Formation and Cycle 1 the Dungate- Tanderagee Formation. The tops of each cycle are marked by an erratic increase in volumetrically minor chert or ironstone, which is taken to represent the culmination of a cycle.

1.4 HISTORY

The DG1 and DG2 licence areas were acquired initially, in 1981, by Ulster Base Metals Limited (later known as Ulster Minerals) an entity which later became a wholly-owned subsidiary of Ennex International plc (Ennex). Ennex conducted exploration on the property between 1986 and 1999. Ennex sold its interest in Ulster Minerals to Nickelodeon Minerals Inc. (Nickelodeon) in January, 2000. In August, 2000, the name of Nickelodeon was changed to Strongbow Resources Inc. and, subsequently, to Strongbow Exploration Inc. (Strongbow).

In February, 2003, Tournigan entered into an option agreement with Strongbow to earn an interest of up to 100% in the Curraghinalt deposit. At the same time, Tournigan entered into a similar option agreement with Strongbow in respect of its ³7\URQHSURMHFW´LQWKH79* Tournigan established Dalradian Gold as a wholly-owned subsidiary through which it would earn its interests in the Curraghinalt and Tyrone properties which were subsequently converted to licences UM-1/02 and UM-2/02.

In the following year (February, 2004), Tournigan entered into a letter agreement with Strongbow for the outright purchase by Tournigan of all of the issued and outstanding shares of Ulster Minerals. A net smelter royalty of 2% held by Ennex was transferred to Minco plc. Full transfer of ownership in Ulster Minerals to Tournigan was completed in December, 2004.

During this period, Tournigan also applied for and received the licences which later became DG3 and DG4 and lie to the northwest of licences UM-1/02 and UM-2/02. As well, licences UM-1/02 and UM-2/02 were converted in name to licences DG1 and DG2 and their internal boundary was moved to reflect the location of the Omagh Thrust, the boundary between the Dalradian metasediments and the TVG.

As described above, in September, 2009, Dalradian Resources entered into an agreement with Tournigan to purchase Dalradian Gold and all of its Northern Ireland assets.

Prior to the work of Ennex/Ulster Base Metals gold was recognized in the gravels of the Moyola River as early as 1652 and, in the 1930s, an English company reported plans for alluvial gold mining in a prospectus, but very little work appears to have occurred. The licences were held by other companies prior to Ulster Base Metals and the Geological Survey of Northern Ireland (GSNI) completed a report on the gold potential of the area.

Between 1983 and 1997 Ennex conducted a significant amount of exploration drilling at Curraghinalt and other targets in the area such as Cashel Rock. Ennex also completed a program of underground development on two of the veins at Curraghinalt. Nickelodeon and Strongbow completed very little work on the property.

Due to the relatively advanced stage of historical exploration completed at the Curraghinalt deposit by the previous operators of the property, in 2003 Tournigan initially moved directly

to infill and deeper drilling and only undertook a small amount of additional exploration. Some geophysical surveys had been carried out on a regional scale to delineate drill targets along strike and some prospecting, sampling and mapping, sufficient to retain the other licences, was completed. Tournigan also compiled an extensive database of available historical exploration information and data.

The drill program allowed Tournigan to report a mineral resource estimate for Curraghinalt in 2007 consisting of both indicated and inferred resources (Mukhopadhyay, 2007). This estimate was performed by Micon. After completion of the resource estimate, and before halting exploration at the deposit, Tournigan completed four more holes to intersect deeper mineralization which were not available to be used in completion of the resource estimate at the time.

1.5 EXPLORATION

After acquisition of the property by Dalradian Resources, Micon modified the geological model to take into account the results from the last four Tournigan holes and published an updated mineral resource estimate (Hennessey and Mukhopadhyay, 2010) in a Technical Report supporting the listing of the company.

Since acquisition of the Tyrone project, Dalradian Resources has embarked upon a campaign of drilling largely concentrated on the Curraghinalt deposit. As of the November, 2011 resource update (Hennessey and Mukhopadhyay, 2012) this program consisted of an additional 45 diamond drill holes totalling 20,561.9 m as described in Section 10 of this report. As a result the company determined that it was then appropriate to update the mineral resource estimate for the Curraghinalt deposit.

Since the completion of the mineral resource estimate employed in this PEA (Hennessey and Mukhopadhyay, 2012) Dalradian Resources has completed an additional 57 holes for 15,677 m (see Section 11). The company has also completed 1,034 line-km of helicopter-borne electromagnetic and magnetic geophysical surveys and commenced an expansion of the soil geochemistry grid (see Section 10).

1.6 MINERAL RESOURCES

In order to prepare the updated mineral resource estimate, the mineralized area was again divided into different zones with statistical analysis carried out separately on all zones. Resources were estimated by the inverse distance interpolation method with selected checks on larger veins by ordinary kriging. Vein thickness and metal accumulation (grade times thickness) were interpolated into the model with block grade determined by division of the metal accumulation by interpolated thickness.

The resource has been classified following the standards and definitions of the Canadian Institute of Mining, Metallurgy and Petroleum (CIM). The mineral resource has been reported at an economic cut-off grade of 5 g/t Au over a minimum horizontal thickness of

1 m. All mineralized blocks in the block model which were greater than 0.10 m in thickness were diluted to 1 m at zero grade (for the diluting material) and reported if they then met the cut-off grade.

The results of the 2011 mineral resource estimate, as originally reported in Hennessey and Mukhopadhyay (2012) are presented in Table 1.1. This estimate is based on exploration data available as of October 10, 2011 and is current as of November 30, 2011.

Table 1.1 Classified Mineral Resource Estimate at 5 g/t Gold Cut-off Grade and 0.1 m Minimum Width, Diluted to a Minimum Width of 1.0 m

Contained Metal Resource Million G rade Million Category Tonnes (g/t Au) Tonnes Oz Measured 0.02 21.51 0.44 0.01 Indicated 1.11 12.84 14.20 0.46 Measured + 1.13 13.00 14.65 0.47 Indicated Inferred 5.45 12.74 69.44 2.23

After the public disclosure of the Curraghinalt mineral resource estimate on November 30, 2011, and its publication in Hennessey and Mukhopadhyay (2012), silver and copper grades were estimated into the block model. This was done to allow for the contribution of silver to WKHFDVKIORZIRUWKH3($DQGWRPRGHOFRSSHU¶VHIIHFWRQF\DQLGHFRQVXPSWLRQ6LOYHUDQG copper grades were estimated using the same techniques as described for gold above. In the PEA base case, no copper concentrate is produced and no revenue from copper is in the cash flow. Silver accounts for approximately 1% of revenue forecast in the economic analysis.

The amended Curraghinalt deposit mineral resource estimate showing copper and silver grades is set out in Table 1.2 below. No changes in tonnages or gold grades have occurred. In order to make it comparable to Table 1.1, Table 1.2 is also reported at a cut-off grade of 5.0 g/t Au. No gold equivalent calculations were made.

Table 1.2 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper

M illion Au Ag Cu Category Notes Tonnes (g/t) (g/t) (%) Measured 0.02 21.51 17.56 0.49 Indicated 1.11 12.84 4.05 0.17 Measured + Indicated 1.13 13.00 4.29 0.18 With Cu* 5.38 12.77 5.48 0.12 Inferred Without Cu* 0.08 10.37 2.00 - * Copper assay data were not available for the 752 and D veins and no copper grades were estimated there. 1 - Original vein thickness blocks of less than 0.10 m were eliminated from the resource tabulation. There were minor differences between a 0.00-m and 0.10-m minimum thickness estimates which disappear with

rounding. For horizontal vein thicknesses of more than 0.10 m, but less than 1.0 m, the grades were diluted to a minimum horizontal width of 1.0 m at zero grade prior to reporting. 2 - Mineral resources are reported at a cut-off grade of 5 g/t Au after dilution. 3 - Mineral resources are not mineral reserves and do not have demonstrated economic viability. 4 - The mineral resource estimate was prepared by Dibya Kanti Mukhopadhyay, MAusIMM (CP), under the overall direction of B. Terrence Hennessey, P.Geo. Mr. Hennessey has taken responsibility for the estimate in this report. The Curraghinalt resource estimate is compliant with the current standards and definitions required under NI 43-101.

To the best knowledge of the authors, the stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues. There are no known mining, metallurgical, infrastructure or other factors that materially affect this mineral resource estimate.

The Curraghinalt mineral resource estimate is compliant with the current CIM standards and definitions as required under NI 43-101 and is, therefore, reportable as a mineral resource by Dalradian Resources.

1.7 MINING

Mineralized veins at the Curraghinalt deposit will be extracted using the Longitudinal Sublevel Retreat Underground Mining method with both paste and waste backfill. This backfill method was chosen in order to maximize the extraction of the high value resources, reduce the surface footprint required for tailings disposal, and allow savings in the volume and transportation cost of waste rock from underground development.

Access to the Curraghinalt deposit will be via a main decline from the surface. The main decline will be connected by cross-cuts to sublevels providing access to the mineralized veins. Sills, excavated along the mineralized veins on each sublevel, provide access to the mining stopes where production drilling, blasting and extraction of the mineralized material take place.

The extraction of mineralized material from underground will be carried out with rubber tired mechanized equipment to maximize production and provide flexibility to the underground operations. Some key parameters are:

Mining Method is longitudinal sublevel retreat. Decline or ramp dimensions are 4.5 m H x 4.5 m W at 15% grade. By-passes at the decline or ramp at 4.5 m H x 4.5 m W x 3.0 m L. Safety bays along the decline and ramp at 3.0 m H x 3.0 m W x 3.0 m L. Sumps at 3.0 m H x 4.0 m W x 3.0 m L. Cross-cut dimension at 4.0 m H x 4.0 m W. Sill or development in ore at 3.0 m H x 3.0 m W. Refuge station/lunch room at 4.0 m H x 4.0 m W x 10.0 m L. Drift to ventilation shafts and remuck bays at 4.0 m H x 4.0 m W. Ventilation raise and ore or waste-pass at 3 m diameter.

Level intervals at 20 vertical metres from the floor of the top sill to the floor of the bottom sill.

The mineral resource considered in the mine plan for Curraghinalt was estimated based on the following parameters:

Tonnages above the 5.0 g/t Au cut-off grade.

Non-recoverable crown pillar extending 20 m below and parallel to the topography.

Exclusion from the mine plan of resources below the -490 m level.

Mining recovery of 95%.

Dilution of mineral resource to minimum mining width of 1.8 m for stopes and 3 m for sills.

Table 1.3 summarizes the mineral resource considered in the mine plan, diluted to 1.8 m mining width at 95% recovery with external dilution.

Table 1.3 Measured, Indicated and Inferred Mineral Resources Considered in the Mine Plan

Avg. Avg. Avg. Tonnage Description Au Ag Cu (Mt) (g/t) (g/t) (%) Measured 0.02 16.46 14.19 0.40 Indicated 1.42 8.76 2.77 0.12 Total Measured & Indicated 1.44 8.88 2.95 0.12 Total Inferred 7.98 7.91 3.38 0.08

Sills at 3.0 m H x 3.0 m W in the mineralization veins will be developed when the cross-cut intercepts the veins. The sills will be developed in both directions along the strike length of the mineralized zone, providing two production faces for each vein on each level.

Mining in a production level will commence from one end of the veins, retreating to the OHYHO¶V FURVVFXW 7KH DYDLODELOLW\ RI QXPHURXV SURGXFWLRQ KHDGLQJV ZKHQ D FURVV-cut intercepts the veins within the level provides multiple production headings to meet the 1,700 t/d production target.

The main decline into the mineralized zones can act as fresh air intake into the mine. Exhausted and contaminated air can be channelled through one of the two ventilation raises or directed into the existing exploration adit on 170 m elevation to the surface.

Ventilation from one level to another will be provided by the ventilation raises located in the cross-cuts. Ventilation drifts connect the ventilation raises to the cross-cuts to supply fresh or exhaust air from the underground workings.

Figure 1.2 presents typical schematic of a longitudinal sublevel retreat mining method with multiple mining levels. The total metreage of development in the LOM plan for the Curraghinalt deposit is presented in Table 1.4.

Figure 1.2 Isometric View of Curraghinalt Project Mine Layout

Table 1.4 Total Underground Development for the Curraghinalt Project

Description Total Length (m) Ramp 5,325 Ramp By Passes 157 Ramp Safety Bays 120 Sump in Ramp 120 &URVV&XWV ,QF5HIXJH6WQ¶V 6HUYLFH'ULIWV 16,614 Sumps in Cross Cut 137 Sill 207,420 Sill (Through Waste) 7,152 Vent, Ore & Waste Pass 2,990 Total 240,035

Table 1.5 summarizes the proposed equipment fleet required to develop and extract 1,700 t/d of mineralized material from the Curraghinalt deposit. The table presents the average number of equipment units over the life of mine.

Table 1.5 Underground Mobile Equipment Fleet

Description Units (Avg. L O M) Stoping Drill (e.g., Boart Stopemate) 4 Stoping Drill (Electric-hydraulic) 1 Sill Narrow Vein Jumbo 5 Development Jumbo (Double Boom) 1 Bolter 1 LHD at 4.0 m3 1 LHD at 3.0 m3 4 Trucks - 30 t 3 Explosive Truck 1 Scissor Truck 1 Fuel and Lube Truck 1 Mechanic Light Pickup 1 Electrician Light Pickup 1 Surveying Light Pickup 1 Light Pick-Up (Service and Supervision) 3 Man Carrier 1 UG Grader 1

A summary of the underground manpower requirements is presented in Table 1.6.

Table 1.6 Estimate M ine Labour Requirement

Description Manpower (Avg. L O M) Stoping Production Driller 12 Sill Narrow Vein Jumbo Driller 13 Development Jumbo (Double Boom) Driller 2 Bolter Driller 3 Trucks Drivers 9 Backfill Crew 4 Grader Operator 3 Blaster - Production 4 Blaster - Development 6 Mechanics and Electricians 18 Surveyor 4 General Miner/Helper 4 Sampler 6 Total 56

,QDGGLWLRQWRWKHDERYHWKHSURMHFW¶VJHQHUDODGPLQLVWUDWLRQDQGWHFKQLFDOPDQDJHPHQWZLOO require staff as outlined in Table 1.7.

Table 1.7 General and Administrative Manpower

Description Manpower (Avg. L O M) General Administration General Manager/Mine Superintendent 1 Health & Safety and Training Officer 6 Mine Technical & Engineering Chief/Senior Mining Engineer 1 Intermediate Mining Engineer 3 Intermediate Draft Person 4 Senior Surveyor 1 Chief/Senior Geologist 1 Intermediate Geologist 3 Senior Mechanical Engineer 1 Mine Technicians 4 Salaried Staff - Mining Admin. Assistant 1 Mine General Foreman 1 Mine Supervisor/Shift Boss (Production) 6 Maintenance Supervisor/Shift Boss 3 Total 36

1.8 METALLURGY AND PROCESSING

1.8.1 Mineralogy and Metallurgical Testing

Mineralogical work has indicated that gold mineralization at Curraghinalt occurs in quartz- pyrite veins and is associated with variable abundances of carbonate, chalcopyrite and tennantite-tetrahedrite. In general, carbonate, chalcopyrite and tennantite-tetrahedrite are paragenetically later than quartz and pyrite, and fill fractures in the latter. Gold occurs mainly as the native metal and more rarely as electrum (>20 wt% Ag), and is found primarily along fractures in pyrite, as inclusions in pyrite, and at pyrite grain contacts with carbonate and quartz.

The PEA base case proposes a conventional crushing, milling and carbon-in-leach gold extraction flowsheet. Comminution tests have returned Bond ball mill work indices in the range 15.2 - 15.4 kWh/t. The abrasion index on composite sample 12-1B returned a value of 0.1278 g. Cyanidation testwork has generally returned very high metal extractions - typically 95% or better for gold and about 80% for silver. A grind of approximately 85% passing 200 mesh (75 µm) and 48 h leach at NaCN 1g/L was generally effective in the earlier testwork.

Additional testwork has investigated options for the pre-treatment of the mill feed using heavy media separation (HMS), the production of gravity and/or flotation concentrates, and the possible sale or intensive leaching of these concentrates. However, evaluation of these suggests the most effective flowsheet is the whole-ore leach reflected in the PEA base case.

1.8.2 Process Flowsheets Considered

The four process flowsheet options considered and economically evaluated for this PEA study are:

Option A: Whole Ore Leach Option B: Gravity concentration, sale of gravity concentrate followed by Bulk Flotation Concentrate, Leach of concentrate Option C: Gravity concentration followed by Bulk Flotation Concentrate - Sale of both concentrates Option D: Copper Flotation, with sale of the copper concentrate and Pyrite Flotation Concentrate with leach of the Pyrite concentrate

Operating and capital costs for each of these four options were developed, as well as revenue estimates. All four options had positive outcomes but, following a preliminary internal review, Options B and C were rejected due to their higher operating costs and lower revenues. Detailed estimates were developed for Options A and D and have been included in the PEA.

Options A, and D both involve a similar set of processing steps including crushing, grinding, and cyanidation. Option D includes 2 flotation steps producing a copper concentrate for sale and a pyrite concentrate for on-site cyanide leaching.

A summary of the project production for options A and D are presented in Table 1.8.

Table 1.8 Production Summary for the Dalradian Resources Tyrone Project Milling Plant

Item Option A Option D Whole O re Copper Flotation Leach and Pyrite Flotation Concentrate Leach Tonnes Processed Annually, Avg. (t/y) 443,179 443,179 Feed G rades (L O M)

Au (g/t) 8.06 8.06 Ag (g/t) 3.31 3.31 Cu (%) 0.08 0.08 Saleable Cu Concentrate Production, Avg. (t/y) - 5,585 Gold Production, LOM Avg. (oz/y) 114,841 102,622 Silver Production, LOM Avg. (oz/y) 47,162 38,681 K ey Operating Parameters

Operating Days Per Year 365 365 Mill Feed Rate, t/d 1,700 1,700 Chemical & Metallurgical Parameters

Cu Concentrate Grade, Cu % - 8.4

Figure 1.3 and Figure 1.4 show the schematic flowsheets for Options A and D respectively.

Figure 1.3 Option A Whole Ore Leach

FEED Cyanidation Tails

Refinery

Gold Bars

Figure 1.4 Option D Copper Flotation and Pyrite Flotation Concentrate Leach

Copper Flotation Pyrite Flotation Float Tails Feed Circuit Tails Copper Flotation

Copper 1st Cleaner Tails Cyanidation Leach Tails

Concentrate

Refinery

Gold Bars

1.8.3 Process Description (Option A)

The ore is delivered by mine trucks to the plant and fed over a grizzly with 400-mm openings to the Run Of Mine (ROM) ore feed hopper. The ore is extracted from the feed hopper by an apron feeder and fed to a jaw crusher. The crushed ore is discharged onto a conveyor which will transport the ore to the plant. The crusher product is 80% passing 150 mm.

The crushed ore reports to the grinding circuit. The primary grinding SAG mill discharge product typically has a size passing of 850 µm.

The secondary grinding ball mill will operate in closed circuit with a cyclone, with the target cyclone overflow 80% size passing 141 µm.

A six stage carbon-in-leach (CIL) circuit is installed based on a retention time of 24 hours and a new carbon concentration of 20g/L. Slurry pH is adjusted to 10.5 with the addition of lime. The first leach tank is aerated, and lead nitrate is added. The oxidation of sulphides in this step reduces cyanide consumption considerably as well as enhancing leach kinetics. The leach circuit feed is pumped to the train of six CIL tanks in series where it is contacted with activated carbon. Each of the tanks is provided with an in-tank screen to allow the slurry to flow by gravity through the CIL train, while retaining the carbon in the tanks. The carbon is transported counter-current to the slurry flow by means of carbon advance pumps. The loaded carbon from the first CIL tank is collected on a screen and is sent to the carbon stripping circuit. The slurry leaving the last CIL tank is passed over a safety screen to capture any carbon particles before it is sent to the cyanide destruction circuit.

A carbon stripping/carbon reactivation circuit is installed to treat the loaded carbon, to recover gold and silver, and to recycle the carbon to the CIL circuit.

The loaded carbon is acid washed and fed to a Zadra elution circuit to strip the gold from the carbon. The pregnant solution is then sent to two electrowinning cells to recover the gold

from solution. The carbon from the stripping circuit is fed to a diesel-fired rotary kiln to reactivate the carbon which is passed over a sizing screen prior to being recycled to the CIL circuit.

The stainless steel wool cathodes carrying the gold from the electrowinning cells are fed to an induction furnace to produce gold doré bars for sale.

A cyanide destruction circuit is installed to treat the slurry discharge from the CIL circuit.

The CIL discharge slurry is pumped to two agitated tanks in series, and SO2 and air are added to destroy the cyanide before the slurry is pumped to the tailings thickener. Thickened tailings at 60% solids are pumped to the pastefill plant or tailings disposal area.

1.8.4 Process Description (Option D)

The flowsheet is similar to Option A except that:

The ball mill ore slurry discharge, at 50% solids, is fed to a conditioning tank, which then RYHUIORZVWRWKHFRSSHUIORWDWLRQFLUFXLW7KHFRSSHUFLUFXLW¶VILUVWFOHDQHUWDLOLQJVUHSRUWWR the cyanide leach circuit. Tails from the circuit are pumped to the bulk flotation circuit.

The concentrate from the rougher cells is fed to the cleaner cells from which the final copper concentrate is thickened to 70% solids in conventional thickener and filtered using pressure filters. This product is then sold to a copper smelter.

Tails from copper flotation are subjected to a bulk sulphide flotation to recover all remaining sulphides, pyrite, etc., as well as gold and silver bearing minerals. This flotation concentrate is thickened and pumped to a CIL circuit for recovery of gold and silver.

1.8.5 Process Manpower

Option D includes the process operations of Option A, but also includes two additional flotation circuits. Thus, Option D (76 personnel) will require nine more than Option A (67 personnel). See Table 1.9.

Table 1.9 Summary of Estimated Processing Plant Operations Personnel

Option A: Whole Option D: Copper Float O re Leach and Pyrite Flotation Description Concentrate Leach Plant Superintendent 1 1 Senior Metallurgist 1 1 Metallurgist 0 1 Chief Chemist 1 1 Sample Prep. Technician 4 4 Assay Technician 2 2 Plant Shift Supervisor 4 4 Crusher Operator 4 4 Plant Operator 4 4 Control Room Operator 0 4 Operator Helper 4 4 Tailings Operator 2 2 Load out Operator 0 4 Cyanidation Plant Operator 4 4 Day Labour Crew 4 4 Maintenance Superintendent 1 1 Maintenance Supervisor 1 1 Instrument Technician 2 2 Electrician 4 4 Plant Mechanic 4 4 Security Personnel 5 5 Senior Accountant 1 1 Accountants/buyer 1 1 HR and Community Relations 1 1 Environmental Technician 1 1 Supervisor 1 1 Bus Driver 1 1 Secretaries/Admin Assistant 2 2 Site Surface Maintenance Crew 4 4 Equipment Operator 2 2 Site Administrator 1 1 Total 67 76

1.9 INFRASTRUCTURE

1.9.1 Access and Site Roads

The property is readily accessible by a network of existing all-weather paved highways and local roads. Micon considers that ultimately the property should be accessed from the south (off the B46 Gortin-Greencastle road).

Site roads will be built to connect the onsite facilities: gate house, administration complex, process plant, mine dry, warehouse, ore and waste storage pads, tailings storage facility, etc.

1.9.2 Power Supply

The base case electrical power requirement is 70 kWh/t which results in a maximum demand of approximately 6 MW for a 1,700 t/d operation.

Power will be supplied by connecting to the national high voltage electricity grid through a 33kV overhead transmission line that will be built from the 110/33kV substation located at Strabane, for approximately 27 km to the site. A substation will be built on site which will transform the electrical power from 33kV to 11kV for distribution around the site.

Emergency power will be provide by a 1.12 MW, 400 V standby diesel generator located on site.

1.9.3 Water Supply

Micon has assumed that supply of the initial and make-up water for the process plant will be through the construction of a rain water catchment system. Rainfall in the project area is in the order of 1,300 to 1,400 mm/y, and Micon anticipates the utilization of the proposed tailings storage facility to capture and store rain water.

Nevertheless, further studies to assess potential water sources should be carried out during further development of the project.

1.9.4 Tailings Storage Facility

In 2011, Golder Associates UK (Golder) was retained by Dalradian Resources to carry out a Tailings Management Facility Site Selection Study aimed at identifying potential sites to store approximately 2.92 million tonnes of tailings. Golder, in its December 2011 report, identified seven potential sites and recommended three of them for further investigation. Golder also recommended the use of conventional tailings disposal technology and to carry out the construction work in two separate phases.

1.10 ENVIRONMENTAL, PERMITTING AND SOCIAL

1.10.1 Studies and Issues

Micon has reviewed available environmental studies completed by SLR Consulting (SLR) and Golder, and compiled a summary of this work. Micon has not prepared a liability assessment of the property and cannot provide a legal opinion on the status of permits.

This area of the Sperrin Mountains is designated an Area of Outstanding Natural Beauty. In addition, there are a number of protected and special interest areas around the project. The nearest Areas of Special Scientific Interest (ASSIs) to the project are Mullaghcarn/ Mountfield Quarry, Murrins, Cashel Rock, Boorin Wood and Black Bog. The nearest

Special Areas of Conservation (SACs) include Drumlea and Mullan Woods, Owenkillew River, and Black Bog.

Environmental baseline studies were initiated by SLR for Dalradian in June, 2010 and have included collecting data on meteorology, hydrology, hydrogeology, water quality, sediment quality, acid rock generation potential of the mineral and waste rock, flora, terrestrial and aquatic fauna, air quality, visual resources, cultural heritage resources, and the local socio- economy.

1.10.2 Waste and Water Management

Geochemical investigations were undertaken by SLR to determine the potential for acid generation in the proposed operation. Thirty-five samples were collected in December 2011 to represent rock types likely to be encountered during the mining operation. The test work undertaken included acid base accounting (ABA), net acid generation (NAG) and acid rain leach procedure (ARLP) tests aimed to determine whether and which of the mined materials would be acid generating and/or metal leaching.

Based on the ABA laboratory results, it was concluded that 5 samples were considered to be acid generating, 23 were classified as non-acid generating and the results of 7 samples were considered to be inconclusive. Potentially acid generating waste rock will be backfilled underground and ultimately flooded after closure which will minimize future oxidation and metal release. The remaining waste rock will be stored in surface dumps, contoured and eventually capped and revegetated in character with the surrounding landscape.

An estimated 40% of tailings will be backfilled underground. The remaining tailings will be disposed of on surface in a conventional tailings impoundment which will be lined with synthetic and/or impermeable material. It is recommended that the project consider paste or dry stack tailings to minimize the footprint and increase the available options for siting the tailings and progressive reclamation to minimize visual impacts.

Mine water will be collected in underground sumps and pumped to surface for use as makeup water in the process plant. This water may need to be treated prior to use in the process, depending on the concentrations of suspended sediments and blasting residues.

The process plant is designed to include an INCO SO2/air cyanide destruction circuit to meet the limit of 10 ppm weak acid dissociable (WAD) cyanide at the point of discharge to the tailings pond as required under Directive 2006/21/EC, Article 13(6). Slurry tailings will be pumped to the tailings impoundment and pond water will be recycled back to the process plant via a reclaim barge and pipeline. The impoundment will need to be designed to minimize seepages. Monitoring wells will then be required at the tailings impoundment to monitor seepages and could be used for seepage collection and pump back to the pond if necessary.

Post-closure, the adit will be plugged and underground workings allowed to flood. For the tailings impoundment at closure, it is assumed tailings pond water will be recycled back through the plant and treated as necessary until the pond water quality is sufficient for direct release of any annual surplus water. Groundwater and surface waters will need to be monitored post-closure.

1.10.3 Social Management

SLR identified social and environmental design criteria should to be implemented to meet the conservation and visual landscapes objectives of the surrounding area. These include:

Tree and shrub plantings for screening and reclamation should use native species and conform to the surrounding landscape; Plant vegetation for screening before construction to allow vegetation to establish and minimize visual impacts during development; Minimize visual impacts from the tailings pond and stockpiles by possibly creating smaller ponds that can be progressively reclaimed and use vegetation for screening around the perimeters; Design buildings in colours, styles and size similar to those in the area and use earth sheltered buildings if the building will exceed the size of a typical farm shed; Use earth berms and stonewalls to screen development and maintain the character of the surrounding landscape; Haul roads should be designed to follow existing field boundaries and landscape contours; Retain existing vegetation in accordance with recommendation in BS5837:2005, Trees in Relation to Construction.

Following good industry practice, it is assumed a social, environmental, health and safety management system will be implemented to meet the FRPSDQ\¶VFRPPLWPHQWVWRSURWHFWWKH environment and the health and safety of the workers and the surrounding communities. The management system and plans should be designed to monitor and maintain permit compliance and a social licence to operate.

1.10.4 Permits

For exploration work, formal notice of intention to enter land to carry out work must be given, and the agreement of landowners sought, before entering any property. Compensation is generally payable to the landowner for any damage caused during exploration.

Project development is subject to legislative requirements from the European Union, England (UK) and Northern Ireland. Applicable environmental legislation is listed in Table 1.10. In addition, there are a number of international conventions that will apply to the project and need to be considered in further planning.

Table 1.10 Key Applicable Environmental Legislation

Name of Legislation Jurisdiction Date Adopted Assessment of Effects of Certain Public and Private Projects on the European Environment (EIA Directive) (Directive 97/11/EC of 3 March 1997 amending Union 1997 Directive 85/337/EEC) The Planning (Environmental Impact Assessment) Regulations 2012 SR 2012 Northern 2012 No 59 Ireland Environmental Protection Act United 1990 Kingdom Climate Change Act United 2008 Kingdom Directive 2006/21/EC Management of Waste from Extractive Industries and European 2006 Amending Directive 2004/35/EC Union Waste Directive (Directive 2008/98/EC) European 2008 Union Ambient Air Quality and Cleaner Air for Europe (Directive 2008/50/EC) European 2008 Union Industrial Emissions (Integrated Pollution Prevention and Control) (Directive European 2010 2010/75/EU) Union Assessment and Management of Environmental Noise (Directive European 2002 2002/49/EC) Union Habitats Directive European 1992 Union Birds Directive (Directive 2009/147/EC) European 2009 Union

At this time it is uncertain how long it will take to permit the project. Land acquisition will also need to be negotiated once decisions are made on the final location of infrastructure and facilities.

1.10.5 Closure

The objective of the mine closure plan is to remove and close down activities in a manner that ensures public safety and to reclaim the land to a usable state consistent with the surrounding land use objectives. In the case of the Curraghinalt project, the land will be restored to productive use for farming and/or heathlands. The visual landscape objective is to minimize disruption of the outstanding natural beauty of the area during operations and restore this value at closure.

Closure will consist of plugging and securing underground openings, removing to licensed facilities any hazardous and contaminated materials from site, decommissioning and demolition of facilities and buildings, and reclaiming waste rock and tailings storage facilities. It has been assumed that the electrical substation is an infrastructural asset that will be left in place post-closure, owned by the utility.

When possible, equipment and machinery will be sold or recycled. Buildings will be demolished and reclaimed, recycled, or disposed of. Concrete foundations will be broken up and removed from site, reclaimed if possible, or buried at a certified waste disposal facility.

The tailings facility seepage and surface waters will be monitored. Seepages will be collected and pumped back to the pond if the quality is not acceptable for release. Tailings pond water will be monitored and pumped back to the plant for treatment if necessary until the pond water is acceptable for discharge to the environment. A spillway will then be constructed in the dam and the dams and beach areas capped with overburden and revegetated. Similarly, any non-acid generating waste rock stored on surface will be capped with overburden and revegetated. It is assumed that any potentially acid generating (PAG) waste rock will be used in backfill and flooded after mine closure to minimize oxidation

Reclamation costs are estimated at $7.5 million in Year 16. Using a real discount rate of 2%, it is assumed for the PEA that the present value (approximately 71%) of this amount will be required to be posted as a financial guarantee at the start of the project. A financial guarantee is required to meet requirements of Directive 2006/21/EC, Management of Waste from Extractive Industries.

1.11 CAPITAL AND OPERATING COST ESTIMATES

1.11.1 Capital Costs

Capital expenditures and capitalized development costs for the base case are summarized as initial and sustaining costs in Table 1.11. The estimates are expressed in second quarter 2012 Canadian dollars, without escalation unless otherwise noted. The expected accuracy of the estimates is ± 30%.

Table 1.11: Capital Cost Summary

A rea Initial Capital Sustaining Capital Cost ($ 000) Cost ($ 000) Capitalized Development 15,745 64,925 Mining Equipment 14,770 33,058 Processing 50,743 - Infrastructure 49,833 16,301 Indirect Costs 12,927 - 2ZQHU¶V&RVWV 17,386 - Contingency 38,341 - Total 199,745 114,284

Expressed as US dollars, initial and sustaining capital amounts to US$192.1 million and US$109.9 million, respectively.

1.11.2 Operating Costs

Estimated cash operating costs over the life of the project are summarized in Table 1.12.

Table 1.12 Summary of Life-of-Mine Operating Costs

Life-of-mine Cost Unit Cost A rea ($ 000) $/t ore treated Production Drilling and Blasting 44,600 4.73 Sill 399,183 42.36 Sill (through waste) 13,764 1.46 Diesel Fuel 22,869 2.43 Backfill 29,280 3.11 Manpower 99,948 10.61 Ventilation & Dewatering 2,279 0.24 Exploration 911 0.10 Major Equipment Maintenance 85,618 9.08 Miscellaneous and Sundries 7,395 0.78 Mining G&A 43,653 4.63 Stockpile rehandle 310 0.03 Sub-total Mining 749,812 79.56 Labour - Metallurgy 5,180 0.53 - Laboratory 6,963 0.71 - Production 27,417 2.81 - Maintenance 14,037 1.44 Power 35,561 3.65 Maintenance 18,346 1.88 Crushing 759 0.08 Grinding 33,760 3.46 Reagents 34,835 3.57 Miscellaneous 13,376 1.37 Sub-total Processing 190,235 20.19 Labour 21,903 2.25 Equipment Maintenance 6,646 0.68 Mobile Equipment Operation 1,472 0.15 Environmental & Social 7,594 0.78 G&A (other) 43,770 4.49 Sub-total General and Administrative 81,385 8.64

Total Operating Costs 1,021,431 108.38

1.12 ECONOMIC ANALYSIS AND SENSITIVITY STUDIES

1.12.1 Basis of Evaluation

accepted within the mining industry as representing the economic value of a project after allowing for the cost of capital invested.

The objective of the study was to determine the viability of the proposed mine and process plant to exploit the Curraghinalt deposit. In order to do this, the cash flow arising from the base case has been forecast, enabling a computation of the NPV to be made. The sensitivity of this NPV to changes in the base case assumptions is then examined.

1.12.2 Macro-economic Assumptions

Unless otherwise stated, all results are expressed in Canadian dollars. Cost estimates and other inputs to the cash flow model for the project have been prepared using constant, second quarter 2012 money terms, i.e., without provision for escalation or inflation. Trailing 36- month average exchange rates of CDN$1.04/US$ and CDN$1.62/GBP are applied in the base case.

Applying the Capital Asset Pricing Model (CAPM) gives an estimated cost of equity for the Curraghinalt project of between 5% and 11%, as shown in Table 1.13. Micon has taken a figure of 8% (i.e., in the middle of this range) as its base case, and provides the results at alternative rates of discount for comparative purposes.

Table 1.13 Estimated Cost of Equity

Range Lower M iddle Upper Risk Free Rate (%) 1.5 2.0 2.5 Market Premium for equity (%) 5.0 5.0 5.0 Beta 0.7 1.2 1.7 Cost of equity (%) 5.0 8.0 11.0

Figure 1.5 shows the monthly average gold and silver prices over the past six years, together with the 3-year trailing averages. At the end of June, 2012, the three-year trailing averages for each metal were $1378/oz gold and $26.28/oz silver, and these metal prices were selected for the base case. These prices were applied consistently throughout the operating period.

Silver contributes approximately 0.6% of the projected total revenue for the base case, so the impact of changing the silver price forecast is minimal.

Figure 1.5 Monthly Average Gold and Silver Prices since July 2006 (source: Kitco.com)

For comparison, Micon also evaluated the sensitivity of the project to using recent (1 month), and 1, 2, 3, 5 and 10-year price averages. The prices used in each of these cases are shown in Table 1.14. As part of its sensitivity analysis, Micon also tested a range of prices 30% above and below base case values.

Table 1.14 Metal Price Averages

Item Units 1-month 1-year 2-year 3-year avg. 5-year 10-year Jun-2012 average average Base Case average average Gold US$/oz 1,597 1,672 1,521 1,378 1,166 814 Silver US$/oz 28.05 33.16 30.98 26.28 21.44 14.66

United Kingdom corporation tax payable on the project has been forecast using current rates, being 20% on the first GBP 300,000 per year, 22% on the increment up to GBP 1.50 million, and 24% on the balance. Depreciation allowances of 18% are assumed to be taken annually on all project capital, on a declining balance basis.

A royalty of 6% of NSR value has been provided for in the cash flow model.

Refining charges are estimated at US$6.00/oz gold and US$0.50/oz silver, based on similar projects. Doré transport charges of US$5,000 per shipment and cash-in-transit insurance, calculated as a percentage of shipment value, are also provided.

1.12.3 Technical Assumptions

The technical parameters, production forecasts and estimates described elsewhere in this report are reflected in the base case cash flow model. These inputs to the model are summarised below. The measures used in the study are metric except where, by convention, gold and silver content, production and sales are stated in troy ounces.

1.12.3.1 Mine Production Schedule

Figure 1.6 shows the annual tonnage of development and stoping material mined, as well as the waste rock tonnage, all of which are held reasonably constant over the LOM period.

Figure 1.6 Annual Mining Schedule

In Figure 1.7, the head grades for gold and silver in the mill feed are shown to remain within very narrow ranges over the LOM period. Tonnages milled reflect the drawdown of stockpiled material during the mill startup and again at the end of the LOM period.

Figure 1.7 Annual Processing Schedule

As a consequence of steady tonnage, grade and recovery from mill feed, annual production of gold and silver remain steady over the LOM period (Figure 1.8). This chart also shows the annual total cash cost per ounce of gold sales. Total cash costs, including refining charges and royalties and net of silver credits, average $532/oz gold over the LOM period. In Years 5 and 7, costs fall below this level owing to a reduction in development costs in those periods.

Figure 1.8 Annual Production Schedule

1.12.3.2 Operating Costs

Direct operating costs average $108.38/t milled over the LOM period, comprised of $79.56/t mining, $20.19/t processing, and $8.64/t general and administrative costs. Figure 1.9 shows these expenditures over the LOM period, compared to the net sales revenue, showing the strong margin maintained over the LOM period.

Figure 1.9 Direct Operating Costs

1.12.3.3 Capital Costs

Pre-production capital expenditures are estimated to total $199.75 million, including $30.5 million for mining equipment and pre-production development, $50.7 million processing, $49.8 million infrastructure, $12.9 million indirect costs, $12.0 million in RZQHU¶V FRVWV D Pillion provision for closure and rehabilitation, and contingencies totalling $38.3 million.

Working capital has been estimated to include 15 days product inventory in the milling/leach circuit, and 15 days receivables from despatch of doré. Stores provision is for 60 days of consumables and spares inventory, less 30 days accounts payable. An average of $22.3 million of working capital is required over the LOM period.

Figure 1.10 compares annual capital expenditures over the preproduction and LOM periods ZLWKWKHSURMHFW¶VFDVKRSHUDWLQJPDUJLQ

Figure 1.10 Capital Expenditures

1.12.3.4 Base Case Cash Flow

The LOM base case project cash flow is presented in Table 1.15 and Figure 1.11.

Table 1.15 Life-of-Mine Cash Flow Summary

L O M total C DN$/t US$/oz ($ 000) M illed Gold Revenue 3,186,233 338.08 1378.03

Mining costs 749,812 79.56 324.29 Processing costs 190,235 20.19 82.28 General & Administrative costs 81,385 8.64 35.20 S/T Direct site operating costs 1,021,431 108.38 441.76 Silver credit (19,044) (2.02) (8.24) Refining & transport charges 37,575 3.99 16.25 S/T Cash operating costs 1,039,962 110.35 449.77 Royalty 190,062 20.17 82.20 Total Cash Costs 1,230,024 130.51 531.98 E BI T D A 1,956,209 207.57 846.05 Capital expenditure 314,029 33.32 135.82 Net cash flow (before tax) 1,642,180 174.25 710.23 Taxation 402,895 42.75 174.25 Net cash flow (after tax) 1,239,286 131.50 535.98

Figure 1.11 Life-of-Mine Cash Flows

The project demonstrates an undiscounted pay back of 2.1 years, or approximately 2.4 years when discounted at 8.0%, leaving a production tail of just over 12 years. Annual cash flows are presented in Table 1.16 (over).

Table 1.16 Base Case Life of Mine Annual Cash Flow

Production Schedule LOM Yr-2 Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12 Yr13 Yr14 Yr15 Yr16 Yr17 Underground Mine Production TOTAL Development 000 t 1,890 - 74 79 95 138 166 89 167 80 143 128 122 107 133 112 134 123 - - Stoping 000 t 7,534 - 236 387 526 483 455 531 453 541 478 492 498 514 488 508 487 459 - - Resources mined 000 t 9,425 - 310 465 621 621 621 621 621 621 621 621 621 621 621 621 621 582 - - U/G Waste Rock mined 000 t 1,260 - 70 47 43 64 74 42 83 43 80 86 79 70 98 83 110 189 - -

Processing Plant Throughput 000 t 9,425 - - 621 621 621 621 621 621 621 621 621 621 621 621 621 621 621 117 - Gold grade g/t 8.06 - - 8.43 8.36 8.04 7.93 8.47 8.03 8.00 7.79 8.01 8.09 8.23 7.98 7.87 7.92 7.68 8.06 - Silver grade g/t 3.31 - - 3.62 3.50 3.28 3.20 3.45 3.29 3.24 3.15 3.19 3.24 3.26 3.28 3.33 3.42 3.24 3.52 -

Payable Metal (imperial) Gold oz 2,223,236 - - 153,180 151,849 146,124 144,149 153,943 145,927 145,353 141,575 145,451 146,980 149,522 145,007 142,965 143,993 139,603 27,613 - Silver oz 715,655 - - 51,406 49,788 46,556 45,509 49,011 46,767 46,084 44,811 45,403 46,009 46,297 46,576 47,305 48,662 46,041 9,429 -

Exchange Rate CAD/USD 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 1.04 Gold price US$/oz 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 1,378.03 Silver price US$/oz 26.280 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28 26.28

Cash Flow Forecast LOM TOTAL Yr-2 Yr-1 Yr1 Yr2 Yr3 Yr4 Yr5 Yr6 Yr7 Yr8 Yr9 Yr10 Yr11 Yr12 Yr13 Yr14 Yr15 Yr16 Yr17 CAD/t ore US$/oz CAD 000 Gross Revenue (Gold) 338.08 1,378.03 3,186,233 - - 219,530 217,622 209,418 206,588 220,624 209,135 208,313 202,899 208,454 210,645 214,288 207,817 204,890 206,364 200,073 39,574 -

Total Cash Costs 130.51 531.98 1,230,024 - 23,919 67,967 74,818 80,450 87,040 71,608 86,521 69,773 84,514 83,429 79,431 78,828 83,055 79,788 86,873 85,946 6,066 - Mining 79.56 324.29 749,812 - 23,919 35,762 42,708 48,847 55,612 39,267 54,943 38,235 53,329 51,864 47,725 46,872 51,564 48,522 55,534 54,992 117 - Processing 20.19 82.28 190,235 - - 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 12,525 2,362 - G&A 8.64 35.20 81,385 - - 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 5,358 1,010 - Sub-total Direct Costs 108.38 441.76 1,021,431 - 23,919 53,645 60,592 66,731 73,495 57,150 72,826 56,118 71,212 69,747 65,608 64,755 69,447 66,406 73,418 72,875 3,489 - By product credit (Silver, net) (2.02 ) (8.24 ) (19,044) - - (1,368) (1,325) (1,239) (1,211) (1,304) (1,245) (1,226) (1,192) (1,208) (1,224) (1,232) (1,239) (1,259) (1,295) (1,225) (251) - Transport & Refining charges (Gold) 3.99 16.25 37,575 - - 2,592 2,568 2,467 2,434 2,603 2,464 2,456 2,393 2,458 2,484 2,525 2,451 2,417 2,437 2,360 467 - Royalties 20.17 82.20 190,062 - - 13,098 12,983 12,491 12,322 13,160 12,475 12,425 12,102 12,432 12,563 12,780 12,396 12,224 12,313 11,936 2,361 -

Operating Margin 207.57 846.05 1,956,209 - (23,919) 151,563 142,805 128,968 119,548 149,016 122,615 138,540 118,385 125,025 131,214 135,460 124,762 125,102 119,491 114,127 33,508 -

Capital Costs 33.32 135.82 314,029 42,831 156,914 2,971 2,601 7,689 4,814 12,328 10,229 3,051 5,615 9,645 17,203 4,665 11,201 5,286 6,719 10,267 - - Mine Development 8.56 34.89 80,670 - 15,745 1,896 1,877 3,154 3,940 2,301 4,549 2,406 4,656 4,490 4,657 4,206 5,902 4,826 6,259 9,807 - - Mining Equip. 5.07 20.69 47,827 - 14,770 1,075 725 460 874 10,027 1,605 645 959 1,080 12,546 460 1,224 460 460 460 - - Processing Capital 5.38 21.95 50,743 15,223 35,520 ------Infrastructure 7.02 28.60 66,134 10,430 39,403 - - 4,075 - - 4,075 - - 4,075 - - 4,075 - - - - - Indirect Capital 7.28 29.69 68,654 17,178 51,476 ------

Change in Working Cap - - - - - 22,153 323 (159) 330 (208) 367 (1,457) 815 330 (158) 224 (126) (494) 709 (513) (3,400) (18,735)

Pre-tax c/flow 174.25 710.23 1,642,181 (42,831) (180,833) 126,439 139,880 121,438 114,403 136,896 112,018 136,946 111,955 115,050 114,170 130,570 113,688 120,311 112,063 104,374 36,907 18,735 Tax payable 42.75 174.25 402,895 - - 21,865 26,955 24,695 23,267 30,975 24,958 29,338 24,962 26,844 28,309 29,432 27,033 27,299 26,146 24,864 5,954 - C/flow after tax 131.50 535.98 1,239,286 (42,831) (180,833) 104,574 112,925 96,742 91,136 105,921 87,060 107,609 86,993 88,206 85,861 101,139 86,655 93,013 85,917 79,509 30,953 18,735 Cumulative C/Flow (42,831) (223,663) (119,089) (6,164) 90,579 181,715 287,636 374,697 482,305 569,298 657,504 743,365 844,504 931,159 1,024,171 1,110,088 1,189,598 1,220,551 1,239,286 Discounted C/Flow (8%) 485,330 (36,721) (143,551) 76,865 76,855 60,964 53,177 57,226 43,552 49,844 37,310 35,028 31,571 34,434 27,317 27,149 23,221 19,897 7,172 4,020 Cumulative DCF (36,721) (180,271) (103,406) (26,551) 34,413 87,590 144,816 188,368 238,212 275,521 310,549 342,120 376,554 403,871 431,021 454,241 474,139 481,311 485,330 Max funding reqmt to positive cashflow (270,652) (42,831) (223,663) (270,652) (148,969) (38,389) ------

Over the LOM period, gross revenues for gold and silver totalling $3,205.8 million are distributed as shown in Figure 1.12. This diagram illustrates the robust nature of the net cash flows forecast to be generated by the project, both on a before- and after-tax basis.

Figure 1.12 Distribution of Gold and Silver Revenues

1.12.3.5 Base Case Evaluation

The base case evaluates to an IRR of 51.7% before taxes and 41.9% after tax.

At a discount rate of 8.0%, the net present value (NPV8) of the cash flow is $664.6 million (US$ 639 million) before tax and $485.3 million (US$ 467 million) after tax. Table 1.17 presents the results in Canadian dollars at comparative annual discount rates of 5%, 8% and 11%. Stated in US dollars, at annual discount rates of 5% and 11%, the after-tax NPV of the project is US$ 655 million and US$336 million, respectively.

Table 1.17 Base Case Cash Flow Evaluation

Base Case $ 000 L O M Discounted Discounted Discounted IRR Total at 5%/y at 8%/y at 11%/y (%) G ross Revenue (Gold) 3,186,233 1,903,535 1,442,827 1,117,046 Mining costs 749,812 448,187 340,171 263,940 Processing costs 190,235 113,237 85,651 66,180 General & Administrative costs 81,385 48,444 36,642 28,312 Transport & Refining, net of Ag credit 18,531 11,074 8,392 6,494 Royalty 190,062 113,548 86,066 66,633 Total Cash Costs 1,230,024 734,490 556,923 431,559 E BI T D A 1,956,209 1,169,046 885,904 685,487 Capital expenditure 314,029 239,989 209,694 185,946 Working Capital - 9,887 11,562 11,893 Net cash flow (before tax) 1,642,181 919,169 664,648 487,648 51.7 Taxation 402,895 238,160 179,317 137,895 Net cash flow (after tax) 1,239,286 681,009 485,330 349,753 41.9

This preliminary economic assessment is preliminary in nature; it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized.

1.12.4 Sensitivity Study

1.12.4.1 Capital, Operating Costs and Revenue Sensitivity

The sensitivity of project returns to changes in all revenue factors (including grades, recoveries, prices and exchange rate assumptions) together with capital and operating costs was tested over a range of 30% above and below base case values. The results show that the project is most sensitive to revenue factors, with an adverse change of 30% reducing after-tax NPV8 from $485 million to $161 million. The impact of changing operating costs is lower, with a 30% adverse change reducing NPV8 to $379 million. The project is least sensitive to capital costs, with a 30% increase in capital reducing NPV8 to $422 million. Thus, NPV8 remains positive within the expected range of accuracy of the estimates. ,Q 0LFRQ¶V DQDO\VLV DSSO\LQJ DQ LQFUHDVH RI PRUH WKDQ LQ ERWK FDSLWDO DQG RSHUDWLQJ costs simultaneously would be required to reduce NPV8 to near zero, further demonstrating the robust nature of the project economics. Figure 1.13 shows the results of changes in each factor separately.

Figure 1.13 Sensitivity Diagram

1.12.4.2 Metal Price Sensitivity

The sensitivity of the project to variation in gold price was tested using 1 month, and, 1, 2, 3, 5 and 10-year trailing averages applied over the life-of-mine period, as shown in Figure 1.14 and in Table 1.18. At US$1,200/oz gold, the project has an IRR of 39.1% pre-tax and 31.8% after tax.

Figure 1.14 Sensitivity to Metal Prices

Table 1.18 Sensitivity to Metal Prices

Pricing Period 1-month 1-year 2-year 3-year 5-year 10-year average average average average average average Pre-tax IRR % 60.1 63.1 57.1 51.7 41.1 24.9 After-tax IRR % 48.6 51.0 46.2 41.9 33.4 20.2

Pre-tax NPV 8 C DN 000 818,095 874,215 762,681 664,648 478,875 218,744

After-tax NPV 8 601,704 644,264 559,678 485,330 344,441 147,160

Pre-tax NPV8 USD 000 786,630 840,591 733,348 639,084 460,456 210,330

After-tax NPV8 578,561 619,485 538,152 466,664 331,194 141,500

1.12.4.3 Sensitivity to Process Flowsheet

Micon conducted a trade-off study to determine the optimum process flowsheet for the Curraghinalt deposit. The four flowsheets considered were:

A) Crushing/milling followed by CIL and gold recovery. B) Crushing/milling/gravity concentration followed by flotation of a concentrate for CIL/gold recovery. C) Crushing/milling/gravity concentration followed by flotation of a concentrate for sale. D) Crushing/milling followed by flotation of (i) a copper concentrate for sale and (ii) a bulk sulphide concentrate for CIL/gold recovery.

A comparison of the cash flows for each of these options suggests that Option A, Milling/CIL provides the best overall economic return, and so this was selected as the base case for this study. Figure 1.15 shows the NPV and IRR for each of the four options.

Figure 1.15 Sensitivity to Process Flowsheet

1.13 CONCLUSIONS

With further exploration there is potential for the conversion of mineral resources from the Inferred and Indicated categories to the Indicated and Measured categories at Curraghinalt. Within the present resource estimate, only the major vein systems have mineral resources classified as Measured or Indicated. While some close-spaced drilling has been undertaken on the property, more is required DQG LW LV 0LFRQ¶V RSLQLRQ WKDW WKHUH DUH PDQ\ VXE-veins which have good potential to be brought into the Indicated resource category.

The mineral resource has been estimated largely for the central part of the Curraghinalt deposit. There are sufficient indications, based on preliminary mapping and geochemical sampling, for exploration potential in either direction from the deposit and there is good potential that some of the veins identified in the present study may contribute further to future resource estimates. Deep drilling has indicated significant potential down dip. The regional geochemical, mapping and sampling work clearly indicates that there are parallel trends of mineralization in conformity with the Curraghinalt trend in the Dalradian metasediments.

The Curraghinalt gold deposit may be regarded as a mid-stage exploration project on which a reasonably-sized gold mineral resource of good grade has been estimated. Additionally there are other good exploration opportunities on the four licence areas for gold and base metal mineralization.

Dalradian Resources has designed a program of exploration and preliminary engineering for the Curraghinalt gold deposit as well as the remainder of the mineral licence areas at the Tyrone project. The program is summarized in Table 1.2. Details for the exploration and development program may be found in Section 26 of this report.

Mining of the Curraghinalt deposit will be undertaken using a Longitudinal Sublevel Retreat Underground Mining method with both paste- and waste rock backfill to maximize the extraction of the high value resources at the Curraghinalt Project, reduce the surface footprint required for tailings disposal and allow savings in the volume and transportation cost of waste rock from underground development. Stopes will have a minimum width of 1.80 m and development of sills along the veins will be restricted to 3.0 m width to minimize dilution. The sills will be developed in both directions along the strike length of the mineralized zone, providing two production faces for each vein on each level. Levels will be mined at 20 m intervals. Haulage ramps will have dimensions 4.5 x 4.5 m to accommodate 30-t trucks.

The PEA considers four possible process flowsheets for the extraction of gold from the Curraghinalt resource. Of these, the whole-ore leach (Option A) was considered the most effective and has been used as the base case for project evaluation. Nevertheless, Micon considers additional testwork, including optimizing grind, reagent strengths and retention times are required, and pilot studies are required before flowsheet development and design specifications can be finalized.

Micon considers the relatively low grade of the copper concentrate produced in Option D will present a challenge with regard to shipping and marketing. However, testwork is ongoing that may validate producing a more valuable, higher grade, copper concentrate, based on a lower weight recovery. Accordingly, Micon is of the opinion that during further development of the project this option should continue be considered along with the base case flowsheet A. It is of critical importance that representative samples of the copper mineralization are used for future testwork. In developing this option, Dalradian Resources will need to confirm off-take terms with a smelter, and also investigate any deleterious content of the copper concentrate.

Micon also suggests that alternative flotation reagents should be investigated in future copper flotation testwork, since the vendor of the collector used in the G&T copper flotation testwork is not accepting fresh orders for this product until their new facility opens in 2014.

Infrastructure required for the project has been identified and is provided for in the evaluation. A site specific layout has not been developed, though, pending discussion with affected landowners over the siting of these works.

The provision of electricity to the mine will require a new overhead powerline, probably from Strabane, approximately 27 km from the mine. A two-year permitting and construction period is expected following a decision to develop the mine.

The PEA assumes that collection of rainfall in the tailings storage facility will provide process make-up water, but further work is required to properly assess water supply sources available to the project.

Future development of the project will need to take into account its location within an Area of Outstanding Natural Beauty, and proximity to Areas of Special Scientific Interest and Special Areas of Conservation.

A small proportion of the waste rock samples tested were acid-generating, so it will be necessary for such material to be segregated and stowed as underground backfill. On closure of the mine, the entrance will be plugged and flooding of the workings will then minimize oxidation.

Based on its economic evaluation of the base case and sensitivity studies, Micon concludes that this PEA demonstrates the viability of the project as proposed, and that further development is warranted.

1.14 RECOMMENDATIONS

0LFRQ KDV UHYLHZHG 'DOUDGLDQ 5HVRXUFHV¶ SURSRVHG H[SORUDWLRQ and development program (Table 1.19) and finds it to be reasonable and warranted in light of the data presented and observations made in this report. Micon recommends that Dalradian Resources proceed with the program.

Table 1.19 Exploration and Development Program Summary Budget

Cost Activity Metres ($ millions) Exploration Resource Drilling and Update 16,300 6.59

In its January, 2012 report on the property, Snowden made recommendations regarding the collection and analysis of geotechnical data. Micon concurs with those recommendations. Also, Micon recommends that during geotechnical data collection the rock mass should be accurately characterized and described. This allows the rock mass attributes, parameters and ratings to be assigned at a later stage enabling the classification to be made in any of the proposed rock mass classification systems (e.g., RMR, Q-system or MRMR).

Regarding development of the mine design, Micon recommends that:

Further optimization of the mining method be performed to evaluate potential additional economic benefits to the project by considering and combining a higher selectivity, non-mechanized mining method with the proposed Longitudinal Sublevel Retreat.

Further studies be performed on the backfill system for the Project and that a systematic laboratory test program be performed on the mine tailings to determine the optimum cement addition and cement type to be used as the binding agent for the paste backfill.

Non-PAG waste rock produced during underground exploration and mine development activities may be utilized as bulk fill materials to meet the construction needs, where appropriate. The structural fill material will be imported from local suppliers, as required.

Future metallurgical testwork should be considered upon review of the final testwork report currently being prepared by G&T.

Based on a review of reports leading up to the current testwork underway, it is recommended that future testwork should include the following:

Pre-aeration and lead nitrate addition should be included in the future leach tests to determine the effect on leach residence time and cyanide consumption Tests for graphitic carbon in the ore should be investigated Additional Bond ball mill work index tests should be conducted Semi-Autogenous Grinding (SAG) amenability testwork should be conducted The effect of grind size on leach extractions should be evaluated further. Alternative flotation reagents should be evaluated, especially the collector 3418A. Carbon in Leach (CIL) testwork should be performed Testwork with higher cyanide concentrations should be conducted Confirmatory testwork on a variety of representative feed samples once a preferred flowsheet is established

Golder recommended three sites for further investigation as potential tailings storage facilities. Although these sites were selected on the basis of conventional (slurry) tailings disposal, Micon suggests that paste or dry stack tailings be considered to minimize the footprint and increase the available options for siting the tailings and employ progressive reclamation to minimize visual impacts.

Potential sources of process water should be investigated, including a study of groundwater in wells close to the proposed underground mine, as recommended by SLR, as well as further study of the proposed collection of rainfall at the tailings storage facility.

In this PEA, the estimated unit cost for electrical power is based on best available information. As the project develops, Dalradian Resources should obtain project-specific pricing.

Negotiations with landowners should take place as the project moves forward so that a site- specific layout of the plant and infrastructure can be developed during the next stages of project engineering.

In further development of the project, Micon recommends that Dalradian Resources obtain offers of specific terms for the purchase of doré, and for the copper concentrate envisioned in process Option D.

2.0 IN T R O DU C T I O N

At the request of Dalradian Resources Inc. (Dalradian Resources), Micon International Limited (Micon) has prepared a Preliminary Economic Assessment (PEA) of the Curraghinalt gold deposit, which is contained ZLWKLQWKDWFRPSDQ\¶V7\URQHSURMHFWLQ&RXQW\7\URQHDQG County Londonderry, Northern Ireland. This Technical Report, prepared in accordance with the reporting standards and definitions required under Canadian National Instrument (NI) 43- 101, summarizes the results of that study. This PEA is based upon an estimate of the Curraghinalt mineral resources originally prepared by Micon in November, 2011.

The scope of the proposed project includes the development of an underground mine to extract the mineralized vein material that comprises the resource, and to treat that material in a conventional crushing, milling and carbon-in-leach plant to extract the gold and silver content, and to dispose of the leached residue underground using paste backfill to the extent possible, with excess material stored in a purpose-built tailings storage facility near the mine site. At the proposed scale of operation, the mine has a projected life of about 15 years.

At least four previous estimates of mineral resources are known to have been completed on the Curraghinalt deposit. The most recent, SULRUWR0LFRQ¶VLQYROYHPHQWZLWKWKHSURSHUW\, was prepared by John V. Tully & Associates Inc. (Tully) on behalf of Tournigan Gold Corporation (now Tournigan Energy Ltd., Tournigan). The independent Technical Report for this estimate is dated January 27, 2005 and was filed on SEDAR on March 24, 2005 (Tully, 6LQFH WKH FRPSOHWLRQ RI 7XOO\¶V PLQHUDO UHVRXUFH HVWLPDWH 7RXUQLJDQ XQGHUWRRN further exploration work on the property, concentrating on the Curraghinalt deposit.

Micon prepared a second estimate of mineral resources on the Curraghinalt deposit for Tournigan in 2007. This took into account the previous exploration work and 20 additional holes completed since the beginning of 2005. This estimate is described in a Technical report dated November, 2007 (Mukhopadhyay, 2007) which was filed on SEDAR by Tournigan on January 17, 2008. After the database for that estimate was frozen, Tournigan completed four additional deep holes on the deposit to test the down plunge extent of those existing resources and then ceased work on the project. After 2007, a small amount of prospecting and sampling was completed on the rest of the property in order to hold the licences.

Dalradian Resources acquired the Tyrone project in the fall of 2009 and retained Micon to update the 2007 resource estimate with the final four holes as part of a Technical Report prepared to support the initial listing of the company on the Toronto Stock Exchange (TSX).

Subsequent to Dalradian 5HVRXUFHV¶ acquisition of the Tyrone project, a diamond drill campaign was started. From the start of the drilling to the resource cut-off date, an additional 20,561.9 m were drilled in 45 surface diamond drill holes. The mineral resource estimate used in this PEA was updated in November, 2011 to incorporate this information.

7RXUQLJDQ¶V DQG WKHUHIRUH 7XOO\¶V DQG 0LFRQ¶V SUHYLRXV ZRUN FRQFHQWUDWHG RQ WKH Curraghinalt gold deposit located on concession DG1. There also exists gold and

volcanogenic massive sulphide exploration potential on the south side of the Omagh Thrust Fault as well as the potential for along strike continuation of the Curraghinalt deposit mineralization. That potential will also be described herein.

The geological setting of the property, mineralization style and occurrences, and exploration history were described in reports that were prepared by Peatfield (2003), Tully (2005), Williams (2007), Coller (2007) and in various government and other publications listed in the references of this report. The relevant sections of those reports are reproduced herein.

The Tyrone project was visited by B. Terrence Hennessey, P.Geo., Vice President of Micon DQGEDVHGLQWKHILUP¶VKHDGRIILFHLQ7RURQWR2QWDULR Canada, on November 6 and 7, 2009. Discussions were held with representatives of Aurum Exploration Services Ltd. (Aurum) and Mr. Patrick Anderson, a director of Dalradian Resources. Aurum provided local project management services on behalf of Tournigan and provided the same to Dalradian Resources. Mr. Hennessey conducted a second site visit to the Curraghinalt project from October 3 to 5, 2010.

For the PEA, authors B. Foo, C. Jacobs, and A. Villeneuve visited the property on April 18- 19, 2012. In addition to an inspection of the underground workings, the team visited the surface of the project site and potential tailings storage sites in the district, Dalradian Resources offices in Gortin, and core storage/logging facility and offices in Omagh.

Messrs. Hennessey)RR'DPMDQRYLü9LOOHQHXYHDQG-DFREV are all Qualified Persons (QP) as defined in NI 43-101.

The mineral resource estimate is based on exploration results and interpretation current as of October, 2011. The PEA has an effective date of July 25, 2012.

3.0 R E L I A N C E O N O T H E R E XPE R TS

Micon has reviewed and analyzed exploration data provided by Dalradian Resources, its consultants and previous operators of the property, and has drawn its own conclusions therefrom, augmented by its direct field examination. Micon has not carried out any independent exploration work, drilled any holes or carried out any significant program of sampling and assaying. However, precious metal-bearing veins are visible in an adit and drifts developed underground and have also been found on surface in a nearby gravel pit.

While no large scale mining of precious metals has taken place in the immediate area there is a small producing gold deposit nearby known as Cavanacaw and held by Galantas Gold Corporation (Galantas) which substantiates the local presence of gold mineralization (see Section 16). Samples collected by Micon come from core or mineralized exposures and independently confirm the presence of gold at grades similar to those claimed by others.

While exercising all reasonable diligence in checking, confirming and testing it, Micon has relied upon the data presented by Dalradian Resources, and any previous operators of the project, in formulating its opinion.

The various agreements under which Dalradian Resources holds title to the mineral lands for this project have not been thoroughly investigated or confirmed by Micon and Micon offers no opinion as to the validity of the mineral title claimed. The descriptions were provided by Dalradian Resources as received from DETI and CEC, the entities responsible for licensing mineral exploration in Northern Ireland (see Section 4.2).

The description of the property is presented here for general information purposes only, as required by NI 43-101. Micon is not qualified to provide professional opinion on issues related to mining and exploration title or land tenure, royalties, permitting and legal and environmental matters. Accordingly, the authors have relied upon the representations of the issuer, Dalradian Resources, for Section 4 of this report, and have not verified the information presented therein.

Micon has relied on the expertise of SLR in its reporting of social, environmental and permitting issues.

7KHFRQFOXVLRQVDQGUHFRPPHQGDWLRQVLQWKLVUHSRUWUHIOHFWWKHDXWKRUV¶EHVWMXGJPHQWLQOLJKW of the information available at the time of writing. Micon reserves the right, but will not be obliged, to revise this report and conclusions if additional information becomes known to it subsequent to the date of this report. Use of this report acknowledges acceptance of the foregoing conditions.

Those portions of the report that relate to the location, property description, infrastructure, history, deposit types, exploration, drilling, sampling and assaying (Sections 4 to 11) are taken, at least in part, from previous Technical Reports prepared by Micon and others as well as updated information provided by Dalradian Resources.

4.0 PR OPE R T Y D ESC RIPT I O N A ND L O C A T I O N

4.1 LOCATION

The Tyrone project is located in County Tyrone and County Londonderry, Northern Ireland. The Curraghinalt gold deposit is found on the property, approximately 75 km west of Belfast and 15 km northeast of the town of Omagh. The property is accessible by paved road, a distance of approximately 127 km from Belfast. See Figure 4.1.

Figure 4.1 General Location Map

Source: Dalradian Resources, 2011

4.2 DESCRIPTION

The Tyrone project is the subject of four Mineral Prospecting Licences and four option agreements. The four Mineral Prospecting Licences are issued by the Department of Enterprise, Trade and Investment (DETI) and the four Option Agreements are issued by the Crown Estate Commissioners (CEC). The Mineral Development Act (Northern Ireland) 1969 makes DETI responsible for the grant of prospecting licences and mining licences for exploration and development of minerals. The CEC is responsible for the licensing of precious metal exploration. Before 2011, the CEC issued Mineral Prospecting Licences similar to DETI. However, in 2011 this was changed to a system of option agreements. The option agreements are similar to the prospecting licences in that they coincide with the DETI licences in area and their renewal periods. DETI notes:

³3URVSHFWLQJ OLFHQFHV IRU SUHFLRXV PHWDOV DUH LVVXHG E\ WKH &URZQ (VWDWH &RPPLVVLRQHUV (CEC) and companies wishing to explore for precious metals should apply simultaneously to the CEC and the Department for licences. Once the Department issues its licence, CEC will QRUPDOO\LVVXHDFRQFXUUHQWOLFHQFHIRUDFRQWHUPLQRXVDUHD´ '(7,0LQHUDOVDQG3HWUROHXP Exploration and Development in Northern Ireland 1997-2000).

Dalradian Resources, through its subsidiary Dalradian Gold, presently has the benefit of the following DETI prospecting licences: DG1/08, DG2/08, DG3/11 and DG4/11. In addition the company, through Dalradian Gold, has the benefit of the following CEC option agreements: DG1, DG2, DG3 and DG4.

The four DETI licences correspond to the four CEC option agreements thereby granting base and precious metal mineral exploration rights to four contiguous pieces of ground. The licences cover a total area of some 844 km2. (See Figure 4.1.)

The approximate centre of licences DG1 and DG2 is located at latitude 54º ¶´QRUWKDQG longitude 7º ¶´ZHVW7KHWZRDGGLWLRQDOOLFHQFHDUHDV'*DQG'*OLHWRWKHQRUWKDQG west of DG1.

Licence and option agreement maps, showing the boundaries of DG1, DG2, DG3 and DG4, have been issued by DETI and the CEC. DETI utilizes the Irish National Grid system of easting and northing which is then used for reference. The Tyrone project is located at approximately 25700 mE and 38600 mN.

Details on the eight licences, including size and expiry dates are presented in Table 4.1 below. The DETI licences and the CEC option agreements expire on the same dates and so only four records are shown in Table 4.1.

Table 4.1 Licence Details

Licensee A rea Licence A rea Minerals Date First Reissued Date of Number (km2) G ranted Expiry Dalradian Gold Curraghinalt DG1/08 167.5 All 1 01/01/2002 01/01/2008 31/12/2013 Limited Dalradian Gold Mountfield DG2/08 184.5 All 1 01/01/2002 01/01/2008 31/12/2013 Limited Dalradian Gold Newtownstewart DG3/11 248 All 1 23/04/2003 23/04/2011 23/04/2017 Limited East Dalradian Gold Sawel-Dart DG4/11 244 All 1 23/04/2003 23/04/2011 23/04/2017 Limited 1 - Concurrent licences/option agreements with DETI & CEC.

The mineral resource estimate presented in this report is located entirely on the property covered by licence DG1. A net smelter return royalty of 2% is payable to Minco plc on a portion of DG1.

Table 4.2 lists the required annual expenditures made on each of the four licence areas since acquisition by Tournigan in 2002. In the reissuing and renewals undertaken in 2011, the requirement for a minimum expenditure for a work program was removed. The licences outline a general work program to be undertaken on the four licences without a minimum spending commitment.

Prospecting licences (and CEC option agreements) in Northern Ireland are issued for a 2-year period and may be renewed (or extended), subject to relevant conditions being met and satisfied, on two occasions for a period of two further years, for a total of six years. At the end of each 2-year period a progress work report is submitted to the licensing authorities in order to support renewal, along with a letter indicating the intention to continue work on the license into the following 2-year period. At the end of each 6-year period a full reissuing application process must be undertaken whereby a full 6-year work report is submitted to the licensing bodies along with a new application for another 6-year period. Renewals, extensions and reissuing are not automatic.

The CEC option agreements allow the option holder the option of taking out a lease on the land. The tenant agrees to pay rent/royalty to the landlord (in this case the CEC) on the production of precious metals.

Table 4.2 Licence Expenditures

F rom To Expenditure Licence Number Reporting Stage (day/month/year) (day/month/year) (£) DG1/02 Reissued (6 years) 01/01/2002 012/01/2003 130,000 01/01/2003 012/01/2004 150,000 DG1/02 First extension 01/01/2004 012/01/2005 180,000 01/01/2005 012/01/2006 225,000 DG1/02 Second extension 01/01/2006 012/01/2007 262,500 01/01/2007 012/01/2008 262,500 DG1/08 Reissued (further 6 years) 01/01/2008 012/01/2009 350,000 01/01/2009 012/01/2010 350,000 DG1/08 First Extension 01/01/2010 01/01/2011 500,000 01/01/2011 01/01/2012 500,000

DG2/02 Reissued (6 years) 01/01/2002 01/01/2003 30,000 01/01/2003 01/01/2004 60,000 DG2/02 First extension 01/01/2004 01/01/2005 75,000 01/01/2005 01/01/2006 100,000 DG2/02 Second extension 01/01/2006 01/01/2007 137,500 01/01/2007 01/01/2008 137,500 DG2/08 Reissued (further 6 years) 01/01/2008 01/01/2009 229,000 01/01/2009 01/01/2010 222,500 DG2/08 First Extension 01/01/2010 01/01/2011 100,000 01/01/2011 01/01/2012 100,000

TG-3/03 (DG3/03) Reissued (6 years) 24/05/2003 23/05/2004 10,000 24/05/2004 23/05/2005 10,000 TG-3/03 (DG3/03) First extension 24/05/2005 23/05/2006 20,000 24/05/2006 23/05/2007 20,000 TG-3/03 (DG3/03) Second extension 24/05/2007 23/05/2008 30,000 24/05/2008 23/05/2009 40,000 DG3/09 Reissued (further 6 years) 24/05/2009 23/05/2010 40,000 24/05/2010 23/05/2011 40,000 DG3/11 Reissue (further 6 years) 24/05/2011 23/05/2012 No Requirement

TG-4/03 (DG4/03) Reissued (6 years) 24/05/2003 23/05/2004 10,000 24/05/2004 23/05/2005 10,000 TG-4/03 (DG4/03) First extension 24/05/2005 23/05/2006 20,000 24/05/2006 23/05/2007 20,000 TG-4/03 (DG4/03) Second extension 24/05/2007 23/05/2008 30,000 24/05/2008 23/05/2009 40,000 DG4/09 Reissued (further 6 years) 24/05/2009 23/05/2010 40,000 24/05/2010 23/05/2011 40,000 DG4/11 Reissued (further 6 years) 24/05/2011 23/05/2012 No Requirement

Licences DG1 and DG2 (and the corresponding CEC option agreements) were reissued in January, 2008. The licences and option agreements were issued for a term which may total up to six years (subject to the conditions described above). A first renewal of these licences and agreements was granted in January 2010 and a second renewal was granted in November 2011 with an effective date of January 2012. Reissuing of these licences and option agreements is scheduled for January 1, 2014.

Licences DG3 and DG4 (and the corresponding CEC option agreements) were due for their first extension in April, 2011. Due to timing problems in the renewal process, the DETI licences were reissued as new licences in April 2011 and the CEC licences were reissued as option agreements. The licences and option agreements were issued for a term which may total up to six years (subject to the conditions described above) with reissuing scheduled again for April 24, 2017.

Dalradian Gold provides annual reports in support of the licences and options agreements, usually around March of each year for DG1 and DG2, and in August of each year for DG3 and DG4. Included in the reports is a summary of the work performed for the year and an audited summary of the spending.

Licensees are required to give notice of their intention to enter land to carry out work and must seek the agreement of landowners before entering their property. Compensation is generally payable to the landowner for any damage caused during exploration.

DETI is required to consult with other departments and with public bodies concerning its intention to issue a licence and is also required to place notices in the Belfast Gazette and at least one local newspaper. This is primarily to allow the owners of surface land within the area under application the opportunity to make their views known.

'(7, QRWHV WKDW D GUDIW OLFHQFH DQG D µOHWWHU RI RIIHU¶ DUH SURYLGHG WR DSSOLFDQWV RQFH DOO comments have been considered. The letter of offer may contain a number of conditions although DETI notes that, at the prospecting stage, it is usually sufficient for the applicant to inform all listed contacts of its plans and progress. When the conditions set out in the letter of offer are accepted and the terms of the draft licence agreed, the licence is executed by DETI and the company.

DETI states that planning permission is not required for early stage exploration although the Planning Service of the Department of the Environment must be informed of the planned ZRUNLQFOXGLQJWKHQDWXUHDQGVFDOHWLPHDQGORFDWLRQRIWKHFRPSDQ\¶VDFWLvities and drill hole locations.

4.3 LOCATION OF MINERALIZED ZONES

Previous exploration has identified the Curraghinalt vein system. An adit was developed by Ennex between 1987 and 1989 to access the deposit. The location of the Curraghinalt gold deposit relative to the boundaries of DG1 and DG2 can be seen in Figure 4.2. Details of the locations of the mineralized zones, the portal and adit are shown on Figure 4.3.

Figure 4.2 Curraghinalt Gold Deposit Location

Source: Dalradian Resources, 2011

Figure 4.3 Locations of Principal Veins and Adit on DG1

Source: Dalradian Resources, 2011

4.4 PERMITS

Licence holders are required to give up to four weeks¶ notice of their intention to enter land to carry out work and must seek the agreement of landowners before entering their property. Compensation is generally payable to the landowner for any damage caused during exploration.

The issuing of a prospecting licence by DETI allows permitted development to take place without additional planning permission. The Planning Service of the Department of the Environment must be notified in writing prior to the work taking place giving details of the location of the proposed development, target minerals, details of plant and operations and anticipated timescale. In addition, DETI must also be notified in writing at least 28 days before work commences if there are any environmentally designated areas in close proximity to the work. This is required to allow DETI and other agencies to determine if any further studies or permissions will be required before the work can take place. Before starting any borehole and upon completion, notification is required to be given to the Health and Safety Executive of Northern Ireland (HSENI).

Dalradian Gold reports that it has negotiated land access and compensation agreements with all the landowners for the present drilling program. Additional access agreements will be required when drilling to the east and west of the known mineralization takes place. Dalradian Gold indicates that it expects to be able to successfully negotiate access to the required land to undertake any future drill programs on the Curraghinalt property.

4.5 ENVIRONMENTAL CONSIDERATIONS

Micon has not undertaken a review of potential environmental concerns or liabilities at the Tyrone project.

$WWKHWLPHRI0LFRQ¶VILUVW VLWHYLVLWLQ $XJXVW GULOOLQJZDV WKHRQO\ DFWLYLW\EHLQJ carried out. Shortly after that, all work ceased and no activity was occurring during the second site visit in 2009. Drilling resumed in early 2010 and has continued, with breaks, up to the time of writing.

Environmental considerations are discussed in detail in Section 20 of this report.

5.0 A C C ESSIBI L I T Y, C L I M A T E, L O C A L R ESO UR C ES, IN FR AST RU C T UR E A ND PH YSI O G R APH Y

Access to the Tyrone project is via a number of highways and local roads. These include the B48 from Omagh to Gortin and the B46 from Gortin to Greencastle. Local county roads, private roads and farm tracks provide generally good access within the property. Figure 5.1 shows the network of local roads and rivers around the Tyrone project area.

Local climate conditions are temperate. The average annual temperature is 9º C. Average daily temperatures vary between 4.1º C and 14.7º C throughout the year. Average annual precipitation is 852 mm, the majority of which falls in the winter months between September and January (average >80 mm per month). Snowfall is usually restricted to areas of higher elevation and occurs on 10 days or less per year. Exploration activities can generally be conducted year-round.

The town of Omagh (population 20,000) provides lodging and local labour, as do smaller local villages. Few experienced mining personnel are available locally although there is a small mining industry in Northern Ireland (salt and gold) and the Irish Republic has a number of underground base metal mines. There is a large quarrying industry in Northern Ireland. The principal economic activities in the area of the licences are sheep farming and, to a lesser extent, the raising of beef cattle.

Figure 5.1 Tyrone Project Licence Map, Major Road Access and Drainages

Source: Dalradian Resources, 2011

Belfast is the capital city of Northern Ireland and supports a population of approximately 300,000. From Belfast, Omagh and the project can be reached by paved road, more than half of which is dual carriageway (limited access highway). The over-road distance is approximately 110 km and requires less than 1.5 hours in good weather. Belfast is served by both a domestic and international airport with frequent daily flights to the rest of the UK and Europe.

The village of Gortin (pronounced Gorchin) is located a few kilometres from the Curraghinalt gold deposit at the western edge of licence DG1 (see Figure 5.1). Dalradian Gold has a small field office there as well as core storage facilities, all of which are rented. Gortin is centrally located within the licence areas and well situated to support the exploration program. Dalradian Gold also leases an office and core storage facility in Omagh near the road leading to Gortin. Geology and administration staff are located there as well as the principal core logging and storage facility.

A principal power sub-station is located at Plumbridge, approximately 22 km north of Omagh and the main 110-KV power line runs just outside Omagh. Local water resources are abundant.

Topography is rolling hills and broad valleys (see Figure 5.2). Glacial deposits and peat cover much of the area resulting in mixed forest and heathlands, as well as the farmland in the valleys. Relief ranges between around 100 masl in the major river valleys to some 550 masl.

Figure 5.2 General View of the Curraghinalt Deposit Area Looking South

6.0 H IST O R Y

6.1 ACQUISITION HISTORY

The property containing the Curraghinalt deposit was acquired initially by Ulster Base Metals (which later became Ulster Minerals) in 1981, an entity which later became a wholly-owned subsidiary of Ennex International plc. Ennex conducted exploration on the property between 1982 and 1999. Ennex sold its interest in Ulster Minerals to Nickelodeon in January, 2000. In August, 2000, the name of Nickelodeon was changed to Strongbow Resources Inc., and subsequently to Strongbow Exploration Inc. (Strongbow).

In February, 2003, Tournigan entered into an option agreement with Strongbow to earn an interest of up to 100% in the Curraghinalt deposit, located within a prospecting licence known as UM-11/96. Terms included staged exploration expenditures of $CDN4.0 million over a period of seven years and the delivery of a bankable feasibility study, and issuing shares to Strongbow at a price based on a 90-day trading average. At the same time, Tournigan entered into a similar option agreement with Strongbow in respect of its ³7\URQH project´ ORFDWHG within prospecting licence UM-12/96. Tournigan established Dalradian Gold as a wholly- owned subsidiary through which it would earn its interests in the Curraghinalt (UM-11/96) and Tyrone (UM-12/96) properties.

In the following year (February, 2004), Tournigan entered into a letter agreement with Strongbow for the outright purchase by Tournigan of all of the issued and outstanding shares of Ulster Minerals through the issue of 5 million shares of Tournigan. The earlier option agreements were terminated and replaced by the letter of agreement. A net smelter royalty of 2% held by Ennex was transferred to Minco plc. Full transfer of ownership in Ulster Minerals to Tournigan was completed in December, 2004.

Tournigan then applied to the licensing authorities, and received licences TG-3 and TG-4 (both minerals and precious metals), to the northwest of UM-11/96 and UM-12/96.

Ulster Minerals licenses UM-11/96 and UM-12/96 were later renamed UM-1 and UM-2, and ultimately DG1 and DG2. During the renaming and reregistering process the internal boundary between DG1 and DG2 was reoriented from east-west to a position which reflects the approximate location of the Omagh Thrust Fault (see geological descriptions in Section 7 and Figure 7.2). Older maps from previous operators may show the prior internal boundary. Details of the current licences and option agreements are provided in Section 4, above.

In October, 2009, Dalradian Resources completed a purchase and sale agreement with Tournigan to acquire all of the issued and outstanding shares of Dalradian Gold which included the licences, mineral rights and surface rights (including easements) in the Area of Interest. The Area of Interest is defined in the agreement as Mineral Prospecting Licences DG1, DG2, TG-3 and TG-4, the latter two being renamed to DG3 and DG4 by DETI after DFTXLVLWLRQ VHH6HFWLRQ $VRIWKHWLPHRI0LFRQ¶VVHFRQGYLVLWLQQRH[SORUDWLRQ work had been completed by Dalradian Resources although Aurum (the mineral exploration

contractor used by Tournigan) had been retained to continue its management of the exploration program over DG1 to 4. During this time Aurum completed a program of data compilation and targeting.

6.2 EXPLORATION HISTORY

The exploration history on the property is described by Tully (2005). Gold was recognized in the gravels of the Moyola River in 1652 and in the 1930s, an English company reported plans for alluvial gold mining in a prospectus. The licences comprising the Tyrone project were held by others prior to Ennex/Ulster Minerals and a report on the gold potential of the area was prepared by the Geological Survey of Northern Ireland (GSNI). However, 7XOO\¶VUHSRUW concentrated on the Curraghinalt gold deposit and his description of the mineral exploration history concentrates on the DG1 licence area and the work completed by Ulster Minerals, Ennex, Strongbow and Tournigan.

Tully (2005) noted:

³'XULQJWKHSHULRGWR5LR7LQWR)LQDQFHDQG([SORUDWLRQ ³5LRILQH[´ FDUULHGRXW an exploration program consisting of geological mapping and geophysical and geochemical VXUYH\V VHDUFKLQJ IRU ³SRUSK\U\ FRSSHU´ GHSRVLWV LQ WKH 0LG-Ordovician Tyrone Volcanic belt of rocks. This work was focused in an area located immediately to the south of the Curraghinalt deposit on Exploration License UM 2/02. The work was followed up with drilling of 6 diamond drill holes in the Cashel (Formil) area. The best intersections were 131 feet grading 0.17% Cu in drill hole H2, and 151 feet grading 0.12% Cu in hole H4. Riofinex ceased exploration work in the area in 1974.

³Clifford et al. (1990) noted that during the mid-1970s, the Geological Survey of Northern Ireland (GSNI) reported on basic geological and geochemical studies that it had conducted in the area which confirmed the presence of gold in stream gravels over a wide area. These reports also recorded anomalous values in stream sediments of elements that are commonly associated with gold.

³On the basis of the above data, Ennex applied for a License in the area in 1980. This license was granted in 1981 and from that date to 1997 Ennex carried out an extensive exploration program consisting of geological mapping, prospecting, geochemical and geophysical surveys and diamond drilling in both the Tyrone Volcanic Group and the Dalradian Metasediments located immediately to the north.

³During 1987 Ennex drilled 25 holes in the Cashel Rock gold deposit (located approximately 4-km southwest of Formil) which was discovered by deep overburden geochemistry, geophysics and trenching. Gold mineralization is located in silicified rhyolite and is generally conformable to the general northwest trending dip of the rhyolite flows. In general long intercepts of very low anomalous gold values were intersected (up to 145 m grading 394 ppb Au), with some shorter better grade zones (5.45 m @ 4.3 g/t Au, 6.9 m @ 1.3 g/t Au and 3.63 m @ 30.6 g/t Au).´

[Tully quotes 25 holes at Cashel Rock, however, examination of old drill records indicates that 14 had been completed by 1990. It is believed that this error is caused by adding in drilling by Riofinex at a different location.]

³In addition, Ennex drilled 11 reconnaissance holes in the Tyrone Volcanic rocks and in the Dalradian strata. Some of these holes returned anomalous base or precious metal assays, but no definite deposits were intersected.

³In addition to the base metal exploration programs carried out by Ennex on the Tyrone Volcanic Group of rocks, they focused on developing and exploring the Curraghinalt deposit, which was discovered in 1983. During the period 1983-1997, Ennex completed 4 major GULOOLQJSURJUDPVWRH[SORUHWKH&XUUDJKLQDOWGHSRVLW´

The four phases of exploration at Curraghinalt, between 1983 and 1997, are summarized as follows:

Phase 1 (1983-July, 1987):

Detailed prospecting, geochemistry and geophysics. 68 trenches (2,856 m) and 63 diamond drill holes (6,387 m).

Phase 2 (August, 1987-March, 1989):

Underground development program including development of an adit (412 m), lateral drifting (325 m) and raising (60 m). Lateral development utilized a Dosco SL 120 road header.

Phase 3 (May, 1995-March, 1996):

Detailed and reconnaissance drilling to test previously inadequately drilled veins. Reconnaissance drilling of veins to the southwest of previously-drilled areas. Total 59 holes (4,980 m).

Phase 4 (June, 1996-May, 1997):

Infill drilling on 25- to 30-m centres in the main vein areas. Drilling of 50 holes (5,400 m).

The locations of the drilling and trenching activities carried out by Ennex are shown in Figures 6.1 and 6.2.

Between 1997, when Ennex transferred its interest to Nickelodeon, and late-2002, when the agreement was signed between Strongbow and Tournigan, little work was done on the property.

Figure 6.1 Ennex Drilling at Curraghinalt

Source: Dalradian Resources, 2011

Figure 6.2 Ennex Trenching at Curraghinalt

Source: Dalradian Resources, 2011

The Tournigan exploration can be broken into several phases.

Phase 1, 2003 to January, 2005:

22,910 soil samples collected. Small geophysical survey conducted. Mapping and prospecting on the DG1 Licence area. 26 diamond drill holes (4,391 m) drilled at Curraghinalt. 7 diamond drill holes drilled at Glenlark.

Phase 2, January, 2005 to 2007:

2 diamond drill holes drilled in the area of the Crows Foot-Bend. 24 diamond holes drilled (infill drilling) on the Southeast Extension target.

Phase 3, 2007 to 2009:

Resource estimate completed by Micon (November 29, 2007). 4 deep diamond drills holes completed

After completion of the 2007 drill program Tournigan ceased all exploration activity at the Curraghinalt deposit and, except for some prospecting on TG-3 and TG-4 in 2008, the property remained inactive until acquisition by Dalradian Resources in 2009.

Williams (2008) described the exploration history of what is now the DG2 licence area as follows:

³0LQHUDOexploration has been completed over the TVG [Tyrone Volcanic Group] in the 1970s by Riofinex and Glencar and in the 1980s and 1990s by Ulster Minerals Limited. This has comprised of geological mapping, prospecting, geochemistry, ground magnetic and IP surveys, followed by trenching and drilling. An airborne INPUT was flown by Geoterrex in 1979 for Moydow Limited. This exploration has led to the discovery of numerous sub- HFRQRPLF\HWVLJQLILFDQWJROGDQGEDVHPHWDORFFXUUHQFHV´

The principle target of interest for Ennex on DG2 was the Cashel Rock showing where a gold-mineralized silicified rhyolite breccia outcrop is exposed at surface. At this location 15 shallow drill holes targeted the zone just below surface. The results and example sections were presented in Hennessey and Mukhopadhyay (2010). They are not relevant to the mineral resource estimate for the Curraghinalt deposit presented in this report and are not repeated here.

6.3 HISTORICAL MINERAL RESOURCE ESTIMATES

In May, 1997, a polygonal resource estimate was prepared on behalf of Ennex by CSA. (See CSA 1997). The estimate was prepared on a minimum mining width of 1.25 m and a cut-off JUDGH RI JW $X 7KH HVWLPDWH UHIHUUHG WR DV D ³SRO\JRQDO UHVRXUFH FDOFXODWLRQ´ LV

summarized below in Table 6.1. It is an historical estimate and is provided for information purposes only. It predates, and is not compliant with, NI 43-101 and should not be relied on.

Table 6.1 Summary of Polygonal Mineral Resource Estimate, 1997

G rade Vein Tonnes (g/t Au) Sheep Dip 21,257 15.68 T17HW 65,676 17.24 T17C 48,788 16.09 No. 1 116,483 17.39 T11F 38,144 13.21 106-16 97,962 21.61 ABB 79,787 13.03 Total 468,097 16.96

Tully prepared a mineral resource estimate in 2005 and, in 2007, Micon also completed a mineral resource estimate on the Curraghinalt deposit for Tournigan (Tully, 2005 and Mukhopadhyay, 2007) and again in 2010 for Dalradian Resources (Hennessey and Mukhopadhyay May 10, 2010). The previous Tully and Micon estimates were NI 43-101 compliant but were not considered to be historic estimates under the previous version of the instrument. While the new instrument (June 30, 2011) allows for the use of estimates made prior to an issuer entering into an agreement to acquire a property, the 2007 and 2010 Micon estimates have been superseded by the estimate presented herein. The 2005, 2007 and 2010 estimates can be found on SEDAR (www.sedar.com) filed under Tournigan and Dalradian Resources.

6.4 H IST O RI C A L PR O DU C T I O N

Micon is not aware of any historical mineral production from the Tyrone project other than the bulk sample taken from the Ennex adit.

7.0 G E O L O G I C A L SE T T IN G A ND M IN E R A L I Z A T I O N

7.1 REGIONAL GEOLOGY

The geological setting of the Curraghinalt deposit area, as set out below, is adapted from Tully (2005).

The geological history of the project area and the immediate surrounding area is related to the closing of the Iapetus Ocean in Ordovician time. This major geological event involved the following individual events:

Obduction of oceanic crust (dated at 484 to 480 million years) from the southeast over the older Dalradian metamorphic rocks (Hutton 1993; Earls et al. 1996);

Building of a volcanic arc with several volcanic cycles in middle Ordovician time (dated at 470 - 464 million years);

Stitching intra-arc tonalites (470 to 466 million years) and younger continental arc granites (467 - 464 Ma);

Overthrusting of older Dalradian (Proterozoic and possibly lower Paleozoic) clastic meta- sedimentary rocks from the northwest (c 470 to 455 million years).

The Oceanic rocks which occur on both sides of the Central Inlier, and to the south of the Tyrone Volcanic Group, form what is known as the Tyrone Ophiolite within the Tyrone Igneous Complex, and are dominated by gabbros and basic dykes. The age of these rocks is has been recently dated as 484 to 480 million years (Cooper et al. 2010).

The volcanic arc package of rocks, the Tyrone Volcanic Group is considered to be middle Ordovician in age (dated at 470 to 464 million years) and forms part of the Tyrone Igneous Complex. The upper part of the volcanics is comprised of bimodal sequences of basaltic and andesite to rhyolitic submarine and subaerial lavas, volcaniclastics with chert horizons and intercalations of graptolitic shales. These lithologies are preceded by submarine basaltic andesite pillow lavas and associated intrusives.

A series of porphyry bodies and a series of calc-alkaline granitic intrusions intrude the volcano-sedimentary sequence. The history of these units began with the northwest-directed obduction of the ophiolite, with its related overlying arc, onto an offshore Proterozoic continental fragment (the Tyrone Central Inlier). During the later stages of this obduction, and before the arc had become inactive, these amalgamated units collided with and were overridden by the Dalradian rocks of the continental margin, causing the c.470 Ma Caledonian orogeny in this area. (Alsop and Hutton, 1993a).

The rocks thrust over the Tyrone Volcanic Group by the Omagh Thrust form part of the Dalradian Supergroup. The Dalradian rocks consist of a metamorphosed clastic sedimentary package of biotite to garnet grade semi-pelites, (siltstone) psammites (impure sandstone) and chloritic-sericitic pelites (shale). The younger Dalradian Supergroup clastic sediments of the Grampian terrane accumulated in passive-margin rift basins between c.800 Ma and the early Cambrian during the breakup of the Late Precambrian supercontinent Rodinia, and the

formation of the Iapetus Ocean. The Glengawna Formation contains a distinctive assemblage of psammites, talcose schists and graphitic pelites that host the Glenlark gold and base metal occurrence. The Curraghinalt deposit occurs in the Mullaghcarn Formation that is comprised of fine grained clastic meta-sedimentary rocks (psammite, semi-pelite and chlorite-rich pelite).

There are many similarities between the Tyrone Volcanic Group rocks and the rocks of the Appalachian Mountains, Atlantic Canada, Ireland, Scotland and Scandinavia. The rocks in these areas were part of the continuous Appalachian-Caledonian Orogen, prior to the opening of the Atlantic Ocean basin in middle Mesozoic time. It is postulated that the Bathurst, New Brunswick area rocks (Tetagouch Group) are correlative with rocks in southern Ireland (Avoca, in County Wicklow), whereas the Tyrone area Ordovician rocks are more closely related to those at Buchans, Newfoundland. Similarly, it is postulated that late Proterozoic and early to late Cambrian sediments present on the Great Northern and Baie Verte peninsulas of Newfoundland are analogous to that of the Dalradian in Northern Ireland.

Figure 7.1 shows the regional geology of the north of Ireland.

7.2 PROPERTY GEOLOGY

Figure 7.2 shows the geology of the area covered by the DG1 and DG2 licences and the location of the Curraghinalt deposit vein swarm.

7.2.1 Curraghinalt Gold Deposit Area - Licence DG1

Williams and Archibald (2009) described the geology of the DG1 licence areas as follows:

7.2.1.1 Stratigraphy

³7KHURFNVXQGHUO\LQJWKHOLFHQFHDUHDEHORQJWRWKH&HQWUDO+LJKODQG *UDPSLDQ 7HUUDQH and consist of mid- to late-Neo-proterozoic rocks of the Dalradian Supergroup and a small area of unconformable Lower Carboniferous sandstone. The Dalradian Supergroup is composed of a thick sequence of clastic marine sediments, with minor volcanic units, deposited in a rift basin, and can be correlated with successions in Scotland (Highlands), Republic of Ireland (Cos. Donegal, Mayo and Galway) and perhaps the Fleur de Lys Supergroup in Newfoundland (Kennedy, 1975) and the Eleonore Bay Supergroup in eastern Greenland (Soper, 1994).

Figure 7.1 Regional Geology of the North of Ireland

60 N

Source: Dalradian Resources, 2011

Figure 7.2 Local Geology in the Area of Curraghinalt

DG4

61 DG2

DG1

Source: Dalradian Resources, 2011

³7KH 'DOUDGLDQ 6XSHUJURXS LV GLYLGHG LQWRIRXU JURXSV LH *UDPSLDQ $SSLQ $UJ\OO DQG Southern Highland [Table 7.1], of which only the Argyll and Southern Highland groups are present in Northern Ireland (Strachan et al. 2002). The rocks in the Argyll Group indicate a change in tectonic and depositional environment, varying from a relatively stable shelf environment (Easdale and Crinan subgroups) to a floundered shelf environment associated basaltic magmatism (Tayvallich Subgroup) caused by the crustal thinning and stretching prior to the opening of the Iapetus Ocean. Erosion and reworking of the basic igneous rocks formed contemporaneous volcaniclastic sediments (green beds) present in the upper part of the Argyll Group and the lower part of the Southern Highland Group. The Southern Highland Group formed in an unstable shelf environment and, as such, has developed turbidite facies. These vary from coarse-grained proximal arenites (conglomerates) to fine-grained outer fan pelites (shales).

³$OORIWKe Dalradian Supergroup lithologies were deformed during the Grampian Orogeny, and now lie on the inverted southern limb of the Sperrins Nappe. The dominant structural feature in the licence area an isoclinal F2 anticline, plunging to the northeast, which has been refolded to form a northwest plunging anticline.

³Based on information from adjacent areas it is certain that the Dalradian Supergroup rocks were covered in younger lithologies. However, the only evidence for the presence of these units in the licence area are [sic] a sandstone-dominated Lower Carboniferous basin that is present 3 km to the north of Newtownstewart and 3 km to the east of Gortin, and another basin DWWKHH[WUHPHVRXWKRIWKHOLFHQFH´

Table 7.1 Correlation of Dalradian Rocks in Scotland and North of Ireland

Scotland South and Central Sperrin Mountains Northeast Co. Co. Donegal North South Antrim G roup Subgroup Formations Londonderry Mullaghcarn Croaghgarrow Glengawna Glendun Southern Ballykelly Highland Glenelly Runabay Shanaghy Mullyfa Claudy Dart Altmore Torr Head Tayvallich Aghyaran Dungiven Limestone Owencam Argyll Crinan Killeter Quartzite Newtownstewart Murlough Lough Esk Psammite, Easdale Lough Mourne Grit Geological Survey of Northern Ireland, 1997.

Southern Highland G roup

³In the south Sperrin Mountains the Southern Highland Group is composed of four formations; Dart, Glenelly, Glengawna and Mullaghcarn. The Glenelly and Glengawna formations are equivalent to the Ballykelly Formation in the north Sperrin Mountains. Dominant lithologies within these formations are psammites, semi-pelites, pelites and amphibole schists.

Glenelly Formation

³The Glenelly Formation outcrops between 4 km northeast and southwest of Gortin and comprises of four mappable members, i.e., Oughtboy Burn, Clogherny, Golan Burn and the Garvagh Bridge. The Oughtboy Burn Member is composed of approximately 70 m of dark green, chlorite- and biotite-rich volcaniclastics semipelite (green bed) with abundant feldspar (albite) porphyroblasts. The overlying Clogherny Member is a calcareous, pale-green to white tourmaline-rich schistose semipelite. Medium- to coarse-grained psammite also comprises this member. A similar lithology is encountered in the overlying Golan Burn Member, except tourmaline is less abundant, and it contains thin beds of impure meta-limestone. The Garvagh Bridge Member is composed of fine- to medium-grained, pale grey, calcareous semipelite schist, and thin bedded meta-limestones. The calcareous units represent basin shallowing during periods of deeper water sedimentation (Cooper and Johnston, 2004).

Glengawna Formation

³The formation is approximately 400 m thick and comprises a distinctive association of graphitic pelites, semi-pelites, chloritic semi-pelites and psammites. These lithologies contrast markedly with the silvery-grey lithologies of the underlying Glenelly Formation.

Mullaghcarn Formation

³The formation is only present in the extreme southeast and south of the licence area, and it consists of psammite, semi-pelite, chloritic-semipelite and pelite schists. Garnets are sometimes present, and may be replaced by chlorite or hematite. Bedding is difficult to recognise, but it is generally acknowledged that the formation is overturned. The Mullaghcarn is economically significant, since both the Curraghinalt and Cavanacaw gold deposits are hosted in this formation.´

Tully (2005), whose report concentrated on the Curraghinalt gold deposit, describes the geology of the deposit area as follows:

³7KH &XUUDJKLQDOW GHSRVLW LV located 3-km to the north of the northeast-southwest striking Omagh Thrust Fault. This major fault has thrust Dalradian Supergroup rocks from the northwest over Tyrone Volcanic Group rocks located to the south. The Dalradian metasediments on the northern side of the thrust strike NE-SW and dip to the northwest. The lithologies are interpreted to lie on the lower limb of a gently to moderately northwest-dipping southeast upward-facing, major recumbent overturned tight isoclinal fold. This fold is referred to as either the Sperrins Overfold or the Sperrins Nappe.

³The Dalradian stratigraphy in the license area is inverted, occupying the lower limb of the overturned Sperrins Nappe. The Dalradian Formations present in the area from the southeast (at the Omagh Thrust) to the northwest are, from the youngest to the oldest, the Mullaghcarn, Glengawna and Glenelly Formations. The Mullaghcarn Formation, which hosts the Curraghinalt deposit, consists of mixed semi-pelites, quartz semi-pelites and psammites. The Glengawna Formation comprises a mixed package of graphitic pelites, graphitic semi-pelites, chloritic semi-pelites and minor psammites. Furthest north is the Glenelly Formation, which is comprised of pelites, semi-pelites and quartz semi-SHOLWHV´

7.2.1.2 Structure

³7KHVWUXFWXUHVLQWKHLPPHGLDWHDUHDRIWKH&XUUDJKLQDOWDUHDKDYHEHHQGHVFULEHGLQGHWDLOE\ M. Boland (1997) and are summarized as follows:

³The Curraghinalt deposit consists of a WNW trending vein swarm. The veins are controlled by a series of high-angled, east-west trending shears that are interpreted to represent hangingwall accommodation structures caused by a ramp in the footwall of the Omagh Thrust. (Earls and al. 1996 [sic]). These structures are expressed geophysically as strong Fraser filtered VLF-EM linears. The shears are vertical to northerly steep dipping structures, which dextrally offset the veins.

³The veins are also offset sinisterly [sic] by ENE to NNE trending normal faults that dip to the NW and with general displacements of approximately 10 metres. Low angle, north-dipping WKUXVWIDXOWVDUHDOVRHYLGHQWLQWKHPDLQDFFHVVDGLW´

In 2007 the operators retained Dave Coller, P.Geo., Euro.Geol., a structural and economic geological consultant, to prepare an initial review of the Curraghinalt vein system. See Coller (2007). Coller concluded:

³$QRYHUDOOPRGHOLVVXJJHVWHGIRUDVKDOORZHDVWSOXQJLQJV\VWHPZLWKVWHHSZHVWSLWFKLQJ ore shoots. The implication is that the SE Extension represents a higher structural level in the system than the Central Resource area such that similar grade and thickness veins are SUHGLFWHGDWPRGHUDWHGHSWKVEHORZWKH6(([WHQVLRQ´

See Figure 7.3 for the relative locations of the Central Resource Area and Southeast Extension, and the relative locations of the principal identified mineralized veins on level 170.

Figure 7.3 Plan of Curraghinalt Vein System on Level 170

NORTH

SOUTH

Grid axes shown in metres. Source: Coller (2007).

Coller undertook a detailed interpretation of the vein geometry for two levels in the Central Resource Area, level 170 and level 220, which are 170 m and 220 masl, respectively. Coller also provided an interpretation for the entire system on level 170.

Coller tabulated the types of veins and their characteristics by means of the definitions in Table 7.2.

Table 7.2 Definitions for Geometry of the Vein System

Type Characteristic Representation Comments Vein complex Series of veins which probably all Envelope encompassing Previously defined as single veins. link in 3D, with one main vein or all the vein branches. Several vein zones and linked several en echelon high grade branches are potentially economic. veins and many branches. Vein zone Large single or closely spaced Width of zone containing In detail internal vein segments and branching veins regarded as a veins defined as one vein associated veinlets have separate single vein. with average grade as on Au assays but would be mined as a drill sections. single vein. Vein branch Veins branch connections Separate veins in drill Larger branches often high grade between veins zones within a vein sections. and may be mined by linkage to complex [sic]. main vein zone. Other veins Major veins which may link or Separate significant vein Could be defined as separate vein occur between vein complexes. zones not named. complexes with more drilling.

7.2.1.3 Central Resource Area

For the Central Resource Area Coller noted:

³0RVWRIWKHlarge high grade veins in the Central Resource Area are relatively distinct vein zones, with a few simple high grade branches and sometimes en echelon segments. The No 1 Vein Complex is more complex and develops multiple branches in the eastern part of the Central Resource Area.

1. Grade is very variable within each of the vein complexes but is consistently over 5 gpt IRUDOPRVWDOOWKHFRPSOH[HVDQGEUDQFKHV RQERWKFRPSLOHGOHYHOV« 7KHH[FHSWLRQLVWKH western part of T17 at the lower 220 level.

2. Over the 300-350 m strike length of the Central Resource Area, all of the vein complexes have three or more very high grade segments >25gpt [grams per tonne]. In the case of No 1 and 106-16 these segments are steep plunging (comparing levels) with a westerly SLWFK«7KDVWKHWKLFNHVWDQGPRVWFRQWLQXRXVKLJKJUDGHVHJPHQWZLWKLQZKLFKWKHUHDUH two very high grade segments of >50gpt.

3. Most of the branches within the vein complexes are high grade and of significant thickness to be included in a resource. No 1 complex has two main branches at the lower level which may separate at the higher level.

4. The main veins (and complexes) have varying dips to the north. The T17 has steep dipping segments of around 70 N although the eastern part of the vein dips more moderately at 55 N. The No 1 complex is generally moderately dipping between 40 and 50 N, whilst the average dip of most of the other veins is around 55-1´

For the other areas Coller made the following observations.

7.2.1.4 Southeast Extension

³$PVWULNHOHQJWKRIGULOOLQJKDVGHILQHGDKLJKGHQVLW\RIKLJK-grade veins which are FOHDUO\FRQWLQXDWLRQVRIWKHPDLQ9HLQ&RPSOH[HVGHILQHGLQWKH&HQWUDO5HVRXUFH$UHD« and which represent a significant Au resource. The main interpretation of the vein geometry showing the high grade distribution and probable linkage of the veins is taken from the 170 level. The other levels have less information but could be worked up for some parts of the SE [Extension].

1 The Vein No 1 and Vein 106-16 complexes develop as multi-branch systems with en echelon, but linked, thick high grade vein segments. ... These two vein complexes essentially merge on 170 level [sic].

2. The high grade segments (>11gpt) of all the vein complexes appears to continue as far east as the current drilling and hence is presumed to continue further.

3. The principal vein T17 appears to be a very straight vein with three or four high grade segments, thickening to the east.

4. The Crows Foot Vein complex in the southern part of the system appears now from the compilation to be an E-W vein system with multi-branches, several of high grade. This vein complex is parallel to the Kiln Shear. This may be a very significant vein complex which LVXQWHVWHG´

7.2.1.5 Attagh Burn Vein Complex

³7KH $WWDJK %XUQGULOOLQJ ZHVW RIWKH &HQWUDO 5HVRXUFH $UHDLV QRZ LQWHUSUHWHG DVD ZLGH complex of veins with the same WNW strike as those in the Central Resource Area. Within the complex there are 3-5 high grade parallel veins or branches. This complex of veins may be bounded to the south by a continuation of the Kiln Shear. However, displacements on the Kiln Shear do not appear to be very significant even in the Central Resource Area so the $WWDJK%XUQYHLQVPD\EHSUHVHQWVRXWKRIWKHVKHDU´

7.2.1.6 Northern Area Veins

³7KUHHVLJQLILFDQWKLJKJUDGHYHLQVKDYHEHHQGLVFRYHUHGLQWKH1RUWKHUQ$UHDLPPHGLDWHO\ north of the Central Resource Area. These veins are not well drill tested although the Mullen Vein appears to have a strike-length of at least 700 m and has three high grade segments and several branches similar to the main vein complexes. The Sheep Dip Vein has a significant high grade segment and at least one main branch and dips [at] around 75-85N than the other veins.

³The Road Vein is essentially untested with one relatively deep level intersection in 04CT 26 RIP#JSW´

Coller concluded:

³7KH FRQVLVWHQW JHRPHWU\ DQG SDWWHUQ WR DOO WKUHH RI WKHVH IHDWXUHV OLQNV ERWK WKH YHLQ branching, the thickness variations and the high grade shoots in a kinematic way that relates to the genesis of the system. The common westerly pitching vector to the system is almost certainly related to the syn-NLQHPDWLF µIDXOW¶ FRQWURO $W WKLV VWDJH WKH LQWHUSUHWDWLRQ LV consistent with a movement direction which is normal to the vein i.e., north block up dip to the west (compressional) as suggested by the geometry of the rolls in plan and consistent with the larger scale en echelon vein style and relay branching (above) which is left lateral.´

The vein system and controlling structures of the Curraghinalt gold deposit are interpreted to strike to west-northwest towards, and possibly onto, the DG3 licence and southwest corner of the DG4 licence as seen in Figure 7.2.

7.2.2 Licence DG2 - Tyrone Volcanic Group

Williams and Archibald (personal communication via e-mail, December, 2009) provides the following description of the Tyrone Volcanic Group rocks on licence DG2.

³7KH7\URQH9ROFDQLF*URXS 79* LVORFDWHGWRWKH6(RIWKH'DOUDGLDQPHWDVHGLPHQWVDQG is Ordovician in age. It is structurally demarcated on the southeast by the much older Pre- Cambrian Complex of the Central Metamorphic Inlier and further to the northeast by an

ophiolite suite. On the northwest side the Ordovician has been overthrust by the Dalradian, late Pre-Cambrian to early Palaeozoic, metasedimentary rocks on the Omagh Thrust. The thrust dips northwest at a moderate dip. Both ends of the northeasterly striking Ordovician belt are covered by younger, Devonian, Old Red Sandstone and the Lower Carboniferous.

³Work by Earls (1987) suggests the presence of three island arc volcanic cycles favourable for VMS [volcanogenic massive sulphide] deposits, adjacent to the back arc basal ophiolites, within the TVG. These cycles are designated from top to bottom as follows:

Cycle 3. Broughderg Formation

Cycle 2. Bonnety Bush - Glencurry Formation

Cycle 1. Dungate - Tanderagee Formation

³The top of each formation is marked by an erratic increase in volumetrically minor chert or ironstone, which is taken to represent the culmination of a cycle. The cherts are variably magnetic and discontinuous and occur at several horizons. For this reason, and their relatively shallow dip to the northwest, the cherts are difficult to trace on the magnetic surveys. The cherts assigned to the two lowermost cycles are jasperoid, indicative of an oxidizing environment, whereas the Cycle 3 cherts are grey to black suggesting reducing depositional conditions. Each sequence progresses from mafic to more felsic upwards.

³Basaltic flows are common in the lowermost parts of the Cycle 1, Dungate-Tanderagee Formation. East of the Dungate Fault the top of the cycle, fairly well marked by cherty zones within the volcanics lies close to extensive ophiolite suite rocks to the southeast. West of the Dungate Fault the Cycle 1 volcanic sequence is much thicker, with some suggestion of ophiolite suite rocks along its contact with the older Metamorphic Inlier.

³The Cycle 2, Bonnety Bush-Glencurry Formation hosts major quartz porphyry intrusions towards the central part of the TVG. These porphyry intrusions are, in all likelihood, the feeders for the rhyolites occurring stratigraphically above at the top of Cycle 2. The intervening tonalites dilate the section and for present purposes may be imagined as being removed to bring the quartz porphyry intrusions closer to their original stratigraphic proximity to the rhyolites.

³The Cycle 3, Broughderg Formation has lesser rhyolite development than the underlying formation, but the ratio of mafic to felsic flows and tuffs is less than in the preceding formations. Black argillite occurs in the upper part of the formation and associated weakly DXULIHURXVFKHUWV7KHWRSRIWKHIRUPDWLRQLVRYHUULGGHQE\WKH2PDJK7KUXVW´

Williams (2008) also notes the similarities between the Tyrone Volcanic Group and lithologies in Scandinavia and eastern Canada and their potential to host VMS style mineralization:

³7KH 7\URQH 9ROFDQLF *URXS FDQ EH URXJKO\ FRUUHODWHG ZLWK VLPLODU YROFDQLc packages in Québec, New Brunswick, Newfoundland and the Scandinavian Caledonides. Similar aged rocks in all four widely separated areas contain significant VMS deposits. Of these, the most significant are the numerous large deposits of the Bathurst District in New Brunswick (Luff

1995), and probably the famous Buchans deposits of Newfoundland (Swinden 1991; Thurlow and Swanson 1981). The latter have in the past been considered to be slightly younger, (Kean et al. 1981; Thurlow 1981; Bell and Blenkinsop 1981), but subsequent information (Nowlan and Thurlow 1984; Swinden 1990; Williams 1995, page 151) has shown them to be most likely time-correlative with the Bathurst area deposits. Other deposits of roughly similar age include several in the Eastern Townships of Québec (Trottier and Gauthier 1991; Gauthier 1995) and a number in the Trondheim district of Norway (Stephens et al. 1984). There are many similarities between the Tyrone Volcanic Group rocks and those in these other regions. Such correlation is not unreasonable, because the rocks of the Appalachian Mountains, Atlantic Canada, Ireland, Scotland and Scandinavia were part of the continuous Appalachian- Caledonian Orogen prior to the opening of the Atlantic Ocean basin in middle Mesozoic time (see, HJ6WHSKHQVHWDO:LOOLDPV:LOOLDPVHWDO ´>6HH)LJXUH@

³More recent work (van der Pluijm et al. 1995) has suggested that the volcanic belts are parts of several island arcs related to separate subduction zones in the Iapetus Ocean. It may well be that the Bathurst area rocks (Tetagouche Group) are correlative with rocks in southern Ireland (e.g., those at Avoca, in Co. Wicklow, Republic of Ireland) whereas the Tyrone area Ordovician rocks are more closely related to those at %XFKDQV1HZIRXQGODQG´

Figure 7.4 Appalachian Correlations

New Brunswick Ireland Nova Scotia

Newfoundland

Red dots are occurrences of significant VMS deposits in the Appalachian-Caledonide mountain belt. After Archibald (2009).

Ongoing work by Hollis (personal communication to G. Earls, 2011) has updated and revised much of the volcanic stratigraphy of the Tyrone Volcanic Group and has allowed a detailed chronostratigraphic framework of the evolution of the Tyrone Igneous Complex to be established. Hollis has interpreted the Tyrone Volcanic Group as forming at the same time as the Annieopsquotch terrane in Newfoundland that is host to the Buchans deposits. In addition

Hollis has recognized a common geological history between the Tyrone Volcanic Group and the Annieopsquotch terranes.

7.2.3 DG3 and DG4 Licences

7.2.3.1 Stratigraphy

Williams and Archibald (2009) described the geology of the DG3 and DG4 licence areas as the same as that presented for licence DG1 in Section 7.2.1.1 above.

The Dalradian Supergroup is divided into four groups, i.e., Grampian, Appin, Argyll and Southern Highland (see Table 7.1), of which only the Argyll and Southern Highland groups are present in Northern Ireland. Detailed descriptions are provided below (Williams and Archibald 2009).

A rgyll G roup: C rinan Subgroup

³The oldest rocks present in DG3 licence are psammites and pelites belonging to the Newtownstewart Formation. The Newtownstewart Formation is the equivalent of the Killeter Quartzite Formation. Both formations belong to the Easdale Subgroup. [G. Earls reports that they are part of the Crinan Subgroup, personal communication, 2012.]

Newtownstewart Formation

³The Newtownstewart Formation is present in the south-western and north-eastern part of the licence area (immediately south of Newtownstewart, and 3 km east of Strabane) and is exposed in the core of a recumbent F2 fold. The formation is composed of fine-grained, pale grey quartzose psammites in beds approximately 50 cm thick, and is interbedded with thin pelite units. Occasionally the formation shows graded pebbly beds. Since the base of the unit is not observed the exact thickness is unknown, but it is probably similar to the Killeter Quartzite Formation, i.e., ~1000 m.

A rgyll G roup: Tayvallich Subgroup

³The Tayvallich Subgroup overlies the Crinan Subgroup and is characterised by two contrasting rock types; limestones and pillowed basalts. Local names for the Tayvallich Subgroup in Northern Ireland include Aghyaran Formation (South and Central Donegal), Dungiven Formation (Counties Londonderry and Tyrone) and the Torr Head Limestone Formation (Co. Antrim). In the DG3 licence the Dungiven Limestone Formation is used to describe the Tayvallich Subgroup.

Dungiven Limestone Formation

³The Dungiven Limestone Formation is present throughout most of the licence, and it attains a maximum extent at Strabane where it is present in the nose of a northwest plunging anticline. Meta-limestone is the dominant lithology, and is associated with pelite, semi-pelite, psammite, quartzite, basaltic pillow lava and volcaniclastic sediments. The best exposure of this unit is 5 km north of Plumbridge on the eastern flanks of Butterlope Glen (GR 249300; 394700).

Basaltic pillow lavas occur at several localities, with individual pillows up to 1 m in diameter, and consisting of fine-grained, equigranular metabasalts. Volcanic textures such as triple junction chilled margins, gas vesicles and radial fracture patterns are all present. Meta- limestones are the dominant lithology at the contact with the Newtownstewart Formation and the overlying Claudy/Dart Formation. The limestone varies in texture from pale saccharoidal to dark crystalline and thickness from a few centimetres to several metres. The sequence has been extensively intruded by metabasite sills, and these are probably conduits to more extensive pillow lavas elsewhere in the Tayvallich Subgroup.

Southern Highland G roup

³Three formations belong to the Southern Highland Group in the north Sperrin Mountains; Claudy, Ballykelly, and Londonderry formations. In the south Sperrin Mountains the South Highland Group is composed of four formations; Dart, Glenelly, Glengawna and Mullaghcarn. The Glenelly and Glengawna formations are equivalent to the Ballykelly Formation in the north Sperrin Mountains. Dominant lithologies within these formations are psammites, semi- pelites, pelites and amphibole schists. All of the south Sperrin Mountain formations are present in the DG3 licence, whereas the Claudy Formation is the only north Sperrin Mountain Formation present.

Claudy Formation

³The Claudy Formation is present in the northern part of the licence area; to the north of Strabane [Figure 5.1]. The formation has a variable thickness, ranging from 750 to 2000 m, and is predominantly composed of coarse-grained quartzo-feldspathic psammite with graded bedding. The quartzo-feldspathic psammite contains coarse grains of blue and white quartz and pink and white feldspar. Beds within the psammite units can attain thickness up to 1 m, alternating with silvery grey phyllitic semipelite and rare greyish-black phyllitic pelite. The presence of biotite porphyroblasts, rare garnets, and chlorite indicate upper greenschist facies metamorphism. Two limestone units within the Claudy Formation are important marker units. These units are termed the Alla Glen Limestone Member and the Bonds Glen Limestone Member. The Alla Glen Limestone Member is pale grey to greyish blue, coarsely crystalline and forms units up to 1 m in thickness. Individual units are separated by black graphitic semipelite and pelite. The Bonds Glen Limestone Member is pale to dark bluish grey, medium- to coarse-grained, in beds up to 0.75 m in thickness. The beds are separated by dark grey to black pelite that is often pyrite-bearing.

Dart Formation

³The Dart Formation is present in the south-eastern quadrant of the licence area. The base of the formation is composed Glenga Amphibolite Member, which is interpreted as a re- sedimented volcaniclastic siltstone and sandstone. Other lithologies within the formation include quartz pebble conglomerate, psammite displaying graded bedding, semipelite and the +HQU\¶V%ULGJH0HPEHU7KH+HQU\¶V%ULGJH0HPEHULVEHWZHHQWRPLQWKLFNQHVVDQG is composed of a dark greenish brown, chlorite-rich volcaniclastic semipelite and is interpreted to record erosion of underlying volcanics, but may also result from contemporaneous volcanism in the hinterland (Strachan et al., 2002). Throughout the Dalradian volcaniclastic VHPLSHOLWHVDUHNQRZQDV³*UHHQ%HGV´

Glenelly Formation

Glengawna Formation

³The Glengawna Formation is restricted to the southeast of the DG3 licence area, forming the low ground in the Gortin Glen Forest Park (GR 249000 381000). The formation is approximately 400 m thick and comprises a distinctive association of graphitic pelites, semi- pelites, chloritic semi-pelites and psammites. These lithologies contrast markedly with the silvery-grey lithologies of the underlying Glenelly Formation.

Mullaghcarn Formation

³The formation is only present in the extreme southeast and south of the licence area, and it consists of psammite, semi-pelite, chloritic-semipelite and pelite schists. Garnets are sometimes present, and are usually replaced by chlorite or hematite. Bedding is difficult to recognise, but it is generally acknowledged that the formation is overturned. The Mullaghcarn is economically significant, since both the Curraghinalt and Cavanacaw gold deposits are hosted in this formation.

Carboniferous: Late Courceyan and Early Chadian

³Two Carboniferous basins are present within the licence area. In the south, the Omagh Sandstone Group represents strata deposited prior to, and during, and marine transgression, whereas in the north the Owenkillew Sandstone Group in the Newtownstewart Outlier represents non-marine strata.

Omagh Sandstone G roup

³The Omagh Sandstone Group rests unconformably on the Dalradian. The basal unit is up to 100 m thick and is composed of non-fossiliferous red sandstone with calcrete nodules, and quartz pebble conglomerates (Mitchell, 2004). Much of the remaining sequence is dominated by channel sandstone and siltstone that contain Courceyan to early Chadian miopores. However, thin algal limestones with evaporite replacement textures occur locally. Some of the limestones contain rare brachiopods. The exact thickness of this group is difficult to estimate based on the amount of uplift, folding, and erosion that has taken place.

Owenkillew Sandstone Group

³The Owenkillew Sandstone Group rests unconformably on the Dalradian and comprises of approximately 1500 m of predominantly non-marine strata present within a half graben. Lithologies include greenish grey and purplish red sandstone and siltstone, with thin beds of algal laminated limestone (Mitchell, 2004). Mudstones containing miospores have indicated an early Chadian age. The group is thought to have formed in an inter-cratonic basin with FXUUHQWLQGLFDWRUVVXJJHVWLQJWKHVHGLPHQWVRXUFHWRWKHQRUWK´

7.2.3.2 Structural History and Setting of Gold Mineralization

Williams (2009) described the structural history and setting of gold mineralization on the DG3 and DG4 licence areas as follows:

³.QRZQJROGGHSRVLWVZLWKLQWKH'DOUDGLDQ6XSHUJURXSHJ&XUUDJKLQDOWDQG&DYDQDFDZOLH on the inverted, gently to moderately northwest-dipping, limb of a southeast facing, major overturned tight to isoclinal fold (Alsop and Hutton; 1993a, 1993b). This fold is known as the Sperrin Fold in the Sperrin Mountains, the Altmore Anticline in northeast Co. Antrim (Alsop and Hutton, 1993a; Cooper and Johnston, 2004). All of these structures are analogous to the Tay Nappe in Scotland, that formed as a result of D2 flattening, tightening and extension of an earlier D1 anticline (Stephenson and Gould, 1995; Strachan et al., 2002). Deformation resulted in lower greenschist facies in the north (Co. Donegal) to lower amphibolite facies metamorphism close to the Omagh Thrust. D3 deformation in the south Sperrin Mountains resulted in a south-easterly-verging minor folds and low angle, northwards inclined thrust surfaces, i.e., the Omagh Thrust Fault, which transported Dalradian rocks over the early Ordovician Tyrone Igneous Complex (Alsop and Hutton, 1993a). The dominant D3 structural feature in west County Tyrone is the Ballybofey Nappe, which is a sub-recumbent structure that closes and faces towards the southeast and is displaced by several Caledonian faults (Alsop, 1991).

³Auriferous quartz-carbonate veins occur throughout the Dalradian Supergroup, with the majority (e.g., Curraghinalt, Golan Burn, Cavanacaw) present close to the Omagh Thrust, where the Dalradian rocks have been overthrust to the southeast over the Tyrone Igneous Complex. The presence of this terrane boundary obviously plays a significant role in the formation of economic mineralization. Gold tends to occur as electrum, gold-bearing hessite (Ag2Te) and as substitutions in pyrite. Associated sulphides in the veins include pyrite, chalcopyrite, arsenopyrite, tennantite, galena and sphalerite. Gangue minerals are quartz, calcite, dolomite, siderite, orthoclase and rare barite. All veins are steeply dipping, but their strike varies from north-south at Cavanacaw in the west, to west-northwest in the east at &XUUDJKLQDOW´

A simplified geological map showing the relationship of the major units of the Dalradian Metasediments, Tyrone Volcanic Group, Central Inlier and the Tyrone Plutonic group can be seen in Figure 7.5.

7.3 MINERALIZATION

7.3.1 The Curraghinalt deposit

The mineralization at the Curraghinalt gold deposit is described by Tully (2005) as follows:

³0LQHUDORJLFDOO\WKHYHLQVDUHFRPSRVHGRITXDUW]ZLWKFDUERQDWHRQODWHFURVVYHLQV3\ULWH and chalcopyrite are the dominant sulphides with accessory tennantite and tetrahedrite plus rare arsenopyrite, galena and molybdenite. Three styles of sulphide mineralization are evident at Curraghinalt; (i) euhedral pyrite crystals with chalcopyrite occur in white quartz, filling vugs or along micro-fractures, (ii) grey pyritic quartz forms the matrix to brecciated white quartz, and (iii) massive pyrite.

³7KH YDULDWLRQV LQ WKH VXOSKLGH FRQWHQW DUH GHSHQGHQW RQ WKH YROXPH RI VSDFH FUHDWHG E\ movements on the E-W shears. The massive sulphide mineralization present at the eastern end of the T17HW drift [Ennex underground exploration program] is due to its position adjacent to the Kiln Shear. On the No. 1 Vein, drill holes 106-16, 90-12, 90-14 and 90-43 exhibit increased mineralization close to the Kiln Shear. Zones of grey pyritic quartz breccia commonly occur along the hangingwall contact of the veins. This is evident in the No. 1 drift. The veins have sharp margins with no mineralization in the country rock. Sericitization and minor chloritization occurs in the hangingwall and footwall of some of the veins, but it is not FRQVLVWHQWO\SUHVHQW´

The gold at Curraghinalt occurs in three forms:

As inclusions within pyrite, remote from fractures. Within fractures in pyrite. As gold-bearing tellurides.

Four stages of quartz deposition (Q1 through Q4) were recognized, with gold mineralization associated with the second and fourth stages, as described by Tully (2005):

³4UHSUHVHQWVWKHLQLWLDOSKDVHRIYHLQIRUPDWLRQDQGKDVDQRSDTXHPLON\ZKLWHDSSHDUDQFH No sulphides are present, and it is interpreted that the quartz precipitated from a metamorphic fluid at greenschist facies temperature.

Q2 is a re-activation phase of the vein system. This resulted in intense brecciation of the existing quartz (Q1), cementation by new quartz (Q2) and the development of significant porosity of the veins. Q2 is recognized by its grey, translucent appearance and its breccia cement texture. Textural relationships indicate that Q2 was synchronous with the main phase of sulphide mineralization. Fluid inclusion and staple isotope geochemical studies have indicated that Q2 was precipitated from a magmatic fluid or from a fluid with a significant magmatic component.

Q3 is limited in occurrence to the larger, most intensely mineralized veins. It can be recognized in hand specimens as coarse, transparent euhedral crystals. The Q3 fluid is interpreted to represent evolved surface-derived water.

Figure 7.5 Simplified Geological Map of the Project Area

Golan Burn Omagh Thrust Fault Curraghinalt Deposit 75

Alwories

Source: Archibald and Williams, 2009 from Geological Survey of Northern Ireland.

Q4 is only common in the thickest mineralized veins. It occurs as cement to anhedral/brecciated aggregates of pyrite and as thin syntaxial overgrowths on euhedral crystals. Q4 is synchronous with the late generation of sulphide and electrum mineralization infilling fractures in pyrite. 6WXGLHVLQGLFDWHWKDW4ZDVSUHFLSLWDWHGIURPEDVLQDOEULQH´

The following description of the vein system at Curraghinalt is reproduced from Tully (2005) who references Boland (1997).

³7KH TXDUW] YHLQV IRUP D VHULHV RI VXE-parallel structures, which crosscut the Dalradian metasediments with a general strike of 090º - 135º, and dip at 55º - 77º to the north. They bifurcate, splay, and pinch and swell along strike and down dip and vary in thickness from less than 5 cm to greater than 2 m. The veins are well developed in the semi-pelites and psammites but have poor continuity in the package of ductile pelites and chloritic semi- pelites.´

The vein swarm has been traced along strike for approximately 1,950 m, across strike for approximately 600 m and down dip for approximately 800 m by prospecting, trenching and GULOOLQJ7HQµPDLQ¶YHLQVDUHLQFOXGHGLQWKHFXUUHQt resource estimate. These are:

1. Road 2. Sheep Dip 3. Mullan 4. T17 5. No. 1 6. 106-16 7. V75 8. Bend 9. Crow 10. Attagh Bridge Burn

A brief description of the veins discovered to date is as follows:

Road Vein - The Road vein was encountered initially in a trench and subsequently by drilling and underground development. It has been traced by drilling over an approximate 500 m strike length and to a depth of approximately 300 m below surface. The vein remains open to depth and to the east and west.

Sheep Dip Vein - The vein lies approximately 60 m south of the Road vein and has been accessed by a short drift from underground, near the portal of the adit and has a defined strike length of 200 m. The vein strikes 110º to 115º and dips between 60º to 65º to the north. It has been traced along strike for approximately 500 m and to a depth of approximately 300 m. The vein remains open to depth and to the east and west.

Mullan Vein - The vein occurs approximately 70 m south of the Sheep Dip vein and has an interpreted minimum strike length of approximately 1,050 m. Drilling has intercepted the vein at approximately 440 m below surface and it has a similar strike and dip to the Sheep Dip

vein. The Mullan Vein is not well developed at the elevation of the adit, but remains open along strike and down dip.

T17 Vein - The vein is located on both the southern and northern sides of the Kiln Shear, approximately 100 m south of the Mullan Vein. It intersects the east-west trending Kiln Shear where it is visible in the T17 drift. The vein strikes west-northwest and dips 60º to 65º to the north. It is interpreted to have a minimum strike length of about 1,400 m and has been delineated by 200 m of drifting. Drilling has tested the vein to an elevation of -350 m (c. 570 m below surface). The vein is well developed, with good continuity in the semi-pelitic lithologies, which were intersected in the eastern 110 m of the T17 drift. The vein is offset by 1 to 2 m in a number of places by northeast-trending cross faults. West of 256980E the lithology is dominantly a pelite and the vein is less well developed. The vein remains open to depth and to the east and west.

No. 1 Vein - The vein is located 90 m south of T17 Vein and strikes 110º to 115º and dips between 60º to 65º to the north. It has been tested by several trenches, numerous drill holes and 95 m of drift, plus one raise. The vein has been interpreted along strike for approximately 1,100 m and down dip for some 680 m. The vein remains open to depth and along strike to the east and west.

106-16 Vein - The vein is located approximately 80 m south of the No. 1 Vein. It strikes 110º to 115º and has a variable dip of between 50º to 65º north. The vein has been interpreted for a length of approximately 1,300 m along strike, and to a depth of approximately 710 m below VXUIDFH7KHYHLQUHPDLQVRSHQWRGHSWKDQGWRWKHHDVWDQGZHVW´

V75 Vein - The vein is located approximately 60 m south of the 106-16 vein and strikes 110º to 115º and has a variable dip of between 50º to 65º north. It has only been encountered in drill holes and has been correlated over an interpreted strike length of 1,100 m and traced down dip for approximately 760 m. The vein remains open to depth and to the east and west.

Bend Vein - The Bend vein is located approximately 60 m to the south of the V75 vein and has been interpreted over a strike length of approximately 1,100 m and down dip for approximately 800 m through trenching and drilling. The vein remains open to depth and to the east and west.

C row Vein - The Crow vein is located approximately 70 m to the south of the Bend vein and has been interpreted over a strike length of approximately 1,100 m and down dip for approximately 840 m through trenching and drilling. The vein remains open to depth and to the east and west.

Attagh Burn Bridge Vein (A BB) - This is the westernmost vein drill tested within the deposit. The vein may represent the western extension of the 106-16 Vein, an offset of the No. 1 Vein, or is a separate entity. It strikes 110º to 115º and dips 60º to 65º to the north and has been tested by drilling and surface trenching for 260 m. It has been tested by four surface trenches and 22 drill holes to a depth of 130 m below surface. The vein remains open to

depth and to the west. Drill hole CT-26 has extended the known depth of the vein to 215 m below surface.

7.3.2 The Tyrone Volcanic Group

Williams (personal communication, December, 2009) has provided the following list of anomalous areas and/or showings in the Tyrone Volcanic Group (TVG) on the DG2 licence. Their locations relative to the DG2 boundaries are shown in Figure 7.6.

³$ QXPEHU RI DQRPDORXV DUHDV KDYH EHHQ GHIined by the exploration completed to date. These include from NE to SW:

1. Tullybrick (273250E, 390000N) The area contains bleached and sericitized tuffs with reported minor fuchsite over several hundreds of metres of section. These contain disseminated pyrite grading up to 5% and a maximum gold value of 2.78 g/t Au.

2. Bonnety Bush (272500E, 388750N) Much of the area is moorland with limited rock exposure. However, there is good exposure of the jasper, magnetite ferruginous chert that marks the top of the Cycle 2, Bonnety Bush-Glencurry Formation. The main target area is located to the north of this. However, there are a number of targets south of the top of the Cycle 2 volcanics that were highlighted by the MPM study that warrant investigation.

3. Crocknahala (Broughderg) (266600E, 387000N) This northeasterly trending area may be on the rhyolite horizon, no outcrop was seen. The presence of much tonalite float, some of it sheared, may indicate that the tonalite extends further northeast than shown on published maps. Trenching to test anomalous geochemistry intersected 1.5m @ 4.36 g/t Au in tectonized tuffs.

4. Aghascrebagh (261000E, 383400N) to Greencastle (258250E, 382200N) The area is between the tonalite and the Omagh Thrust and contains altered tuffs with disseminated pyrite-galena-sphalerite-chalcopyrite mineralization. Drilling in the area has intersected siliceous tuffs with maximum values of 0.5m @ 2% Pb/Zn and 0.5 g/t Au.

5. The Formil (Cashel) Copper Prospect (262000E, 382200N) Exploration by Riofinex in the 1970s established the presence of low-grade copper porphyry-type mineralization at Formil (Cashel) and additional low-grade copper mineralization was intersected further west.

At Formil, drilling by Ennex was in quartz porphyry within an area of argillic alteration. At the contiguous Slievemenagh copper occurrence drillhole H-7 (Riofinex) intersected northwesterly dipping mineralization, 94.2m @ 0.16% Cu, including 1.83m @ 0.32% Cu, in banded mesocratic, dark, tuffs cut by small acid and intermediate intrusions. The tuffs are silicified near the intrusions. Mineralization reportedly occurs as veinlets and blobs in quartz veins or as disseminations in the surrounding rock.

About 3km west-southwest of Formil, Riofinex intersected further low copper values. Here drillhole H-2 intersected 39.6m @ 0.17% Cu in tuffs with coincident copper soil anomalies, drillhole H-4 intersected 46m @ 0.12% Cu in tuffs and drillhole H-6 intersected 24m @ 0.17% Cu in tuffs.

Figure 7.6 Tyrone Volcanic Group Mineralized Locations

DG1

DG2

Source: Williams 2009, personal communication.

6. The Cashel-Leaghan (Cashel Rock) Gold Prospect (258950E, 380100N) In the 1980s Ennex discovered and drilled the Cashel-Leaghan (Cashel Rock) gold prospect. The prospect was discovered by geochemistry/prospecting and tested by trenching/drilling.

Significant gold was outlined by Ennex (max. 3.63m @ 30.5 g/t Au, including a 1.23m interval @ 1.14% Cu and 1.85% Pb), along a 100m zone of silica flooding within rhyolite flows. The silica zone, about 2m thick, dips west at 30º and may occur along a flow top or alternatively at the contact between a felsic intrusion and a rhyolite flow.

7. Murrins (256000E, 377500N) This area is located to the SW of the Cashel-Leaghan gold prospect and occurs in an area with extensive gravel development with a number of operating sand and gravel pits and no bedrock exposure. No rhyolite is mapped as occurring beneath the sands/gravel but rhyolite may exist here on the northern side of the tonalite and this would enhance the potential of the target area. A further positive aspect is that if the tonalite, which dilates the section, is abstracted it brings the Murrins area into close depositional proximity with the Loughmacrory quartz porphyry.

8. Mountfield (253600E, 378000N) to Glencurry (251000E, 377000N). The volcanics in this area are located between the rhyolite and the graphitic pelite that occurs along the Omagh

Thrust. The bedrock exposure in the area is very limited but prospecting in the area in the 1980s located tuff float with disseminated sphalerite.

9. Loughmacrory (259000E, 377300N). This target area is located at the top of Cycle 1 and straddles the area between the Loughmacrory quartz porphyry and the tonalite to the north. A large boulder of rhyolite volcaniclastic occurs in a dry stone wall of a laneway, near the foot of the hill. There is a suggestion that the quartz porphyry may have vented onto the sea floor in this vicinity, thereby creating a favourable environment for the deposition of massive sulphide deposits. If this is so, it makes areas south of the tonalite more prospective than ZRXOGRWKHUZLVHEHWKHFDVH´

Beaghbeg (see Figure 7.6) is another important anomalous area in the TVG. At this location surface jasperoid and rhyolite breccia outcrops have been identified. Soil sampling by Ennex has identified soil geochemical anomalies with anomalous levels of copper, lead, zinc and gold.

Crosh (see Figure 7.6) is a possible gold exhalite target where silicified volcanics and black shale gold mineralization have been found. Grab samples in siliceous tuff have returned gold values of 0.57, 0.42 and 0.37 g/t. Soil sampling by Ennex has also discovered anomalous levels of copper, lead, zinc and gold.

Although the TVG has been recognized as a terrane suitable for the possible discovery of VMS-style mineralization, and some supporting evidence has been found, no massive sulphide mineralization has been discovered to date. Sulphides of volcanic origin have been found though. Analogies with the TVG have been drawn above to the host rocks of the Buchans, Newfoundland area. The following description of Buchans type ores is taken from Thurlow and Swanson (1981).

³7KH%XFKDQVJURXSLVDUHJLRQDOO\extensive calc-alkaline volcanic suite erupted at a point in time near the Ordovician-Silurian boundary. It is subdivided in Lower and Upper Subgroups, each of which consists of interbedded mafic to felsic pyroclastics, flows and breccias.

³The Buchans ores are high grade Kuroko-like polymetallic massive sulphide deposits associated with submarine felsic volcanic rocks near the top of the Lower Buchans Subgroup. Three genetically related ore types occur: stockwork ore, in situ ore and mechanically transported fragmental ore. Stockwork ore consists of networks of veinlet and disseminated base metal sulphides within highly silicified and locally chloritized host rocks. In situ ore lies stratigraphically above stockwork mineralization and is composed of banded to streaky yellow ore, black ore and barite. Transported ore consists of discrete orebodies composed of sulphide and barite fragments which were derived from in situ ore, transported by density flows and deposited in paleotopographic depressions. They maintain ore grade at distances greater than 2 km from the source area, well beyond the limits of stockwork mineralization and related alteration.

³The Buchans ores are outstanding among massive volcanogenic sulphides because of their very high grade DQGWKHH[WHQVLYHGHYHORSPHQWRIPHFKDQLFDOO\WUDQVSRUWHGIUDJPHQWDORUHV´

Thurlow and Swanson (1981) report that 17,426,000 tons of ore grading 14.62% Zn, 7.60% Pb, 1.34% Cu 3.69 oz/ton Ag and 0.043 oz/ton Au had been mined from five major orebodies in the Buchans area. Strong (1981) reports that ores commonly contain pyrite, chalcopyrite, sphalerite, galena, quartz, barite and calcite. The sphalerite is very low in iron (<1%) and the silver occurs essentially in tetrahedrite and not galena.

8.0 D EPOSI T T YPES

Tully (2005) described the types of mineral deposits that have been identified on DG1 and DG2 (identified as UM-1/02 and UM-2/02 in 2005):

³2URJHQLFVWUXFWXUDOO\FRQWUROOHGPHVRWKHUPDOJROG-bearing quartz and quartz-sulphide veins in Dalradian clastic meta-sedimentary rocks - e.g., the Curraghinalt deposit.

³Possible stratabound and perhaps stratiform semi-massive and massive bands of base metal (lead, zinc and pyrite) sulphides with significant gold values in fine grained Dalradian clastic meta-sedimentary strata - e.g., the Glenlark occurrence.

³3RUSK\U\´-style bodies in altered intermediate intrusive rocks within the Tyrone Volcanic Group. Sulphide veinlet stockworks contain pyrite and chalcopyrite, and some quartz veins host rare galena and sphalerite. Molybdenite occurs on shear surfaces - e.g., the Formil occurrence.

³Siliceous breccias with gold values and base metal sulphides in felsic units of the Tyrone Volcanic Group - e.g., WKH&DVKHO5RFNRFFXUUHQFH´

8.1 THE CURRAGHINALT DEPOSIT

The Curraghinalt deposit, an orogenic structurally-controlled, mesothermal gold deposit is comprised of a swarm of gold-bearing quartz veins within a meta-sedimentary sequence of pelites, semi-pelites and psammites. The veins range from a few centimetres wide to over 3 m wide and are aligned along a west-northwest trend. Average width is approximately 1.2 m. The strike length of the vein swarm at Curraghinalt has been traced by prospecting, trenching and drilling to a minimum length of 1,950 m and remains open to the east and west. The veins included in the present resource occur over a width of approximately 600 m from north to south and more veins are known to occur to both the north and south. The veins dip between 60º and 75º to the north and depth has been traced to approximately 840 m below surface The vein system remains open to depth.

As noted above, the structures are interpreted to be related to high-angle, east-west- to west- northwest-trending shears which are probably related to the thrusting of the Dalradian metasediments over the Tyrone Igneous Complex.

A technical report prepared for the DETI and Crown Mineral Agent by Aurum (Williams, 2008) describes the veins as follows:

³7KHYHLQVDUHGRPLQDQWO\TXDUW]ZLWKVRPHFDUERQDWHV/RFDOO\abundant sulphides are mostly pyrite and commonly contain variable amounts of other accessory minerals [sic], notably chalcopyrite, tetrahedrite and tennantite, and trace amounts of galena, sphalerite, molybdenite and several telluride minerals present. Gold is reported to RFFXUDVHOHFWUXPLQXQVSHFLILHGWHOOXULGHPLQHUDOVDQGZLWKLQWKHS\ULWH´

8.2 THE TYRONE IGNEOUS COMPLEX

The Tyrone Igneous Complex (TIC) comprises the Tyrone Plutonic Group and the Tyrone Volcanic Group (TVG). The TIC is known to host both gold and porphyry copper style mineralization and is interpreted to have good potential to host VMS style mineralization similar to the Buchans type.

Williams and Archibald (personal communication via e-mail, December, 2009) observe the following:

³7KHPDLQSRWHQWLDOIRU906PLQHUDOL]DWLRQLVLQ&\FOH>RIWKH79*@DQGH[WHQGVDORQJD- 3 km wide corridor on the southern side of the Omagh Thrust. Southwest from Crocknahala (267000E, 387000N) the volcanics are intruded [by] tonalites and quartz porphyries with associated rhyolites. It is likely that the rhyolites trending southwest from the Slievemenagh area (Formil) (262500E, 382700N) continue under glacial cover along the northerly side of the tonalite towards Mountfield and then continue on either side of the tonalite for some distance to the southwest. The rhyolite may be assumed to continue southwest from the Cashel- Leaghan gold prospect along the northern flank of the tonalite.

³With regard to the licensed area, the most prospective VMS locations would be along strike from these central areas, both north and south of the tonalite. The northerly striking Dungate fault makes the volcanic rocks on its northeast side towards Tullybrick a favourable locus for 906GHSRVLWV´

As described above, there is also potential for porphyry-style copper mineralization and gold- mineralized silicified breccias. The potential target types in the TVG, on licence DG2, are summarized in Figure 8.1.

Ongoing research by Hollis has indicated that the newly defined Greencastle Formation (equivalent to Cycle 3 quoted by Williams and Archibald) is the most likely correlation with the Annieopsquotch Group in Newfoundland that is host to the Buchans VMS deposits. Hollis bases this on detailed geochemistry and age dating (personal communication with G. Earls of Dalradian Resources, 2011).

Figure 8.1 Mineralization Models and Target Types on DG1

After Clifford et al. (1992).

9.0 E XPL O R A T I O N

The exploration programs carried out by Ennex between 1983 and 1997 have been described in Section 6, above. Between 1997 and 2002, only a small amount of exploration was undertaken on the property during a period of relatively low metal prices and poor markets.

Following the option agreement of 2003 between Tournigan and Strongbow, Tournigan carried out geochemical and geophysical surveys, mapping and prospecting and a 26-hole diamond drilling program.

Since 2010, Dalradian Resources has completed 100 drill holes at Curraghinalt (DG1) and 16 regional drill holes. In addition airborne and ground geophysical surveys, prospecting, mapping and geochemical surveys have been completed. Drilling is discussed in Section 10 below.

9.1 NICKELODEON/NAVIGATOR/STRONGBOW 1997 - 2003

Nickelodeon focused on validating drill core and underground channel sample gold assays. A program of quarter drill core sampling and analysis successfully validated previous sampling and analyses. Several channel samples were sawn on the back of the T17 and No. 1 drifts in lower, medium and higher grade sections. The results from this program compared very closely with the existing samples collected during underground development in the late 1980s.

Strongbow focused exploration on two areas, Glenlark in DG1 and Crosh in DG2. Mobile Metal Ion geochemistry was completed in both areas (Glenlark 236 sites; Crosh 308 sites) and defined coherent precious and base metal anomalies. An induced polarization (IP) geophysical survey at Glenlark identified a series of conductors.

The mineralization encountered in the first trench (T10) comprised stratiform and crosscutting zones of massive and semi-massive zinc, lead and iron sulphides. These zones were also associated with significant gold and silver values (tabulated in Table 9.1 below

Table 9.1 Selected Results from Trench T10, Glenlark, Sampled Widths

F rom To Zn Pb Au Ag (m) (m) (%) (%) (g/t) (g/t) 98.5 99.5 8.03 2.87 1.27 19.6 99.5 100.5 10.69 0.36 6.88 36.2 100.5 101.5 0.40 0.06 0.10 1.3 101.5 102.5 5.05 0.10 5.74 16.2 102.5 104.0 0.08 0.13 0.10 1.4 104.0 106.0 3.74 0.96 3.91 27.6 106.0 108.0 7.36 0.83 6.98 41.7

The sampled interval from 98.5 m to 108.0 m (9.5 m) assayed 4.89% Zn, 0.75% Pb, 3.78 g/t Au and 22.53 g/t Ag.

The second trench (T11) encountered very deep overburden conditions and bedrock was not encountered over much of its length. Several angular mineralized boulders, which are interpreted to be related to a nearby source, were encountered.

In addition prospecting was undertaken at Greencastle and Carnanransy in DG2.

9.2 TOURNIGAN EXPLORATION, 2003 - JANUARY, 2005

7KHIROORZLQJGHVFULSWLRQRI7RXUQLJDQ¶VH[SORUDWLRQZRUNLVWDNHQIURP7XOO\

³7RXUQLJDQVWDUWHGH[SORULQJWKHSURSHUW\LQ$SULOand has completed geochemical and geophysical surveys, plus mapping and prospecting programs as well as a 26-hole diamond drilling program on Curraghinalt.

Geochemical Sampling

³7RXUQLJDQFROOHFWHGVRLOVDPSOHVIURPQRUWKVRXWKRULHQWHGJULGOLQHVon 100m centres along the Curraghinalt trend. Samples were collected at a depth of 50-75 cm, roughly the B- soil horizon, on 30m spacing along the lines. The area covered included the area to the north of the known vein structures, as well as in both directions along strike from the known veins.

³Many anomalous areas were identified in the sampled areas, indicating the continuity of the vein structures along strike as well as the possible presence of additional veins to the north of the known veins. The main targets identified cover several areas north, northwest and northeast of the adit that combined, measures 1.5-km north to south and which extends for approximately 1.0-km along strike. This area represents 419 sample points with values ranging from 400 - 12300 ppb Au, plus 5 smaller areas with similar values. These areas outside the immediate area of the known veins will require follow-up deep overburden sampling plus diamond drilling to develop the individual vein structures.

Geophysics

³$QDWWHPpt to conduct a VLF survey over the grid area was eventually abandoned when the Cuttler Maine transmitter station was shut down. A total of 652 readings taken at 30m spacing along 100m spaced lines were completed prior to the shut down of the Maine station. The results were inconclusive. The Rugby station, located in the UK was never turned on during the time of the survey.

Geology

³7RXUQLJDQKDVFRPSOHWHGSUHOLPLQDU\PDSSLQJDQGSURVSHFWLQJWKURXJKRXWWKHOLFHQVHDUHD One of the discoveries resulting from this work is the Alwories Vein, located 2.5 km southeast of Curraghinalt. Previous work in this area had identified float samples of vein material only. Vein outcrop was mapped in a newly developed rock quarry. Hand stripping subsequently revealed a 50-cm wide vein that averaged 15 g/t Au. The Alwories vein is a quartz-sulfide vein very similar to the veins that comprise the Curraghinalt deposit. Future work in this area will include detailed stripping and sampling, detailed mapping and possibly follow-up GLDPRQGGULOOLQJ´

7RXUQLJDQ¶VH[SORUDWLRQZRUNRQWKHSURSHUW\VWDUWLQJLQ$SULOZDVFDUULHGRXWE\LWV subsidiary, Dalradian Gold. Aurum, based in Kells, County Meath, Ireland, provided local project management services and technical staffing for the project.

Tully reported that Tournigan completed 33 HQ drill holes between 2003 and January, 2005. Of these, seven were drilled in the Glenlark area and 26 infill holes were drilled in the Curraghinalt deposit.

Infill drilling, holes 03-CT-05 to 04-CT-26, was completed on the central resource block.

9.3 TOURNIGAN EXPLORATION FROM JANUARY, 2005

The drilling programs carried out by Tournigan since the preparation of Tully (2005) are described in Section 11 below.

7RXUQLJDQ¶VH[SORUDWLRQSUogram up to the time of writing of the Tully report is described in Tully (2005). Tully concluded:

³7KHZRUNFRPSOHWHGE\7RXUQLJDQWRGDWH>LHWRWKHGDWHRI7XOO\¶VUHSRUWLQ-DQXDU\@ has consisted of 26 diamond drill holes (4,391m) on the Curraghinalt deposit, 7 holes [drilled] on the Glenlark project area (NE of Curraghinalt) for a total 830.5m, shallow depth geochemical surveys, detailed compilation work, plus one trench (228m) and a satellite LPDJHU\DQDO\VLVRIWKHDUHD´

In 2005, Tournigan completed two diamond drill holes in the area of the Crows Foot-Bend structural zone (05-CT-27 and 05-CT-28).

In 2006 to 2007, a 25-hole program of infill drilling was completed on the Southeast Extension target.

A program of underground structural mapping and sampling was undertaken by consultant structural geologist, John Cuthill in 2002-2003. The T17 and No. 1 drifts were the principal areas of emphasis. Tully (2005) reported on the structural analysis as follows:

³7KHGDWDVXJJHVWVWKDWIDXOWLng has played a significant role in ground preparation within the vein structures. The faulting appears to have facilitated the injection of the Q4 style sulphide rich quartz veining material in the veins where they are in close contact with the Kiln Shear. The high-grade nature of most of the drill holes piercing the T17HW vein in the area where the vein intersects the Kiln Shear suggests a high-grade shoot in this area.

³Similarly, the apparent low-grade sections occurring within the plane of the veins can be shown in some cases to be associated with the pelite horizons that tended to deform plastically, rather than in brittle fashion. The presence of the pelitic units has resulted in rather ³WLJKW´URFNFRQGLWLRQVWKDWKDYHLQKLELWHGWKHLQMHFWLRQRIquartz veining. Within these low- grade horizons mostly barren to low grade Q1 quartz has formed. The nature of the enclosing rock has prevented the formation of Q2 and Q4 gold bearing veining. Consequently the quartz

veining within these semi-pelite units tend to contain only narrow, Q1 relatively barren quartz. Since stratigraphy strikes NE, and dips at a moderate angle across the NNW striking vein sets, the low grade intervals are created by the pelitic units where such units intersect the strike of tKHYHLQV´

The location of holes drilled by Tournigan between 2003 and 2007, and the adit developed by Ennex, are shown on Figure 9.1.

Figure 9.1 Location of Holes Drilled by Tournigan

53 Holes

26 SE Extension Extension

Map provided by Tournigan. Date unknown.

Data from trench samples had not been digitized and precise sample locations had not been mapped. No data from trench samples were used in the previous mineral resource estimates.

Tournigan planned to collate the data from trench sampling but, WR0LFRQ¶VNQRZOHGJH, this was not completed.

9.4 TOURNIGAN EXPLORATION 2007 TO 2009

After completion of the 2007 Micon resource estimate for Tournigan (Mukhopadhyay, 2007), the company completed several more drill holes at the Curraghinalt deposit. The results of that work are discussed in Section 11 below. After cessation of drilling at Curraghinalt Tournigan had Aurum engage in a small program of data compilation, prospecting and sampling on the DG3 and DG4 licence areas. Williams (2009) describes the work completed as follows:

Geological Data - Geology, Prospecting and Sampling

³All available historical data (which includes geological mapping, field prospecting and geochemistry) was collected and catalogued during the earlier phases of work. This data was re-assessed and new data, where possible, was added to the database. This data included outcrop data, prospecting data, and general geological data.

³Historical prospecting had previously identified several areas where significant auriferous quartz vein material was evident either in bedrock or float. The results ranged between 15- 30g/t Au in the Rylagh area, >100g/t in the SW corner of TG-3 (now DG 3) at Bessy Bell, and 187g/t Au in Glensass Burn on TG-4 (now DG 4).

³Satellite imagery completed by Tournigan in 2004 identified structural trends on TG-3 (DG 3) and TG-4 (DG 4) which are very strongly analogous to those at Curraghinalt, and along with the anomalous geochemistry that characterizes the area, several important target areas have emerged.

Geochemistry - Stream Sediments \ Panned Concentrates

³All existing geochemical data was compiled for the TG-3 (DG 3) and TG-4 (DG 4) licences and this included panned concentrate data collected historically by Ennex, and also British Geological Survey (BGS) stream sediment data. This data was compiled and re-evaluated in the target generation.

³Ennex also carried out detailed panned concentrate surveys on TG-4 [DG4, the licence bordering DG1 to the north], in an area around Plumbridge and Eden Hill. Several very anomalous streams had been identified with values up to 97-105 ppm Au in panned concentrates. No equivalent data sets for historical panned concentrate have so far been located for TG-3.

Geophysics

³Since the last renewal period, the Geological Survey of Northern Ireland released the Tellus Airborne Geophysical Survey Data for Northern Ireland. This data was purchased [technically Tellus data is leased] and work was carried out in-house by Aurum Exploration Services using this Geophysical database to interpret geological features including faults, conductive units and specific anomalous targets.

Geological Appraisal

³Licences TG-3 and TG-4 [DG3 and DG4] have a significant potential to host Curraghinalt style mineralization as well as several other distinct styles, most of the licence area is underlain by Dalradian metasediments interbedded with volcanics and intrusive lithologies. Structurally there are some major features known both historically and through more recent satellite imagery work and also through interpretive use of the Tellus airborne geophysical data. There have been several reported float and outcrop occurrences of gold mineralization and also the high level of anomalism for gold within the panned concentrate surveys carried RXWE\(QQH[LQWKH¶V

³Work during the last reporting period 2005 - 2007 concentrated primarily on the stream sediment geochemical data set and detailed basin analysis work was carried out on the anomalous drainage basins. Limited follow up shallow soil geochemistry failed to identify a mineralizing structures associated with these very high values and though most evidence has started to point to a glacially derived origin for this alluvial gold. A buried bedrock source however, VWLOOFDQQRWEHUXOHGRXW´

9.5 DALRADIAN RESOURCES EXPLORATION

Since acquiring the Tyrone project, Dalradian Resources, in addition to the drill program at Curraghinalt, began a program of exploration on others areas of its licenses. After carrying out a review of the Northern Ireland Government Tellus geophysical data, regional prospecting and target generation was undertaken.

Prospecting was carried out on all four licences in the first and second quarters of 2011. A total of 929 samples were collected, 143 of which assayed greater than 0.25 g/t Au. A summary of the samples is provided in Table 9.2. The locations of the prospecting samples are shown in Figure 9.2.

Table 9.2 Summary of 2011 Prospecting Samples

License Total Outcrop Float Gold Value Significance A rea Samples Samples Samples Range (g/t Au) DG1 316 139 177 0.01 - 44.96 88 samples over 0.25 g/t Au DG2 270 144 126 0.01 - 5.48 31 samples over 0.25 g/t Au DG3 184 86 98 0.01 - 14.08 22 samples over 0.25 g/t Au DG4 159 97 62 0.01 - 2.07 2 samples over 0.25 g/t Au

Integration and re-evaluation of previous exploration and the new regional prospecting information identified 19 areas for detailed follow-up and drill testing. These are shown in Figure 9.3.

These targets can be split into two distinct groups, those within the igneous TVG, mainly located within licence DG2, and those within the Dalradian Supergroup (DSG) of predominantly metasediments in DG1, DG3 and DG4.

The TVG is an environment favourable for the formation of VMS deposits. It is interpreted to have a similar origin as the Buchans VMS camp in Newfoundland and contains an abundance of float with VMS-style mineralization. Gold and base metal mineralization is most prevalent within the upper part of the volcanic sequence. Historic prospecting results include siliceous tuff assaying 16.1% Pb and 1.5 g/t Au and chloritic tuff assaying 8.8% Zn, 1.2% Pb and 1.7 g/t Au.

Figure 9.2 Location of Prospecting Samples, 2011

Source: Dalradian Resources, 2011.

Figure 9.3 Exploration Targets

Source: Dalradian Resources, 2011.

In the DSG, the highest priority exploration targets are within the Curraghinalt structural corridor, an area of approximately 8 km of strike length and 3 km width, centred on the Curraghinalt mesothermal gold deposit.

The targets in the TVG and DSG are described in Tables 9.3 and 9.4 and their locations are shown in Figures 9.2 and 9.3.

Table 9.3 Tyrone Volcanic Group Exploration Targets

Target Significant Sample Historic A rea Target defined by Geology Name Samples Type Drilling TVG1 Historic drilling, 750 Auriferous chert- trenching, magnetic high, m x 1.5 m at 4.36 magnetite horizon. 2 shallow Au soil geochemistry, Trench 350 g/t Au Interpreted to be drill holes mineralized outcrop and m exhalative unit. float. TVG2 2.9 EM, Au and Zn - Pb - Cu 1.63 g/t Au Outcrop of altered km x soil geochemistry, and 4.3% Cu 15 shallow Prospecting tuffs in poorly 1.9 mineralized outcrop and + Pb + Zn drill holes exposed area. km float. from outcrop. TVG3 Rhyolite - tonalite 750 5.48 g/t Au contact with m x Au soil geochemistry and from quartz Prospecting abundant angular None 500 mineralized float. float. quartz float with m visible gold. TVG4 Rhyolite breccia Historic drilling and Historic 350 with gold and base trenching, EM, Au soil shallow drill m x metals. Airborne 15 shallow geochemistry, and hole: 3.63 m Drill hole 300 geophysics shows drill holes mineralized outcrop and at 30.12 g/t m new EM float. Au. anomalies. TVG5 EM, IP, magnetic 4.5 Historic Mineralized geophysics, Zn-Cu soil km x Prospecting: ironstone geochemistry, and Prospecting None 700 4.54 g/t Au overlying altered mineralized outcrop and m in ironstone tuffs and basalts. float. TVG6 Auriferous 3.5 EM, Au soil 2.19 g/t Au rhyolite breccias km x geochemistry, and and 2.99% Prospecting and tuffs with None 2.2 mineralized outcrop and Cu + Pb + Zn galena and km float. from outcrop. sphalerite. TVG7 EM and IP, Zn-Cu soil Historic 1.2 Altered volcanic geochemistry, and Prospecting: km x Prospecting tuffs with quartz None mineralized outcrop and 1.87 g/t Au 1 km veins. float. in float

Table 9.4 Dalradian Supergroup Exploration Targets

Target Sample Historic A rea Target defined by Significant Samples Geology Name Type Drilling DSG1 Trench: 9.5 m at 5.64% Metasediment hosted Zn + Pb. Historic 7 km EM, IP, and quartz vein and Prospecting: 141.2 g/t Trench and 12 shallow x mineralized outcrop stratiform gold and Au from float. Recent Prospecting drill holes 600 m and float. base metal prospecting: 33.94 g/t mineralization. Au from float. DSG2 Au soil Graphitic pelite- 2 km geochemistry, hosted breccia zone x 10.52 g/t Au from float. Prospecting None mineralized subcrop up to 50 m wide and 2 50 m and float. km long. DSG3 Quartz float, 200 m 1.8 km 44.96 g/t Au and 32.80 east of drilled x Mineralized float. Prospecting None g/t Au from float. mineralization at 375 m Curraghinalt. DSG4 Au soil 1.5 km Quartz vein, 2 km geochemistry, Channel: 0.88 m @ x Channel east along strike from None mineralized outcrop 39.43 g/t Au. 700 m Curraghinalt. and float. DSG5 1 km Historic drilling, and Historic shallow drill Quartz vein 4.5 km 44 shallow x mineralized outcrop hole: 0.6 m @ 61.43 g/t Drill hole west along strike from drill holes 600 m and float. Au Curraghinalt. DSG6 22.4 g/t Au from quartz 7.5 km Au soil float. 3.36 g/t Au from Wide range of float x geochemistry and Prospecting None silicified metasediment styles in river valley. 600 m mineralized float. float. DSG7 2.5 km Au soil Quartz float in river x geochemistry and 11.68 g/t Au from float Prospecting None valley. 650 m mineralized float. DSG8 Au soil 14.08 g/t Au from quartz 2.3 km Float train of silicified geochemistry, float with pyrite. 8.18 x Prospecting metasediments and None mineralized outcrop g/t Au from silicified 1.7 km quartz veins. and float. metasediment float. DSG9 Au soil 2 km Silicified quartzite geochemistry, 2.96 g/t Au from x Prospecting with disseminated and None mineralized outcrop outcrop. 1.5 km fracture fill pyrite. and float. DSG10 Au soil 1.63 g/t Au from 2.3 km Silicified graphitic geochemistry, graphitic pelite outcrop. 3 shallow x Prospecting pelite and quartz mineralized outcrop 1.88 g/t Au from drill holes 0.9 km veins. and float. outcropping quartz vein. DSG11 Au soil 4 km Quartz breccias with geochemistry, x 2.35 g/t Au from float. Prospecting pyrite in None mineralized outcrop 1 km metasediments. and float. DSG12 Au soil 2 km geochemistry, 2.07 g/t Au from Quartz vein with x Prospecting None mineralized outcrop outcrop. pyrite. 1 km and float.

Following the target identification, a program of regional scout-drilling was started, with two targets having been drilled. Two drill holes were completed at TVG-1 (Broughderg) and one

drill hole was completed at TVG-7 (Tullybrick). There were no significant intercepts in these three drill holes.

In late 2011 Dalradian Resources suspended the scout drilling program in order to gather additional geological information to better define drill targets. Dalradian Resources completed a helicopter airborne 1,034.3 line-km Versatile Time Domain Electromagnetic (VTEM) and magnetic survey along 10 km of the Curraghinalt Trend. The area of the concessions surveyed is shown in Figure 9.4. The survey was flown at a 50-m line spacing. The final interpretation of this work remained outstanding at the time of writing.

Figure 9.4 Airborne Geophysical Survey Coverage

Source: Dalradian Resources, July, 2012

In addition to the airborne geophysical survey, Dalradian Resources initiated a detailed soil geochemistry survey to extend the existing geochemistry grid. The area to be sampled, relative to the existing soil sampling, can be seen in Figure 9.5.

Figure 9.5 Area to be Soil Sampled in Relation to Existing Grid

Source: Dalradian Resources, July, 2012

The soil sampling program remained in progress at the time of writing.

'DOUDGLDQ 5HVRXUFHV¶ SODQ LV WR FRQWLQXH VFRXW GULOOLQJ WKH PRVW SURPLVLQJ WDUJHWV LQ WKH Dalradian Supergroup once the information from the geophysics and geochemistry have been interpreted and drill targets identified.

10.0 DRI L L IN G

The drilling completed by Irish Base Metals/Ennex/Strongbow was described in Section 6, History. The holes drilled by those companies, Tournigan and Dalradian Resources are listed in Appendix 1 of this report. The relevant mineralized drill intersections are listed in Appendix 2.

10.1 TOURNIGAN DRILL PROGRAMS

Irish Drilling Limited (Irish Drilling) was contracted to carry out all drilling on the Curraghinalt area property on behalf of Tournigan and, according to Tully (2005), was one of the principal contractors engaged by Ennex in the 1980s. From 2002, Aurum managed the drilling program carried out by Irish Drilling for Tournigan. The programs were designed primarily by Aurum and Tournigan structural geologist, John Cuthill.

Drilling at Curraghinalt between 2003 and the date of the Micon resource estimate in 2007 is summarized in Table 10.1.

Table 10.1 Summary of Drilling by Year

Drill Holes Number Year Total Metres of Holes 2003* CT-01 to CT-11 11 1,888.8 2004* CT-12 to CT-26 15 2,499.8 2005 CT-27 to CT-28 2 183.0 2006 CT-29 to CT-43 15 2,826.9 2007 CT-44 to CT-53 10 2395.4 Total 53 9,793.9 * - Strongbow-Tournigan joint venture

For the period 2005 to 2007, up to the completion of the 2007 resource estimate, the metreage and details for each hole drilled by Tournigan are shown in Table 10.2. Details for the earlier CT-series holes are found in Appendix 1.

Table 10.2 Summary of Drilling 2005-2007

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 05-CT-27 257135.5 385981.8 245.2 182.9 -50 84.0 05-CT-28 257553.5 385953.0 266.6 176.6 -47 99.0 06-CT-29 257377.5 386301.1 230.4 182.7 -45 240.0 06-CT-30 257436.9 386227.6 241.1 189.3 -47 172.1 06-CT-31 257398.6 386172.2 239.8 185.5 -45 150.0 06-CT-32 257443.5 386147.2 245.0 186.8 -47 120.0 06-CT-33 257404.0 386126.0 242.0 188.8 -49 149.0 06-CT-34 257435.1 386097.8 245.7 190.9 -47 72.0 06-CT-35 257441.3 386304.6 237.0 183.4 -47 263.0 06-CT-36 257507.1 386280.8 243.7 189.3 -47 228.0 06-CT-37 257559.1 386256.7 247.9 190.2 -45 264.0 06-CT-38 257506.6 386209.1 247.3 186.5 -46 180.0 06-CT-39 257503.2 386153.1 250.1 191.8 -47 146.0 06-CT-40 257514.0 386090.0 254.0 193.8 -44 233.8 06-CT-41 257560.4 386196.2 251.9 189.2 -46 207.0 06-CT-42 257600.1 386208.0 252.3 186.3 -45 171.0 06-CT-43 257600.3 386274.6 248.2 199.7 -46 231.0 07-CT-44 257600.0 386136.9 256.1 192.7 -44 131.3 07-CT-45 257558.6 386096.2 256.7 186.7 -45 207.2 07-CT-46 257600.0 386074.9 260.2 195.4 -47 75.0 07-CT-47 257649.6 386216.4 252.9 194.6 -46 230.3 07-CT-48 257650.3 386152.2 256.6 190.1 -45 195.0 07-CT-49 257650.0 386099.0 260.6 197.7 -45 290.0 07-CT-50 257375.1 386536.8 202.7 191.3 -44 398.0 07-CT-51 257342.0 386778.6 156.3 107.3 -46 185.5 07-CT-52 257381.4 386942.1 123.9 194.7 -46 182.0 07-CT-53 256936.6 386810.4 147.5 189.5 -45 501.1 Total 5,405.3

After completion of the 2007 resource estimate the additional holes set out in Table 10.3 were drilled by Tournigan.

Table 10.3 Summary of Drilling 2008

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 08-CT-54 257050.3 386800.2 153.0 195.3 -71 130.0 08-CT-54a 257043.4 386800.5 152.9 210.8 -71 545.5 08-CT-55 256764.5 386753.6 131.5 208.8 -69 633.2 08-CT-56 256907.8 386866.5 135.4 209.7 -71 138.0 08-CT-56a 256907.8 386866.5 135.4 218.4 -72 663.1 08-CT-57 256845.5 386825.3 136.4 213.3 -71 661.0 Total 2,770.8

Tully (2005) describes the program up to January, 2005 as follows:

³7KHGULOOW\SHXVHGIRUWKH'DOUDGLDQ*ROGSURJUDPLVD%R\OHV%URWKHUVPRXQWHGRQD Go-Tract 1000, which allows access to boggy areas. The rig is operated on a 10-hour shift by experienced drillers and helpers. The core size is mostly HQ, although NQ has been utilized when necessary to allow continuation of drilling in difficult conditions when HQ-sized casing is required. Earlier drilling by Ennex produced both HQ and NQ sized core as well. Core recovery is very good and averages in the 95-UDQJH´

Each drill site was fenced off prior to the start of drilling to keep livestock away from the rig. Personal protective equipment was used by drillers and visitors to the site.

Each drill site was located with GPS, staked and then photographed prior to the rig being moved in. The holes were down-hole VXUYH\HGE\XVLQJDQ(QFRUH³5HIOH[LW´VPDUWPXOWL- shot tool instrument. Tully reports that a Tropari survey instrument was used for the first four holes in the Tournigan program. On completion of the drill hole, the collar was surveyed by Celtic Surveys Ltd. using the Irish National Grid, to the nearest centimetre. All holes have been cemented except 04-CT-26, which was left open for potential future use. All sites were cleared on completion and re-photographed.

10.2 DALRADIAN RESOURCES DRILLING

Dalradian Resources began a new drilling campaign following the purchase of Dalradian Gold from Tournigan. The program commenced in March, 2010 with the mobilization of two Boyles 37 track-mounted drills with the capability of drilling to a depth of 500 m using HQ rods. These drills were supplied by Irish Drilling. A second diamond drill contract was signed with Major Drilling and two drills were mobilized to site in December, 2010, with a third Major drill mobilized in August, 2011. The Major drilling fleet consists of one UDR200 track-mounted drill, one AVD6000 track mounted drill and one AVD6100 track-mounted drill. The UDR drill is capable of drilling to 800 m depth using HQ rods and the two AVD rigs are capable of drilling to a depth of 1,500 m using HQ rods. The contract with Irish Drilling concluded in early 2012.

The majority of the drilling with all rigs is HQ diameter. NQ diameter is sometimes used when a reduction from HQ is required due to ground conditions.

Drilling is carried out on a single shift, 7 days per week. Drilling was suspended in March, 2012 and resumed in July, 2012.

Prior to drilling, access agreements are negotiated with the owners of the surface lands and, generally, compensation is paid for use of the land and for disturbances caused by drilling. Before drilling, all drill sites are photographed. They are also photographed upon completion of a drill hole and again after rehabilitation of the drill site.

Drill hole locations are marked out using a GPS and later all drill collars are independently surveyed. Down-hole surveys are carried out on all drill holes using a REFLEX multi-shot

tool, with a reading taken every 6 m. This survey information is stored in the drill hole database.

All drill core is stored at the drill location during the shift, and brought to Gortin at the end of each day. The drill core is stored in a secure shed. For the majority of the drill campaign, the drill core was logged, sampled and photographed at the core facility in Gortin and then moved to storage at one of two facilities located outside Gortin. Dalradian Resources has recently leased a larger office and core facility in Omagh, approximately 15 km from the Gortin facility. Presently core is stored overnight in the secure facility in Gortin and transported the following morning to Omagh, where it is logged, sampled and photographed and stored.

All sampled material and pulps previously stored in Gortin are being transferred to the new facility in Omagh.

)RU0LFRQ¶V2010 Curraghinalt resource estimate, the drill hole database was frozen at hole CT-57. For the 2011 estimate, an additional 45 holes were drilled totalling 20,561.9 m. The locations of these holes are shown in Figure 10.2 and a summary of the then new Dalradian Resources drilling is presented in Table 10.4. This drilling was a combination of holes to the east and west of the previous resource and testing the structures at depth. In addition to the 45 holes completed, five holes, totalling 318.4 m, were drilled, but were not included in the resource estimate as these holes were abandoned at various depths due to drilling difficulties (see Table 10.5).

Drill results for the mineralized intersections in these 45 holes are provided in Appendix 2 of this report.

Vein thicknesses are discussed in Section 14, Mineral Resources of this report.

100

Figure 10.1 North-South Schematic Section at 258555 E, Looking West

101

Source: Dalradian Resources, 2011.

Figure 10.2 Curraghinalt Project, Drill Hole Locations

102

Source: Dalradian Resources, 2011.

Table 10.4 Summary of Dalradian Resources Drilling From Hole CT-58

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 10-CT-58 257649.4 386036.2 265.4 194.9 -47 245.5 10-CT-59 257287.3 386107.6 233.4 197.3 -45 251.0 10-CT-60 257287.5 386108.2 233.3 195.6 -69 244.0 10-CT-61 257294.8 385935.8 248.5 194.5 -70 275.4 10-CT-62 257287.8 385736.1 263.9 200.2 -70 245.5 10-CT-63 257199.5 386246.3 223.4 193.5 -70 324.5 10-CT-64 257370.3 386094.3 241.0 187.9 -71 252.0 10-CT-65 257199.0 386095.2 234.5 198.6 -70 193.0 10-CT-66 257470.6 386119.7 248.7 197.4 -71 250.7 10-CT-67 257206.0 386128.6 232.5 198.6 -70 270.0 10-CT-68 257470.2 385920.5 261.6 193.2 -71 250.0 10-CT-69 257388.9 385925.8 255.3 200.7 -69 250.0 10-CT-70 257110.4 386189.5 231.2 199.8 -71 251.0 10-CT-71 257384.0 386030.6 246.6 198.4 -69 256.0 10-CT-73 257120.4 386427.7 202.2 187.1 -66 266.6 10-CT-74 257033.5 386638.7 171.8 198.4 -42 408.0 10-CT-75 257120.3 386427.7 201.9 189.4 -49 263.0 10-CT-76 257086.6 386828.6 149.1 191.7 -46 762.8 10-CT-77 257289.8 386500.3 202.0 191.4 -45 335.0 11-CT-77a 257289.8 386500.3 202.0 191.4 -45 551.0 11-CT-78 257033.7 386639.7 171.6 186.8 -68 550.0 11-CT-79 257048.8 386204.7 231.3 199.2 -45 250.0 11-CT-80 257086.8 386829.6 148.8 184.7 -69 981.0 11-CT-82 257290.0 386499.2 202.1 190.1 -70 114.0 11-CT-82a 257290.3 386501.2 202.2 190.1 -71 594.0 11-CT-83 256928.8 386273.1 226.5 199.9 -44 338.5 11-CT-84 257048.7 386206.2 234.5 197.5 -67 369.0 11-CT-85 257469.5 386522.4 215.2 192.7 -45 963.0 11-CT-86a 257270.9 386766.4 163.0 201.4 -65 969.0 11-CT-87 256928.9 386274.1 226.5 198.5 -69 393.7 11-CT-88 257847.9 386239.2 244.0 203.7 -69 446.0 11-CT-89 257046.1 386364.3 217.4 165.1 -43 62.0 11-CT-90 257271.5 386767.2 162.9 200.9 -74 981.0 11-CT-91 257046.4 386364.3 214.2 157.0 -45 49.0 11-CT-92 257843.7 386051.5 266.1 195.0 -45 355.0 11-CT-93 257046.4 386364.3 214.2 145.1 -52 73.0 11-CT-94 257469.5 386523.3 215.0 190.0 -60 948.0 11-CT-95 256682.1 386490.5 186.1 190.0 -45 517.0 11-CT-96 257843.5 386050.5 266.2 195.0 -70 469.0 11-CT-97b 257488.9 386813.1 153.5 200.0 -45 957.0 11-CT-98 257681.4 386467.0 227.2 190.0 -45 1,005.0 11-CT-99 258220.4 386474.3 187.8 200.0 -60 1,121.0 11-CT-100 256682.1 386491.6 186.1 190.0 -70 500.4 11-CT-102 256695.1 386305.5 214.0 175.0 -45 344.0 11-CT-103 257671.5 386468.0 226.8 190.0 -70 1,067.3 Total 20,561.9

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Table 10.5 Summary of Abandoned Holes Not Used in Resource Estimation

Drill Hole Hole Depth Number (m) 10-CT-72 8 11-CT-81 35 11-CT-86 44 11-CT-97 188 11-CT-97a 43.4 Total 318.4

After freezing the database for the 2011 mineral resource estimate, diamond drilling continued at Curraghinalt. In total, an additional 15,677 m were drilled in 57 holes as of the date of this report. While they were not used for the resource estimate in this PEA the drill hole information is included in Table 10.6 for completeness. Collar locations of these new drill holes are shown in Figure 10.3. Available assay results for the mineralized (>1.0 g/t Au) intersections in these 57 holes are provided in Appendix 3 of this report.

Figure 10.3 Location of Drill Holes Drilled After the 2011 Resource Estimate Cut-off Date

Source: Dalradian Resources, July, 2012

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Table 10.6 Summary of Dalradian Drill Holes Completed Since the 2011 Resource Update

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 11-CT-101a 258244.3 386086.1 234.5 195 -45 558.0 11-CT-104 257491.1 386814.0 147.5 200 -75 1,001.0 11-CT-105 258222.7 386474.9 182.5 200 -75 1,047.0 11-CT-106 257725.7 386446.7 224.1 190 -52 429.0 11-CT-107 257727.3 386303.0 243.9 190 -50 313.0 11-CT-108 257368.0 386257.2 231.0 185 -50 179.0 11-CT-109 257428.6 386179.6 241.7 185 -50 149.0 11-CT-110 257309.1 386250.8 225.1 185 -50 186.0 11-CT-111 257489.6 386273.6 242.0 185 -50 277.0 11-CT-112 257005.9 386658.8 167.6 182 -50 170.0 11-CT-113 256946.2 386630.6 171.4 182 -50 179.9 11-CT-114 256978.1 386618.4 173.8 182 -50 182.0 11-CT-115 256915.8 386656.6 164.2 182 -50 175.1 11-CT-116 256947.4 386604.4 171.7 182 -50 206.0 11-CT-117 256977.8 386619.2 173.6 182 -70 120.0 11-CT-118 256912.9 386630.3 166.6 182 -50 171.0 11-CT-119 257032.5 386701.2 163.2 182 -50 165.0 11-CT-120 257457.9 386235.8 240.7 185 -48 172.0 11-CT-121 256854.9 386629.4 162.9 182 -50 163.0 11-CT-122 257340.7 386206.5 231.7 185 -50 147.0 11-CT-123 256912.0 386607.2 170.0 182 -50 177.0 11-CT-124 256888.3 386641.4 164.0 182 -50 165.0 11-CT-125 256887.4 386618.6 166.1 180 -50 169.0 11-CT-126 256885.8 386589.2 170.0 180 -50 174.0 11-CT-127 257459.2 386276.1 239.2 185 -49 271.0 11-CT-128 256883.6 386554.0 174.0 180 -50 179.0 11-CT-129 256856.9 386581.3 170.1 180 -50 171.0 11-CT-130 256858.0 386613.9 165.3 180 -50 169.0 11-CT-131 256855.1 386547.7 174.1 180 -50 177.0 11-CT-132 257035.6 386606.5 174.6 182 -50 180.0 11-CT-133 257520.0 386219.0 246.9 185 -50 273.0 11-CT-134 257487.2 386230.3 243.8 180 -50 296.0 11-CT-135 257006.4 386570.9 179.5 180 -50 184.9 11-CT-136 256938.7 386690.4 162.1 182 -50 168.0 11-CT-137 257185.5 386569.5 185.2 182 -50 201.0 11-CT-138 257006.8 386533.1 185.2 182 -50 196.2 11-CT-139 257213.8 386564.1 187.9 182 -50 195.0 11-CT-140 257036.5 386570.2 180.2 182 -50 189.0 11-CT-141 257006.9 386493.3 192.0 180 -50 83.0 11-CT-142 257549.7 386270.5 245.3 180 -50 308.0 11-CT-143 257219.5 386518.7 193.0 180 -50 201.0 11-CT-144 257035.1 386524.3 187.4 181 -50 120.0 11-CT-145 256977.2 386461.7 196.0 180 -50 211.0 11-CT-146 257039.1 386486.1 193.0 180 -50 81.0 11-CT-147 256975.9 386422.7 200.5 182 -50 216.0 11-CT-148 257214.4 386478.7 197.7 182 -50 210.0 11-CT-149 256944.0 386433.0 198.2 182 -50 212.0 12-CT-150 256948.2 386357.6 213.2 180 -50 332.50 12-CT-151 256976.3 386346.5 216.2 180 -45 361.00 12-CT-152 257633.7 385853.3 280.4 180 -45 400.00

105

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 12-CT-153 256563.7 386634.8 163.0 180 -55 501.00 12-CT-154 258602.5 385792.0 262.3 180 -45 420.00 12-CT-155 257282.0 385602.7 276.0 180 -45 452.00 12-CT-156 257255.0 386844.0 150.0 200 -75 720.00 12-CT-157 257574.0 386568.0 195.0 200 -60 639.00 12-CT-158 259278.0 385557.0 243.0 180 -55 384.00 12-CT-159 256192.0 386530.0 196.0 180 -50 300.00 Total 15,676.6

10.3 CORE LOGGING

Prior to sampling, the drill core is washed and logged by the site geologist. This process results in a recovery log, and core annotated with metre-marks. Drill core was logged to record lithology, structure, alteration, rock quality designation (RQD) and mineralization.

Core logging was carried out by geologists assigned to the project by Aurum (until July, 2011) and by Dalradian Resources staff (after July, 2011). The length of core brought from the drill site is confirmed against the drill report. Logging commences with the calculation of core recovery. Geology is then marked with particular reference paid to the identification of mineralized zones. Structural data collected include orientation of mineralized quartz veins against the longitudinal axis of the core with particular care given to mineralized veins that are sub-parallel to the longitudinal axis. Alteration and geotechnical information (RQD) are noted.

In general, core recoveries vary from 90 to 97%, and RQDs are generally in the high 80s or better, except in certain wall rock zones (pelites and faults). In general, the core size drilled is HQ which results in excellent recoveries and, thus, the samples are considered to be representative of the veins. Some drill holes are reduced to NQ to allow for completion of a hole. The overall recovery difference between NQ and HQ core is quite similar.

Logging is carried out on paper with data then being input into a series of digital data logs which are entered into the computer database with other drill hole data logs.

During its first site visit, Micon recommended that all drill core be photographed in order to provide a permanent record and to conform to accepted industry best practice. Tournigan initiated photography of all vein material immediately following the site visit in August, 2007 and, at the time of writing of the 2007 report, this was substantially complete. Presently, all core is routinely photographed and the photographic record, along with copies of the paper drill logs, geotechnical logs and assay certificates are stored in a separate file for each drill hole. Figure 10.4 shows an example photograph of mineralized drill core.

106

Figure 10.4 Quartz-Sulphide Vein in Drill Core

The quality assurance/quality control (QA/QC) procedures for the drilling and logging processes are shown in Table 10.7. This table has been based on information provided by Dalradian Resources. 0LFRQ¶VREVHUYDWLRQRIFRUHORJJLQJGXULQJWKHYLVLWVKRZHGLWWR be carried out to industry standards.

107

Table 10.7 Drill Set-up and Core Logging Controls

Procedure Comments Drill Rig Setup Drillers to set up rig. Check location of small ditch to be dug from the pierce point for the drill return water. Returns are directed to a large excavated sump or a series or two sumps in series, where the returns settle out. Clear water is then pumped from the sump back to the drill rig for reuse. Drill rig to be checked daily (early afternoon); vein material, in core box, to be taken directly to core logging shed. Drillers to deliver drill core to the logging shed each evening unless collected by the geologist. Drill Core Approximate depths to be reported as core comes in for ongoing section correlation. Logging Sequence of core boxes to be checked for correct order; boxes to be checked for correct labelling. Core markers to be checked every 3 m. Core loss to be checked if markers are not evenly spaced. Orientate drill core so that the general schistosity is consistent. Complete RQD on every 3-m run down the core using detailed recording sheet. RQD, joint frequency and orientation, roughness, joint infill, fractures, faults etc. to be recorded. Log lithology using pelite/semi-pelite/psammite system. Record host rock mineralogy/alteration. Record nature of hanging wall and footwall contacts. Log veins; measure hanging wall and footwall contact angles if natural and not faulted. Record mineralogy in detail; identify zoning/banding. Separate vein into areas of similar mineralogy and percentage sulphides. Log structure and alteration. Enter data into digital data book on an ongoing basis. Surveying Drill Down the hole surveys are carried out with reflex multi-shot tool. Two digital files are Hole produced and converted into Microsoft Excel file format using proprietary software. These files are then reviewed by the geologists for errors and then loaded in to the drill hole database.

108

11.0 SA MPL E PR EPA R A T I O N, A N A L YSES A ND SE C URI T Y

11.1 SAMPLING

The Dalradian Resources sampling approach and methodology is as follows.

All sulphidic quartz vein material intersected in the drill core is sampled.

Core sample intervals are selected by the geologist logging the hole. The maximum sample interval is currently 0.25 m in mineralized veins with no minimum width (intervals are as short as 0.02 m). The 0.25-m sample length was set so that the entire sample is pulverized at the laboratory, eliminating the need to split the crushed sample. Samples in weakly mineralized or unmineralized wall rock are usually significantly larger (i.e., up to 1 m).

Sample intervals tend to be split by the geology and or mineralization noted (i.e., sulphide- rich zones in veins are separated from quartz-rich zones), and wall rock samples are only taken if mineralized (i.e., stockwork veins are present).

During the logging process the geologist marks the intervals to be sampled on both the core and the core box. The core is then oriented along the line of symmetry and sawn with a diamond-tipped saw. This process allows for an accurate depth measurement to be assigned to each sample. Once an interval is cut, both halves of the core are placed back in the box, and the next interval is sawn.

A sample ticket book is used to record sample intervals and a brief mineralogical description. Each sample is assigned an identifying number as printed on each ticket. A plastic sample bag is then clearly numbered with the matching ticket number placed inside with the split core. A duplicate ticket is stapled to the core box at the start of the sample interval above the remaining half core in the core box.

Once bagged, all the samples split during each work shift are placed on the floor in numerical order. When all the individual samples have been counted, and the tag numbers verified, the samples are placed into large plastic bags. Each larger bag of samples is sealed with a wire tag and an address label. These bags are stored in the locked core shed until shipment to the laboratory by a commercial courier service

The QA/QC procedures for the sampling processes are shown in Table 11.1, which has been based on information provided by Dalradian Resources.

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Table 11.1 Drill Set-up, Sampling Procedures and Controls

Procedure Comments Sampling Orient the core. Mark sample boundaries using mineralogy and percentage sulphide as guidelines. Record intervals on sample sheet. Sample main veins at 0.25-m intervals. Split the actual vein width only if vein intercept is less than 10 cm. Maximum sample width 0.25 m. Insert one analytical standard packet as a sample amongst each main vein; approximately 20% of all samples per hole. Insert one blank sample of drill core as a sample amongst each main vein; approximately 20% of all samples per hole. Photograph sampled vein intervals once sampled drill core has been returned to shed.

11.2 SAMPLE ANALYSES

Currently, all exploration samples at Curraghinalt are analysed at OMAC Laboratories (OMAC), Athenry Road, Loughrea, County Galway, Republic of Ireland. The OMAC laboratory was also used by Tournigan. Before Tournigan, sample analysis was carried out by Ennex at OMAC as well, except for the underground sample analyses from the development drifts which were completed at the Ennex Mine Laboratory, essentially a satellite OMAC laboratory.

OMAC joined the Alex Stewart Assayers group in 2000 and operates as its principal exploration laboratory. In July, 2011, The Alex Stewart Group was acquired by the ALS Group, the testing services business of Campbell Brothers Limited. OMAC states that it is accredited to ISO 17025 by the Irish National Accreditation Board (INAB). ISO 17025 relates to general requirements for the competence of testing and calibration laboratories. INAB is a member of the International Accreditation Cooperation (ILAC) and is a signatory to the ILAC Mutual Recognition Arrangement whose signatories include Canada, the United States, Australia, South Africa, Japan and countries of the European Union, among others. OMAC states that its accreditation schedule can be accessed through the INAB website under Registration No. 173T (www.inab.ie).

OMAC participates in twice-yearly round robin programs run by Geostats of Perth, Australia, and under the Proficiency Testing Program for Materials Analysis Laboratories (PTP-MAL) run by CANMET in Canada. (See www.omaclabs.com/QA-QC.htm).

The sample preparation and analytical methods used by OMAC are shown in Table 11.2

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Table 11.2 OMAC Laboratory Assay Procedures

Procedure Procedure Code P5 Dry. Crush to <2 mm using Rhino jaw crusher. 5LIIOHNJDQGSXOYHUL]HWRȝXVLQJ%LFRSXOYHUL]HU. Au4 50 g sample fused with lead oxide/carbonate/borax/silica/flux at 1,100oC using silver a carrier. Fusions producing lead buttons <30 g are rejected. After de-slagging, buttons are cupelled at 950º C. Prills are parted in dilute nitric acid and finally dissolved in aqua regia. Au read by flame AA to 0.01ppm using a Varian Spectr AA 55 AA instrument. Samples run in batches of 50 consisting of a) 48 samples, b) blank and in-house standard or international reference material. 10% of the samples repeated subsequently by procedure Au5. MA/ES Multi Acid Digestion with ICP-AES. Acid attack in open PTFE beakers. Multi element suite measured by injecting sample into Inductively Coupled Plasma (ICP). Resulting emission spectra measured with Atomic Emission Spectrometry (AES). Element suite comprises: Ag, Al, As, Ba, Be, Bi, Ca, Cd, Ce, Co, Cr, Cu, Fe, Ga, Ge, Hg, K, La, Li, Mg, Mn, Mo, Na, Nb, Ni, P, Pb, Rb, S, Sb, Sc, Se, Sn, Sr, Ta, Te, Th, Ti, Tl, U, V, W, Y, Zn, Zr.

Micon confirmed with the site geologist and a director of Aurum, and representatives of Tournigan during the 2007 site visit, that the procedures described by Tully for sample preparation and assaying remain in place.

11.3 SAMPLE SECURITY

All drill core is brought from the drills at the end of shift, and is stored in a rented lockable storage facility on the main street of Gortin, across the road from the field office. Until November, 2011 core was sawn at a nearby farm and unmineralized core boxes were stored in a concrete-floored barn at that location. Mineralized sample boxes were returned to the industrial building where they were kept under lock and key. In November, 2011, Dalradian Resources leased a new office and core facility in nearby Omagh. The drill core is brought from the drills at the end of the shift and stored overnight in the secure facility on the main street in Gortin as previously was done. The following day, the core is brought to the Omagh facility where it is logged, sawn and stored. The Omagh facility is located in a secure fenced area, which is locked in the evenings and the weekends. The core storage and logging facility is kept locked when unoccupied. Unshipped samples are also stored at this location.

11.4 QUALITY CONTROL AND QUALITY ASSURANCE

11.4.1 Early QA/QC Programs

The QA/QC procedures have varied slightly over the history of the exploration of the deposit, and can be broken into three main phases, the Ennex work, the Tournigan work and the

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Dalradian Resources work. The database used in this resource estimate contains assay information from all drilling carried out on the property

Tully (2005) describes sample Tournigan QA/QC procedures as follows:

³4XDOLW\FRQWUROLV PDLQWDLQHG E\LQVHUWLQJDEODQNVDPSOH RIOLPHVWRQH drill core from the midlands of Ireland (Lisheen area), known to contain no gold, after each mineralized interval, DQGJLYHQWKHQH[WVHTXHQWLDOLGHQWLI\LQJQXPEHU7KLVVDPSOHLVUHFRUGHGDVD³EODQN´LQWKH ticket book. At least 15% of the samples submLWWHGDUH³EODQNV´.

³In addition, Tournigan uses six different standards of known gold grade that it submits as control samples. The standard name is recorded in the sample ticket book and when the assay results from the lab are received, they are comparHGZLWKWKHNQRZQJUDGHRIWKHVWDQGDUG« 7KHUHVXOWVIURPWKH³6WDQGDUG´FKHFNDQDO\VLVKDYHDJUHHGYHU\ZHOO

³Sample pulps are returned to the mine office in Gortin once analysis is complete. Fifteen to twenty percent of the returned pulps are reanalyzed. These samples are sent back to OMAC [Laboratories] for re-analysis under a new sample number. A recent change to the procedure is to send half of the duplicates to OMAC and the remaining half of the 20% to an outside laboratory as an additional check on OMAC.´

Drilling carried out in the Curraghinalt area totaled about 27,299 m, up to August 15, 2007. In addition, an approximate total of 740 m of underground workings was established from an adit developed by Ennex and which included drifts, cross-cuts and raises. This drifting helped confirm continuity of mineralized structures. After August 15, 2007 an additional 2,502.8 m of core were drilled in the final four deep holes.

Assaying for all drilling completed by Tournigan was carried out by OMAC at its facility at /RXJKUHD&RXQW\*DOZD\,UHODQG3ULRUWR7RXUQLJDQ¶VSURJUDPVDVVD\LQJof underground samples was carried out by Ennex at its internal laboratory which is referred to in this report as the Ennex Laboratory. Ennex drill core was assayed at OMAC.

Out of the total 27,299 m drilled, 8,923 m of drilling was carried out during 2006-2007. 7RXUQLJDQ¶V4$4&SURJUDPFRQVLVWHGRIDVVD\LQJRIEODQNVUHIHUHQFHPDWHULDOVGXSOLFDWH samples, and the re-assay of samples in a reference laboratory. Check assaying undertaken by Tournigan included samples from the Ennex drill programs assayed at the Ennex Laboratory. 7KHGHWDLOV RIVDPSOHV DVVD\HGDVSDUWRI7RXUQLJDQ¶V4$4& SURJUDP DUH JLYHQLQ 7DEOH 11.3.

Table 11.3 'HWDLOVRI7RXUQLJDQ¶V4$4&6DPSOHV

Number of Type of Q A/Q C Samples Samples Blanks 315 Standard 134 Duplicate 174 Re-assay at External Laboratory 43

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11.4.2 Dalradian Resources QA/QC Program

The QA/QC program was under the supervision of Aurum from the restart of drilling in 2010 until July, 2011, at which time it was under the supervision of Dalradian Resources staff.

The Dalradian Resources sampling and QA/QC procedures can be described as follows:

1) Core sample intervals are selected by the geologist as described above.

2) Sample intervals tend to be split by geology as described in Section 11.1 above.

3) Samples are cut by diamond saw down the ellipse of the vein. The same half is always taken for the sample and the other half is left in the core tray for reference. Cross cuts at sample extents are along the contact of the vein and wall rock, where this is possible on both vein contacts (with the interval measured to the centre of the core axis) or, if this is impossible, perpendicular across the core axis.

4) To check for contamination during sample preparation at the laboratory, blanks are inserted where the geologist estimates the sample will be high grade. These are inserted so that approximately 10% of the submitted samples are blanks. The blank material is chunky (fist-sized) limestone from a local source that has been used extensively enough so that there is confidence that the material contains negligible amounts of gold.

5) A certified Standard Reference Material (SRM or referred to in this report as a standard) is inserted into the sample stream so that they make up approximately 10% of the samples submitted. These are 50-g sachets of pulverized, homogenized rock with gold and sulphide added. The accepted gold grade of each standard is determined by round-robin assaying at multiple laboratories. These are supplied by Rocklabs (New Zealand), with around five or six different standards in circulation at any one time and a current gold grade range from 1.34 to 30.04 g/t Au. The selection has been chosen to reflect the typical range of vein assays as it is the accuracy within this grade range that is of interest. The standard that is inserted is randomly selected.

6) Before being dispatched the samples are sealed in bags and an additional signed and dated, tamper proof, security tag is placed around the wire tie. All samples and core that will be sampled is stored in a locked core shed.

7) Specific gravity (SG) measurements are attempted on each vein sample, using the water submersion method.

Out of the total 51,575 m drilled at Curraghinalt, 20,561.9 m of drilling was carried out during 2010-2011. Dalradian Resources¶ 4$4& SURJUDP FRQVLVWV RI DVVD\LQJ RI EODQNV and standard reference materials and a duplicate assay program at OMAC. No sample re-assaying

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has been conducted at a second reference laboratory. The details of samples assayed as part of Dalradian Resources¶4$4&SURJUDPDUHJLYHQLQ7DEOH11.4.

Table 11.4 Details of Dalradian Resources¶ QA/QC Samples

Type of Q A/Q C Samples Number of Samples Blanks 347 Standard 242

As part of its review of the quality control data, Micon conducted a number of checks to assess precision, accuracy and bias. This is described in the following sections of the report.

11.4.3 Blank Sample Results

Pre-Dalradian Resources

Blank samples form part of all analytical quality control procedures and are used to assess the accuracy of the assay results and to identify any possible contamination and mixing during analysis. A total of 315 blanks were inserted along with the other samples. As blanks are close to the detection limit of the laboratory, it is expected that the error in analysis at that level can be +50%. There are two blanks which had considerable error and these are designated as outliers. All the results were plotted in Figure 11.1. The accuracy of the blanks is considered to be within an acceptable limit of error.

Figure 11.1 Results for Blanks Analyzed by Ennex

) 0.1 u A

t / g (

e d a r G 0.01

0.001 0 50 100 150 200 250 300 350 No. of Samples

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Dalradian Resources

A total of 347 blanks were inserted along with other samples by Dalradian Resources. There were three blanks which had considerable error and these are designated as outliers (Table 11.5). All the results were plotted in Figure 11.2. The accuracy of the blanks is considered to be within an acceptable limit of error.

Table 11.5 Dalradian Resources Blank Samples Designated as Outliers

Dalradian Au Sample Resources (g/t) O rder # B0762 0.16 DAL236 A8487 0.47 DAL203 A7915 0.11 DAL179

Figure 11.2 Results for Dalradian Resources Blanks Analyzed by OMAC

11.4.4 Reference Standard Sample Results

Pre-Dalradian Resources Results

A total of six different types of reference material were used to check the accuracy of assay results at the Tyrone project. All standards were acquired from CDN Resource Laboratories Limited (CDN Resource). The average value for each standard, along with its tolerance, is given in Table 11.6. A total of 134 standard samples were submitted along with the drill core samples. The average value received for each of the standards is given in Table 11.7. It may

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be seen that there is no significant bias while analyzing the standards. Figure 11.3 shows an error plot compiled considering all the standards with the mean recalculated to zero and deviation plotted against the mean. These results indicate that the assay results likely do not have any significant bias or accuracy issues. All the results for the individual analyzed standard samples along with their certified value and acceptable range of standard deviation are plotted in Figures 11.4 to 11.10.

There are some outliers from assays of standard samples that are beyond the acceptable limit of 2 standard deviations but these do not indicate bias or misallocation. Micon had recommended that Tournigan review the outliers and, if necessary, re-assay the complete sample batch.

Overall, Micon considers that the Tournigan results are within the acceptable range of error.

Table 11.6 Certified Value of CDN Resource Standard Samples

C DN- C DN- C DN- C DN- C DN- C DN-

GS-11 GS-5A GS-14 GS-12 GS-15 GS-20 Mean 3.4 5.1 7.47 9.98 15.31 20.6 +/- 2 Standard Deviation 0.27 0.27 0.31 0.37 0.58 0.67 Standard Deviation % of Mean 8% 5% 4% 4% 4% 3% Number of Samples 84 84 84 84 84 84

Table 11.7 Analyzed Value of CDN Resource Standard Samples

C DN- C DN- C DN-GS- C DN- C DN- C DN-

GS-11 GS-5A 14 GS-12 GS-15 GS-20 Mean 3.45 5.05 7.56 10.16 15.30 20.61 +/- 2 Standard Deviation 0.15 0.29 0.55 0.88 1.14 1.06 Standard Deviation % of Mean 4% 6% 7% 9% 7% 5% Number of Samples 19 20 29 20 27 19 Bias 1% -1% 1% 2% 0% 0%

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Figure 11.3 Error Plot for the Standard Samples Used by OMAC

80% e 70% u l a

V 60% d r a 50% d n a

t 40% S e

h 30% t o t

20% e c n

e 10% r e f f

i 0% D -10% %

-20% 0 20 40 60 80 100 120 140 160 No. of Samples

Figure 11.4 Repeatability by OMAC on Reference Material (3.4 g/t Au)

3.80

3.70 +2 Std. Deviation 3.60

3.50

3.40 Grade (g/t Au) Grade 3.30

3.20 -2 Std. Deviation 3.10 0 2 4 6 8 10 12 14 16 18 20 No. of Samples

117

Figure 11.5 Repeatability by OMAC on Reference Material (5.1 g/t Au)

5.50

5.40 +2 Std. Deviation 5.30 ) u 5.20 A t / g (

5.10 e d a r 5.00 G

4.90 -2 Std. Deviation 4.80

4.70 0 5 10 15 20 25 No. of Samples

Figure 11.6 Repeatability by OMAC on Reference Material (7.47 g/t Au)

8.40

8.20

8.00 ) u 7.80 A

+2 Std. Deviation t / g (

7.60 e d a r 7.40 G

7.20 -2 Std. Deviation

7.00

6.80 0 5 10 15 20 25 30 35 No. of Samples

118

Figure 11.7 Repeatability by OMAC on Reference Material (9.98 g/t Au)

11.40

11.20

11.00

10.80 ) u 10.60 A t / 10.40 g ( +2 Std. Deviation e

d 10.20 a r

G 10.00

9.80 -2 Std. Deviation 9.60

9.40

9.20 0 5 10 15 20 25 No. of Samples

Figure 11.8 Repeatability by OMAC on Reference Material (15.31 g/t Au)

16.5

16 +2 Std. Deviation

) 15.5 u A

t / g (

15 e

d -2 Std. Deviation a r G 14.5

13.5 0 5 10 15 20 25 30 No. of Samples

119

Figure 11.9 Repeatability by OMAC on Reference Material (20.6 g/t Au)

22.00

21.50

+2 Std. Deviation

) 21.00 u A

t / g (

20.50 e d a r G 20.00 -2 Std. Deviation

19.50

19.00 0 2 4 6 8 10 12 14 16 18 20 No. of Samples

Dalradian Resources Standard Results

A total of 10 different types of reference materials were used to check the accuracy of assays. All standards were acquired from Rocklabs Ltd. (New Zealand). The average value for each standard along with its tolerance is given in Table 11.8. A total of 288 standard samples were submitted along with the drill core samples. Three samples had clear evidence of being attributed in the database with the incorrect reference standard code. These were discarded from the graphical analysis (see Table 11.9 for details). The average received value for each of the standards is given in Table 11.8.

It may be seen that there are no significant bias or accuracy issues in analyzing the standards. This indicates that the assay results would not have any significant bias. An error plot (Figure 11.10) was also compiled considering all the standards with the mean recalculated to zero and deviation plotted against the mean. All the results for analyzed standard samples, along with their certified value and acceptable range of standard deviation, are plotted in Figures 11.11 to 11.20.

There is one outlier from the assay results of the standard samples that is beyond the acceptable limit of 2 standard deviations (Sample A8366, Standard SH41) but this does not suggest any systematic bias or misallocation. Dalradian Resources regularly reviews the outliers and, if necessary, will re-assay the complete sample batch, however, this has not yet been deemed to be necessary.

Overall, Micon considers that the Dalradian Resources results are within the acceptable range of error.

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Table 11.8 Certified & Analyzed Value of Rocklabs Standard Samples

Certificate O M A C Analyzed Standard No. Used Mean 2SD SD 2SD 3SD Median Average Bias % SH35 25 1.323 0.088 0.03 0.07 0.1 1.27 1.28 -3.25 SH41 25 1.344 0.082 0.13 0.25 0.38 1.28 1.27 -5.51 SJ39 24 2.641 0.166 0.12 0.24 0.36 2.56 2.54 -3.82 SJ53 26 2.637 0.096 0.07 0.15 0.22 2.54 2.55 -3.30 SL46 25 5.867 0.34 0.2 0.4 0.61 5.85 5.85 -0.29 SL51 21 5.909 0.272 0.21 0.43 0.64 5.86 5.89 -0.32 SN38 25 8.573 0.316 0.18 0.37 0.55 8.52 8.51 -0.73 SN50 23 8.685 0.36 0.23 0.46 0.69 8.44 8.44 -2.82 SP37 48 18.14 0.76 0.54 1.07 1.61 17.92 17.92 -1.21 SQ36 47 30.04 1.2 0.9 1.81 2.71 29.3 29.38 -2.20

Table 11.9 Samples Discarded From QA/QC Analysis Due to Likely Data Entry/Recording Error

G rade Analyzed SR M G rade Hole ID Sample ID Standard Code Batch (g/t Au) (g/t Au) 10-CT-71 A4993 SQ36 DAL172 7.29 30.04 11-CT-100 B1728 SP37 DAL283 5.68 18.14 11-CT-103 B1912 SL51 DAL284 17.44 5.909

Figure 11.10 Error Plot for Standard Samples Used by OMAC

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Figure 11.11 Curraghinalt Drilling Standard SH35 X-Chart (OMAC) (SR M = 1.323 ppm)

Figure 11.12 Curraghinalt Drilling Standard SH41 X-Chart (OMAC) (SR M = 1.344 ppm)

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Figure 11.13 Curraghinalt Drilling Standard SJ39 X-Chart (OMAC) (SR M = 2.641 ppm)

Figure 11.14 Curraghinalt Drilling Standard SJ53 X-Chart (OMAC) (SR M = 2.637 ppm)

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Figure 11.15 Curraghinalt Drilling Standard SL46 X-Chart (OMAC) (SR M = 5.867 ppm)

Figure 11.16 Curraghinalt Drilling Standard SL51 X-Chart (OMAC) (SR M = 5.909 ppm)

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Figure 11.17 Curraghinalt Drilling Standard SN38 X-Chart (OMAC) (SR M = 8.573 ppm)

Figure 11.18 Curraghinalt Drilling Standard SN50 X-Chart (OMAC) (SR M = 8.685 ppm)

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Figure 11.19 Curraghinalt Drilling Standard SP37 X-Chart (OMAC) (SR M = 18.14 ppm)

Figure 11.20 Curraghinalt Drilling Standard SQ36 X-Chart (OMAC) (SR M = 30.04 ppm)

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From a review of the different standard samples, it is evident that OMAC Laboratory has consistently under reported the grades of the reference material except for four standards. In the case of these four, three look reasonable and one (2.641 g/t Au) over reports the grade. As a result, the bias is consistently negative. However, when reviewed in absolute terms, the bias is small and there is no reason to raise an alarm with the laboratory. This is something to keep a close eye on in the future. Micon concludes that the QA/QC procedures conform to industry standards and the data generated as a result are suitable for use in the present mineral resource estimate. No significant bias has been identified as a result of the above analysis and which warrants further study.

11.4.5 Duplicate Sample Results

As part of the internal quality control by OMAC, the laboratory conducts duplicate pulp assays on selected samples (there being no coarse reject to go back to as a result of total pulverization of the sample). A total of 765 re-assays were completed (including multiple re- assays of one sample). Dalradian Resources indicates that it has recently started submitting samples from the returned pulps for their own its check.

A scatter plot (Figure 11.21) was generated to understand the relationship between the original assays and the repeat assays. The plot clearly shows good repeatability of original samples during re-assay.

Figure 11.21 Scatter Plot between Original and Re-submitted Samples

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An error plot was generated to understand the degree of error of individual samples between the original samples with the repeat samples. This was done by plotting the mean of the two results with the absolute difference to the mean. The plot is shown in Figure 11.22. The plot shows that 87% of samples with a mean value above 5 g/t Au fall within a 10% absolute error.

Figure 11.22 Error Plot for Gold Assays

100 r or or o r r rr < 5% Absolute E rror (56%) r r E E E or % % % rr 5-10% Absolute E rror (18%) 0 5 0 E 10 2 1 10- 25% Absolute E rror (15%) 5% 10

) >25% Absolute E rror (11%) u A

t / g (

e 1 c n e r e f f i D 0.1 e t u l o s b A 0.01

0.001 0.001 0.01 0.1 1 10 100 1000

Mean Gold Grade (g/t Au)

Micon concludes that the QA/QC procedures used conform to industry standards and the data generated as a result are suitable for use in the present mineral resource estimate. No significant bias has been identified as a result of the above analysis and which warrants further study.

11.4.6 Dalradian Resources QA/QC Since October, 2011

Since the October, 2011 database freeze date for the mineral resource estimate employed in this PEA Dalradian Resources has continued the QA/QC program as per the methodology described above. Summary statistics are provided below for the numbers of standards, blanks and duplicates that have been submitted for analysis since the database was frozen for the November, 2011 resource estimate. The results of this QA/QC program have no influence on the mineral resource estimate used herein.

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Dates: October 10, 2011 - Aug 16, 2012 Standards = 599 Blanks = 616 Duplicates = 202

Additional reference standard materials have been supplied by Rocklabs during this time to make up for depleted stock levels. These new standards have similar grade distributions to those described in Table 11.8.

11.5 SPECIFIC GRAVITY DETERMINATIONS

Tully (2005) provided a tabulation of specific gravity (SG) determinations carried out by Tournigan in 2004. Two samples from drill core from each of the veins were sent to OMAC for specific gravity determinations by the water displacement method. Table 11.10 summarizes the specific gravity determinations for the veins as shown.

Table 11.10 Specific Gravity Determinations

Sample Number SG Sheep Dip Vein SD41 3.01 SD43 2.54 Average 2.78 Attagh Burn Vein (ABB) ABB34 2.49 ABB45 3.65 Average 3.07 T17C Vein 17C-50 3.65 17C-11 2.86 Average 3.05 T17HW Vein 17HW-4 2.62 17HW-54 3.51 Average 3.07 T11F Vein T11F-14 2.90 T11F-28 2.97 Average 2.93 No. 1 Vein No1-21 3.16 No1-52 2.89 Average 3.03

A total of 583 specific gravity samples have been generated since the Tournigan work. A scatter plot was also generated in order to understand the relationship between gold grade and

129

specific gravity. The plot is shown in Figure 11.23. The plot indicates that generally there is no significant relationship between the gold grade and specific gravity. However, there is a positive correlation when the grade is above 10 g/t. This result prompted Micon to use the average value of the specific gravity for each vein.

Samples measured for specific gravity were selected for each vein/sub-vein/veinlet. It was found that not all 35 veins have been analysed for specific gravity. In the absence of some data, samples were regrouped into major veins and average values were calculated for each major vein as shown in Table 11.11. The specific gravity assigned for minor veins and veinlets is the same as for the major veins. For veins D and G, specific gravity values used were the same as those from the last resource estimate.

Figure 11.23 Plot of Gold Grade against Specific Gravity

4 ) 3 m / t (

y 3 t i s n e D

1 0.001 0.01 0.1 1 10 100 Grade (g/t Au)

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Table 11.11 Mean Specific Gravity and Standard Deviation for Major Veins

Standar Numb Co- Numb Minimu Maxim Averag d Densit Vein er of efficient er of Vein m um e Deviatio y Used ID Sampl of Sampl G roup (t/m3) (t/m3) (t/m3) n (t/m3) es Variation es (t/m3) 10 22 2.43 3.40 2.70 0.22 0.08 22 10 2.70 20 18 2.48 3.18 2.81 0.19 0.07 18 20 2.81 30 22 2.43 3.68 2.88 0.35 0.12 22 30 2.88 40 78 2.47 4.12 2.85 0.32 0.11 403 8 2.41 3.02 2.73 0.21 0.08 98 40 2.84 DL426 12 2.69 3.36 2.84 0.21 0.08 B 50 7 2.61 3.41 2.98 0.31 0.10 50E 5 2.68 3.49 2.96 0.28 0.09 14 50 2.96 55 2 2.68 3.14 2.91 0.23 0.08 60 104 2.43 3.99 2.86 0.25 0.09 604 10 2.66 2.88 2.79 0.08 0.03 605 3 2.68 2.89 2.80 0.09 0.03 121 60 2.85 606 3 2.76 3.00 2.85 0.11 0.04 607 1 3.09 3.09 3.09 - 70N 27 2.44 3.63 2.87 0.26 0.09 70S 97 2.45 7.39 2.83 0.49 0.17 701 11 2.67 3.06 2.78 0.13 0.05 202 70 2.83 703 2 2.52 3.20 2.86 0.34 0.12 75 65 2.51 3.49 2.83 0.19 0.07 80 32 2.67 3.27 2.82 0.15 0.05 DL829 37 80 2.81 5 2.72 2.84 2.75 0.05 0.02 0 90 49 2.53 3.59 2.81 0.20 0.07 49 90 2.81 D 2.57

G 2.57

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12.0 D A T A V E RI FI C A T I O N

12.1 DALRADIAN RESOURCES DATA VERIFICATION

The QA/QC and data verification programs employed by Dalradian Resources and their results have been discussed in Section 11.4 above.

12.2 MICON DATA VERIFICATION

Micon was provided with the drill hole data from Curraghinalt and is satisfied that they are suitable for the present mineral resource estimate.

During the second site visit in 2009, Micon collected five samples to confirm the presence of gold and copper mineralization at the Curraghinalt deposit. A sixth sample was collected from silicified and mineralized outcrop at the Cashel Rock showing on licence DG2. The results of this sampling are set out in Table 12.1.

Table 12.1 Micon Check Samples

Sample Au Ag Cu Location Type No. (g/t) (g/t) (ppm) 75117 T17 Vein, Curraghinalt Underground grab 15.1 6.7 1,790 sample 75118 T17 Vein, Curraghinalt Underground grab 4.2 4.0 2,150 sample 75119 #1 Vein, Curraghinalt Underground grab 21.9 5.1 54 sample 75120 Hole CT55, Curraghinalt Duplicate ¼ core sample 42.6 9.2 97 75121 Hole CT54a, Curraghinalt Duplicate ¼ core sample 45.2 9.0 25 75122 Cashel Rock Surface grab sample 8.71 10.9 555

The original assay results for the duplicate quarter core samples 75120 and 75121 were 44.96 g/t Au and 23.68 g/t Au. The samples collected by Micon have confirmed the presence of gold mineralization at the expected grades on both the DG1 and DG2 licences.

The checks carried out by Micon confirm the logging procedures utilized at the Curraghinalt site. All drill core is stored in wooden boxes with proper numbering to indicate the drill hole number and metreage. Random checks were carried out by Micon on the stored core and no discrepancies were identified.

The drill hole database is maintained in Microsoft Access and work is carried out at ǯ ǡ Ǥ ensure that there are no errors in data entry. Assay results are provided in Excel files

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from OMAC, thereby allowing direct electronic data transfer. This eliminates any potential error arising from manual data entry.

Micon utilized the data verification protocols available in Datamine software to check the database for errors such as transposed or crossed from and to entries for sample intervals, or incorrect entries in data fields.

Hanging wall and footwall contacts of mineralized veins were developed by contractors of Dalradian in constant consultation with Micon. Several intermediate amendments and modifications were suggested by Micon which were adopted to create the final set of wireframes for each vein. Wireframes were also verified for intersections between hanging wall and footwall and between veins. In addition, Micon carried out spot checks to confirm that the changes have been adopted. Micon did not review each intersection for each vein.

The data submitted to Micon appear reliable in light of the checks it has carried out. Micon has not independently verified the statements and data contained in historical reports or in assay data provided to Micon for the purpose of this mineral resource estimate, other than those steps described herein.

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13.0 M IN E R A L PR O C ESSIN G A ND M E T A L L UR G I C A L T EST IN G

Reports dating from 1985 to 1999 have previously been prepared describing metallurgical testwork completed on samples from the Sperrin Mountain and Curraghinalt project. More recently, testwork has been completed by SGS Mineral Services (Lakefield) in 2012, and some work is currently being carried out at G&T Metallurgical Services Ltd., Kamloops, BC, Canada (G&T).

A Bond ball mill index test was carried out in 1985 and repeated in 2012 and an Abrasion index test was completed on a single composite. Gravity concentration testing has been completed which indicates the samples are amenable to this process. All samples tested responded well to cyanidation but cyanide consumption was high. In 1986 Lakefield examined pre-aeration and lead nitrate addition which resulted in improved recovery rates and lower cyanide consumption. Alternatively, the Gravity Recoverable concentrate could be suitable for sale to a smelter.

Flotation testwork has been undertaken that suggested that a copper concentrate could be produced for sale to a smelter. A Bulk flotation concentrate was also produced during testwork which was had high gold recovery and was amenable to cyanidation, or could possibly be suitable for sale to a smelter.

Testwork Reports reviewed are referenced in section 27.0.

It was observed that there was a wide variation in gold, silver and copper head assays in the samples used for earlier testwork. The exact origin of samples from the earlier studies was not clearly identified. This earlier testwork was primarily conducted on samples from the T-17, No 1 and Sheep Dip veins.

13.1 SAMPLE CHARACTERISATION - MINERALOGY

Mineralogical studies were conducted by Lakefield on a sample in 2011 and are being undertaken by G&T as part of an ongoing testwork program. This work included QEMSCAN and chemical assaying.

Mineralogy work conducted in 2004 indicated that gold mineralization at Curraghinalt occurs in quartz-pyrite veins and is associated with variable abundances of carbonate, chalcopyrite and tennantite-tetrahedrite. In general, carbonate, chalcopyrite and tennantite-tetrahedrite are paragenetically later than quartz and pyrite, and fill fractures in the latter. Gold occurs mainly as the native metal and more rarely as electrum (>20 wt% Ag), and is found primarily along fractures in pyrite, as inclusions in pyrite, and at pyrite grain contacts with carbonate and quartz. Most native gold grains are associated paragenetically with carbonate, chalcopyrite, tennantite-tetrahedrite and telluride minerals. The report examined mineralized samples from all the veins and concluded that generally the mineralogy was similar in all the veins found at Curraghinalt.

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Initial results from G&T (final report not received), indicate that 96.0% of the copper-bearing minerals occur in the sample as chalcopyrite, with chalcocite/covellite (1.4%) and tetrahedrite/tennantite (2.6%) making up the remainder.

Pyrrhotite was not identified in the sample analysed in this study by G&T nor was it identified in the 2011 Lakefield work. Pyrrhotite was identified in a single sample in the 1986 testwork. This may explain the lower levels of cyanide consumption during the G&T testwork compared to testwork conducted during the period 1985-1999. Also, the copper grades in the historic testwork were generally higher than the average of the deposit, which would explain the elevated cyanide consumption observed during those tests.

In the G&T tests copper occurrence, as reported through QEMSCAN, was 0.35%, with iron assaying 5.00%, and sulphur occurring as 3.24%.

13.2 METALLURGICAL TESTWORK RESULTS

13.2.1 Comminution

G&T completed a single Bond ball mill work index on composite sample 12-1B and returned a value of 15.2 kWh/t (metric). This agrees well with a Bond ball mill work index test completed by Lakefield in 1986 on a composite sample which returned a value of 15.4 kWh/t (metric).

G&T also completed a single Abrasion index on composite 12-1B which returned a value of 0.1278 g.

13.2.2 Gravity Concentration

The earlier testwork (1985 to 1989) examined the amenability of the samples to gravity concentration using a Wilfley table with cleaning on a Mozley concentrator. The gravity recovery of gold varied widely in these tests, with gold recovery results ranging from 26 to 52%. Testwork done in 1999 utilized a Knelson concentrator and had gold recovery from 50 to 52%.

Gravity concentration testwork was carried out by G&T in 2012 using a Knelson concentrator. Gold recovery into the gravity concentrate of 81% and 76% into a concentrate of 6% and 5% mass pull was achieved from composite samples 12-1A and 12-1B respectively.

13.2.3 Heavy Media Separation (HMS)

Testwork carried out in 2011 by Lakefield to examine Heavy Media Separation (HMS) as a means of reducing the amount of feed material reporting to the grinding section of the plant by rejecting the waste portion of the plant feed that would come from the mine. The tests indicated that, using this pre-treatment of the plant feed, it is possible to reject 50% of the feed

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material into the waste stream; however, gold loss into the reject material was about 4%. As part of this testwork program, the sink portion was subjected to bulk flotation and cyanidation tests. HMS is not considered for this PEA Study.

13.2.4 Cyanidation

Cyanidation testwork has generally returned very high metal extractions - typically 95% or better for gold and about 80% for silver. A grind of approximately 85% passing 200 mesh (75 µm) and 48 h leach at NaCN 1g/L was generally effective in the earlier testwork.

During the historic testwork, cyanide consumptions in direct cyanidation tests have been variable but generally high. These tests suggested NaCN consumption of between 1 to 2.4 kg/t of feed.

Where solution assays were available, they have shown high Cu and CNS in solution and copper sulphides are most likely the cause of the high cyanide consumption.

Lakefield, as part of the HMS testwork, completed cyanide leach tests on samples of the rougher concentrate and rougher tailings from the sinks portion and from the float portion of the HMS test. Extractions were fairly consistent (~90%) and indicated that there is likely a strong dependency on particle size (independent of grade).

Recent test results from G&T with leach feed having lower copper levels more closely matching the current mine production model has indicated lower cyanide consumption of 0.2- 0.6 kg/t as shown in Table 13.1.

Table 13.1 Gravity Recovery Tailings Cyanidation Testwork

Sample Cyanide Consumption Reference Composite 12-1A 0.6 kg/t G&T June 2012 Composite 12-1B 0.2 kg/t G&T June 2012

13.2.5 Flotation

Several flotation tests have been conducted both historically and as part of the Lakefield and G&T work to produce a copper concentrate and a bulk flotation concentrate. Several tests cyanide leaching tests of the bulk flotation concentrate were completed.

During the recent SGS Lakefield study, the sinks concentrate from the HMS test was submitted for flotation tests. Gold and silver rougher recovery was 99% and 95%, respectively (relative to the flotation feed), into 42% of the mass. This testwork suggested that a relatively coarser grind is likely possible prior to flotation. The flotation results suggest that the gold occurrence is strongly associated with sulphides, which is consistent with the conclusions in the 2004 mineralogy report.

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A copper flotation process option was considered to produce a saleable copper concentrate and to reduce cyanide consumption during the leaching process. Micon notes that some gold would report to the copper concentrate. At reduced copper feed levels this option appears less attractive.

13.3 PROPOSED FUTURE TESTWORK

Future testwork should be considered upon review of the final testwork report currently being prepared by G&T.

Based on a review of reports leading up to the current testwork underway, it is recommended that future testwork should include the following:

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14.0 M IN E R A L R ESO UR C E EST I M A T ES

The mineral resources used to prepare the PEA presented in this report were first published in Hennessey and Mukhopadhyay (2012). The description of the methodology used to prepare the estimate is repeated here.

14.1 MICON MINERAL RESOURCE ESTIMATE

Micon prepared the 2012 estimate of mineral resources for the Curraghinalt deposit using geological information and assay data from 301 drill holes and 277 underground channel samples. This section of the report provides the methodology for estimation of the gold mineral resource.

Primary, or raw, assay data were composited for gold and were analyzed to determine the basic statistical and geostatistical characteristics. This information has been used in several modelling algorithms, which have been compared and checked for validity. A total of 583 specific gravity measurements were utilized.

14.2 GEOLOGICAL MODEL

Several mineralized veins have been identified at Curraghinalt and these have been drilled to different degrees. The present resource estimate encompasses all the known mineralized veins, which are comprised of 12 major veins, 4 secondary veins and 19 veinlets which bifurcate out of the major veins.

For the purpose of resource definition and modelling, each of the identifiable veins was designated with a code. These codes were different from those used previously by Tully for WKH UHVRXUFH HVWLPDWH 7KH DGGLWLRQDO YHLQV LGHQWLILHG VLQFH 7XOO\¶V PLQHUDO UHVRXUFH estimate are the result of further drilling and re-interpretation of the data. A general correlation of the previous and present codes is shown in Table 14.1.

Table 14.1 Previous and Present Codes Used for Mineralized Veins

Previous Present Name Code Road 10 Sheep Dip 20 Mullen 30 T17 H/W 40 T17C 50 T11 55 No. 1 60 106-16 70 ABB D & G

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The details of each mineralized vein, along with the codes used in the mineral resource model, are given in Table 14.2.

Table 14.2 Mineralized Veins and Codes

Primary Secondary Small Veinlets associated with Veins Vein Primary Veins 10 20 30 35 302, 302E 40 45 403, 451E, DL426B 50, 50E 55 552 60 61,62, 601,604,605,606,607 70S 75 70N, 701,703,708, 752 80 DL8290 90 D G

The hanging wall and footwall contacts of each of the veins were calculated from the identified zone intersection from each drill hole. Two dimensional (2D) surfaces were then created for the hanging wall and footwall of each vein, with an additional extension 30 m to 40 m beyond the last contact point derived from the drill hole intersection. A vertical section derived from the hanging wall and footwall wireframes is shown in Figure 14.1. The disposition of the different veins is illustrated schematically in Figure 14.2.

Figure 14.1 Typical Geological Section for the Curraghinalt Area (Section O rientation is North-South Looking West at X = 257437.5 m)

Vertical and horizontal scale shown by grid axes in metres.

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Figure 14.2 Schematic 3D Disposition of Different Mineralized Veins for the Curraghinalt Area (View from Northeast Underneath)

Scale shown by marked Line = 340.0 m.

14.2.1 Basic Statistics

Basic statistics were calculated for gold for each of the individual veins with statistics calculated for the raw, capped and composited samples. The results are presented in Table 14.3 through Table 14.10. The tables show classical statistical parameters (i.e., minimum, maximum, mean and standard deviation) for assays, composites and block model grades. Statistics on linear metal accumulation are similarly given in Table 14.8 and Table 14.10. Both composited samples and estimated blocks are shown in the tables. Linear metal accumulation was calculated for each intersection by multiplying the total length of the samples by grade.

Table 14.3 Classical Statistics for the Mineralized Veins for Raw Assays for Gold from Underground Channel Samples

M inimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 40 620 0.01 451.00 20.00 49.64 60 191 0.01 111.30 11.90 20.40

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Table 14.4 Classical Statistics for the Mineralized Veins for Raw Assays for Gold from Drill Core

M inimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 51 0.01 900.00 32.49 128.62 20 65 0.01 117.44 18.80 27.21 30 102 0.01 204.80 15.05 32.93 40 356 0.01 501.76 19.52 47.80 50 46 0.01 110.40 12.09 20.83 50E 49 0.01 45.93 5.46 10.91 60 357 0.01 246.62 15.86 31.27 70S 327 0.01 221.70 12.95 23.38 80 89 0.01 101.76 15.14 23.61 90 118 0.01 136.41 13.75 24.03 D 65 0.01 133.75 15.32 25.59 G 38 0.01 109.44 17.90 24.74 35 24 0.01 30.70 4.33 9.61 45 15 0.01 27.68 5.25 8.95 55 148 0.01 138.20 9.46 21.50 75 147 0.01 109.44 10.40 17.76 302 25 0.01 29.40 2.63 7.22 302E 12 0.01 64.70 9.49 17.76 403 69 0.01 130.35 4.15 18.87 DL426B 16 0.01 32.00 4.74 8.17 451E 13 0.01 17.26 3.79 6.16 552 27 0.01 25.74 2.98 6.76 61 30 0.01 90.20 9.98 18.51 62 38 0.01 141.25 13.59 33.85 601 25 0.01 59.75 9.90 15.81 604 46 0.01 118.00 11.54 24.58 605 28 0.01 58.92 6.70 13.57 606 30 0.01 39.20 5.83 10.31 607 55 0.01 90.88 5.76 16.87 70N 75 0.01 67.84 9.47 16.11 701 52 0.01 133.21 8.09 20.78 703 19 0.01 54.08 7.92 13.12 708 27 0.01 35.25 5.46 8.84 752 22 0.01 14.80 3.97 5.50 DL8290 8 0.01 25.53 4.75 8.30

Table 14.5 Classical Statistics for the Mineralized Veins, Capped Assays for Gold from Underground Channel Samples

M inimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 40 620 0.01 110.00 15.72 30.90 60 191 0.01 110.00 11.77 20.30

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Table 14.6 Classical Statistics for the Mineralized Veins for Capped Assays for Gold from Drill Core

M inimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 51 0.01 72.00 12.39 21.16 20 65 0.01 50.00 15.23 18.04 30 102 0.01 65.00 11.04 17.78 40 356 0.01 110.00 15.75 27.83 50 46 0.01 70.00 11.21 17.20 50E 49 0.01 45.93 5.46 10.91 60 357 0.01 110.00 14.72 25.63 70S 327 0.01 85.00 12.17 19.07 80 89 0.01 75.00 14.48 21.55 90 118 0.01 70.00 12.34 18.59 D 65 0.01 75.00 13.73 19.45 G 38 0.01 60.00 16.48 20.64 35 24 0.01 30.70 4.33 9.61 45 15 0.01 27.68 5.25 8.95 55 148 0.01 70.00 8.48 17.16 75 147 0.01 85.00 10.23 16.91 302 25 0.01 29.40 2.63 7.22 302E 12 0.01 64.70 9.49 17.76 403 69 0.01 110.00 3.86 16.96 DL426B 16 0.01 32.00 4.74 8.17 451E 13 0.01 17.26 3.79 6.16 552 27 0.01 25.74 2.98 6.76 61 30 0.01 90.20 9.98 18.51 62 38 0.01 110.00 12.27 29.32 601 25 0.01 59.75 9.90 15.81 604 46 0.01 110.00 11.25 23.38 605 28 0.01 58.92 6.70 13.57 606 30 0.01 39.20 5.83 10.31 607 55 0.01 90.88 5.76 16.87 70N 75 0.01 67.84 9.47 16.11 701 52 0.01 85.00 7.16 15.61 703 19 0.01 54.08 7.92 13.12 708 27 0.01 35.25 5.46 8.84 752 22 0.01 14.80 3.97 5.50 DL8290 8 0.01 25.53 4.75 8.30

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Table 14.7 Classical Statistics for the Mineralized Veins for Composited Assays for Gold

M inimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 23 0.01 72.00 12.19 18.39 20 31 0.01 50.00 12.39 12.47 30 57 0.01 65.00 10.42 17.00 40 347 0.01 110.00 13.43 20.17 50 33 0.01 34.50 7.77 10.04 50E 25 0.01 25.31 3.76 5.83 60 233 0.01 110.00 12.00 18.52 70S 143 0.01 85.00 12.90 17.90 80 52 0.01 75.00 14.72 20.42 90 50 0.01 70.00 11.46 15.39 D 39 0.01 53.20 9.44 12.38 G 31 0.01 60.00 13.47 17.98 35 21 0.01 30.70 3.12 8.62 45 10 0.01 13.44 2.93 4.37 55 92 0.01 70.00 6.25 13.72 75 76 0.01 52.48 8.80 12.29 302 20 0.01 16.10 1.45 3.73 302E 5 0.01 13.68 4.69 5.20 403 57 0.01 110.00 3.21 15.63 DL426B 8 0.01 12.64 4.25 4.66 451E 8 0.01 14.24 2.84 4.69 552 18 0.01 8.86 1.45 2.61 61 19 0.01 90.20 11.91 21.68 62 35 0.01 110.00 12.31 30.40 601 20 0.01 59.75 9.96 14.49 604 22 0.01 110.00 11.21 22.82 605 18 0.01 32.00 5.74 9.57 606 25 0.01 27.04 5.18 8.70 607 39 0.01 38.25 2.48 7.08 70N 44 0.01 67.84 5.53 12.66 701 33 0.01 53.17 4.86 11.55 703 9 0.01 12.04 3.51 4.76 708 20 0.01 17.50 3.35 5.23 752 10 0.01 11.38 2.83 3.88 DL8290 6 0.01 25.53 5.23 9.18

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Table 14.8 Classical Statistics for the Mineralized Veins for Composited Assays for Linear Metal Accumulation

Minimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 23 0.00 66.09 8.24 14.34 20 31 0.00 27.28 7.37 8.13 30 57 0.00 42.68 4.67 8.62 40 347 0.00 157.53 18.16 26.65 50 33 0.00 40.88 4.67 9.51 50E 25 0.00 17.71 3.20 4.97 60 233 0.00 792.43 12.77 53.18 70S 143 0.00 111.28 9.06 14.34 80 52 0.00 50.73 5.97 9.47 90 50 0.00 39.63 7.56 10.91 D 39 0.00 39.66 7.59 11.56 G 31 0.00 57.01 7.58 12.51 35 21 0.00 10.87 0.83 2.46 45 10 0.00 5.57 1.38 2.12 55 92 0.00 22.85 2.92 5.55 75 76 0.00 29.80 4.63 7.08 302 20 0.00 12.40 0.91 2.76 302E 5 0.00 17.92 4.05 6.95 403 57 0.00 110.00 3.12 15.75 DL426B 8 0.00 15.67 3.97 5.67 451E 8 0.00 4.01 1.07 1.52 552 18 0.00 8.86 0.83 2.09 61 19 0.00 56.83 5.91 12.93 62 35 0.00 16.50 2.22 4.56 601 20 0.00 20.32 3.46 5.99 604 22 0.00 27.91 5.56 8.44 605 18 0.00 17.33 2.86 5.06 606 25 0.00 9.84 1.26 2.27 607 39 0.00 36.31 2.03 6.42 70N 44 0.00 54.78 4.06 10.07 701 33 0.00 36.75 2.50 6.76 703 9 0.00 34.07 5.42 10.75 708 20 0.00 17.31 2.35 4.32 752 10 0.00 6.78 1.77 2.41 DL8290 6 0.00 6.89 1.55 2.45

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Table 14.9 Classical Statistics for the Mineralized Veins for Estimated Blocks for Gold

Minimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 88810 0.25 72.00 22.00 15.15 20 189495 0.01 47.01 10.48 7.31 30 267485 0.01 65.00 12.15 13.89 40 553296 0.01 107.98 9.98 10.94 50 23246 0.01 34.50 8.26 8.30 50E 27358 0.01 25.30 4.42 4.25 60 684408 0.01 110.00 14.55 16.51 70S 612559 0.01 85.00 12.14 12.66 80 567767 0.01 75.00 18.75 17.56 90 548826 0.01 70.00 12.27 11.03 D 34315 0.01 53.14 9.89 7.70 G 32903 0.01 60.00 14.41 15.23 35 16034 0.01 30.70 12.43 13.48 45 14095 0.01 13.44 4.38 3.57 55 70886 0.01 70.00 6.17 8.62 75 555991 0.01 52.48 11.74 11.40 302 26423 0.01 16.10 3.12 3.44 302E 13347 0.01 13.68 7.44 4.52 403 11413 0.01 110.00 4.20 12.92 DL426B 36715 0.66 11.98 6.38 0.95 451E 16615 0.01 14.24 3.62 4.40 552 12945 0.01 6.30 2.75 1.85 61 18243 0.01 90.20 21.31 25.43 62 28315 0.01 110.00 13.55 20.12 601 28112 0.01 59.75 12.01 13.84 604 34356 0.01 110.00 10.03 17.03 605 39483 0.01 27.80 8.16 5.89 606 28442 0.01 27.04 6.69 7.01 607 38252 0.01 38.25 3.27 5.69 70N 474674 0.01 67.84 9.20 7.81 701 44053 0.01 53.17 4.72 9.06 703 7871 0.01 12.04 5.59 4.84 708 21576 0.01 17.50 5.54 4.65 752 10149 0.01 11.38 4.19 3.76 DL8290 26036 0.01 25.53 4.28 7.41

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Table 14.10 Classical Statistics for the Mineralized Veins for Estimated Blocks for Linear Metal Accumulation

Minimum Maximum Mean Standard Deviation Vein No. of Samples (g/t Au) (g/t Au) (g/t Au) (g/t Au) 10 88810 0.18 59.28 11.17 5.92 20 189495 0.00 27.28 6.06 5.67 30 267485 0.00 42.68 4.38 5.94 40 553296 0.00 157.50 10.09 11.79 50 23246 0.00 40.88 4.91 7.49 50E 27358 0.00 17.71 3.63 4.05 60 684408 0.00 791.00 12.95 24.78 70S 612559 0.00 111.24 7.71 7.95 80 567767 0.00 26.10 5.86 5.70 90 548826 0.00 39.21 7.21 7.34 D 34315 0.00 38.98 9.01 9.77 G 32903 0.00 33.89 7.42 7.84 35 16034 0.00 10.87 3.14 3.99 45 14095 0.00 5.57 2.48 2.05 55 70886 0.00 22.04 3.77 4.56 75 555991 0.00 24.81 4.27 4.82 302 26423 0.00 12.40 1.68 2.52 302E 13347 0.00 17.92 7.73 6.76 403 11413 0.00 110.00 4.33 13.12 DL426B 36715 0.82 14.86 6.67 3.37 451E 16615 0.00 4.01 1.40 1.41 552 12945 0.00 4.56 1.38 1.18 61 18243 0.00 56.83 12.45 16.23 62 28315 0.00 16.08 3.42 4.10 601 28112 0.00 20.31 4.72 6.19 604 34356 0.00 27.91 5.68 7.62 605 39483 0.00 17.33 3.89 3.45 606 28442 0.00 9.84 1.64 1.71 607 38252 0.00 36.31 3.16 6.59 70N 474674 0.00 54.78 10.27 13.46 701 44053 0.00 36.75 3.06 5.66 703 7871 0.00 34.07 8.52 10.05 708 21576 0.00 17.31 4.74 5.23 752 10149 0.00 6.78 2.46 2.10 DL8290 26036 0.00 6.89 1.33 1.97

14.2.2 Raw Assay Intervals

Samples were selected within the mineralized wireframe for all the veins. The statistical analysis and the resource estimation were carried out using only these samples.

A study on capping of high grade gold assays was carried out. There are not enough samples for each sub-zone (veinlets) to apply capping individually. All the samples from each veinlet were included with the samples from the major veins in order to assess capping for that particular vein.

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While a few of the major veins do not have enough samples to create a probability plot, these plots were created for almost all of the major veins and a few of the sub-zones. The inflection point in the probability plot indicates the break in continuity of the distribution for each vein. Capping was applied since these inflections indicate the beginning of the outlier population for the respective veins. The details of grade capping including percentile at which capping has been applied along with number of samples affected by capping for each vein are given in Table 14.11. The probability plots are shown in Figure 14.3 through Figure 14.13.

Table 14.11 Grade Capping Including Number of Samples Affected and Percentile at Which Capping is Applied

Vein Total Number of Capping No. of Samples Percentile at Which Code Samples (g/t Au) Affected By Capping Capping is Applied 10 51 72 4 92% 20 65 50 7 89% 30 102 65 7 93% 40 356 110 11 97% 50 46 70 1 98% 50E 49 70 0 100% 60 357 110 9 97% 70S 327 85 6 98% 80 89 75 4 96% 90 118 70 6 95% D 65 75 3 95% G 38 60 2 95% 35 24 65 0 100% 45 15 110 0 100% 55 148 70 5 97% 75 147 85 2 99% 302 25 65 0 100% 302E 12 65 0 100% 403 69 110 1 99% DL426B 16 110 0 100% 451E 13 110 0 100% 552 27 70 0 100% 61 30 110 0 100% 62 38 110 3 92% 601 25 110 0 100% 604 46 110 2 96% 605 28 110 0 100% 606 30 110 0 100% 607 55 110 0 100% 70N 75 85 0 100% 701 52 85 1 98% 703 19 85 0 100% 708 27 85 0 100% 752 22 85 0 100% DL8290 8 75 0 100% 40 (UG) 620 110 34 95% 60 (UG) 191 110 1 99%

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Figure 14.3 Probability Plot for Vein 10 of Curraghinalt Deposit

100000

10000

1000

100 Capping 72 g/t ) u 10 A t / g (

1 e d a r 0.1 G 0.01

0.001

0.0001

0.00001 -4 -3 -2 -1 0 1 2 3 4

Cumulative Normal Distribution Function

Figure 14.4 Probability Plot for Vein 20 of Curraghinalt Deposit

100000

10000

1000 Capping 50 g/t 100 ) u A

t 10 / g ( e d 1 a r G 0.1

0.01

0.001

0.0001 -3 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

148

Figure 14.5 Probability Plot for Vein 30 of Curraghinalt Deposit

100000

10000

1000

100 Capping 65 g/t ) u A

t 10 / g ( e d 1 a r G 0.1

0.01

0.001

0.0001 -3 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

Figure 14.6 Probability Plot for Vein 40 of Curraghinalt Deposit

100000

10000

1000 Capping 110 g/t 100 ) u A

t 10 / g ( e d 1 a r G 0.1

0.01

0.001

0.0001 -3 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

149

Figure 14.7 Probability Plot for Vein 50 of Curraghinalt Deposit

10000

1000

100 Capping 70 g/t ) u A

10 t / g ( e d

a 1 r G

0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

Figure 14.8 Probability Plot for Vein 60 of Curraghinalt Deposit

100000

10000

1000 Capping 110 g/t )

u 100 A t / g (

10 e d a r

G 1

0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

150

Figure 14.9 Probability Plot for Vein 70 of Curraghinalt Deposit

100000

10000

1000

) Capping 85 g/t u 100 A t / g (

10 e d a r

G 1

0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

Figure 14.10 Probability Plot for Vein 80 of Curraghinalt Deposit

100000

10000

1000

) Capping 75 g/t

u 100 A t / g (

10 e d a r

G 1

0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

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Figure 14.11 Probability Plot for Vein 90 of Curraghinalt Deposit

100000

10000

1000

) Capping 70 g/t

u 100 A t / g (

10 e d a r

G 1

0.1

0.01

0.001 -2 -1 0 1 2 3 Cumulative Normal Distribution Function

Figure 14.12 Probability Plot for Vein D of Curraghinalt Deposit

1000

100 Capping 75 g/t

10 ) u A

t / g (

1 e d a r G 0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

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Figure 14.13 Probability Plot for Vein G of Curraghinalt Deposit

1000

100 Capping 60 g/t

10 ) u A

t / g (

1 e d a r G 0.1

0.01

0.001 -2 -1 0 1 2 3

Cumulative Normal Distribution Function

14.2.3 Compositing

Compositing is the first step in mineral resource estimation. The intersection defined by each drill hole for each zone represents the thickness of the quartz vein for that zone. In places, the width of the quartz vein is as low as 10 cm. It is not possible to carry out uniform length compositing with a length of 10 cm, since it would generate a huge number of artificial samples which is not supported by the theory of compositing. A larger composite length on the other hand would dilute the grade of the vein. Under such circumstances, it is better to group all the samples from a single drill intersection through a vein into one composite. The problem of variable composite length may be overcome by estimating linear metal accumulation rather than grade for the purpose of interpolation.

Compositing was carried out based on the above method. Linear metal accumulation was calculated for each intersection by multiplying the total length of the samples by the grade. The classical statistics for linear metal accumulation are given in Table 14.8, above.

14.3 GRADE INTERPOLATION METHOD

14.3.1 Estimation Parameters and Search Distances

The density of drilling in the principal zones of mineralization is generally 30 m by 30 m. This drill density, together with good understanding of the geology and mineralization, provides a reasonable level of confidence in the resource estimate. The method of

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compositing resulted in a single sample point per intersection. This results in a two- dimensional estimation protocol. The method of compositing has resulted in reduction of the nugget effect to a minimum (see semi-variogram in Figure 14.19). A very low nugget effect prompted the selection of the inverse distance method of estimation. The power was raised to 5 so that the variation of grade remains local. The analysis of classical statistics and compositing has dictated the selection of the estimation parameters. The orientation of the search ellipse for all the veins along with the estimation parameters are given in Table 14.12 through to Table 14.14.

Table 14.12 Summary of Search Volume Reference Number, Estimation Method and Vein

Search Search Estimation Volume Estimation Vein Volume Power Vein Power Method Reference Method Reference No. No. 10 1 Inverse Distance 5 601 19 Inverse Distance 5 20 2 Inverse Distance 5 604 20 Inverse Distance 5 30 3 Inverse Distance 5 605 21 Inverse Distance 5 35 4 Inverse Distance 5 606 22 Inverse Distance 5 302 5 Inverse Distance 5 607 23 Inverse Distance 5 302E 6 Inverse Distance 5 70N 24 Inverse Distance 5 40 7 Inverse Distance 5 70S 25 Inverse Distance 5 45 8 Inverse Distance 5 75 26 Inverse Distance 5 403 9 Inverse Distance 5 701 27 Inverse Distance 5 451E 10 Inverse Distance 5 703 28 Inverse Distance 5 DL426B 11 Inverse Distance 5 708 29 Inverse Distance 5 50 12 Inverse Distance 5 752 30 Inverse Distance 5 50E 13 Inverse Distance 5 80 31 Inverse Distance 5 55 14 Inverse Distance 5 DL8290 32 Inverse Distance 5 552 15 Inverse Distance 5 90 33 Inverse Distance 5 60 16 Inverse Distance 5 D 34 Inverse Distance 5 61 17 Inverse Distance 5 G 35 Inverse Distance 5 62 18 Inverse Distance 5

Table 14.13 Summary of Search Parameters

Parameter Value Minimum No. of Samples 3 Maximum No. of Samples 5 Search Enlargement Factor 2 Minimum No. of Samples 3 Maximum No. of Samples 5 Search Enlargement Factor 10 Minimum No. of Samples 1 Maximum No. of Samples 5

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Table 14.14 Summary of Search Volume Reference Number, Direction and Angle of Rotation

Search Search Search Search Angle of Angle of Angle of Volume Along Along Along Axis of Axis of Axis of Rotation Rotation Rotation Reference X - Z - Y - Rotation Rotation Rotation (o) (o) (o) No. Direction Direction Direction 1 30 30 250 25 Z -30 X 0 Y 2 30 30 250 20 Z -35 X 0 Y 3 30 30 250 20 Z -25 X 0 Y 4 30 30 250 15 Z -20 X 0 Y 5 30 30 250 25 Z -25 X 0 Y 6 30 30 250 5 Z -25 X 0 Y 7 30 30 250 10 Z -30 X 0 Y 8 30 30 250 10 Z -25 X 0 Y 9 30 30 250 0 Z -25 X 0 Y 10 30 30 250 20 Z -40 X 0 Y 11 30 30 250 10 Z -20 X 0 Y 12 30 30 250 10 Z -30 X 0 Y 13 30 30 250 15 Z -50 X 0 Y 14 30 30 250 20 Z -40 X 0 Y 15 30 30 250 10 Z -45 X 0 Y 16 30 30 250 10 Z -35 X 0 Y 17 30 30 250 10 Z -25 X 0 Y 18 30 30 250 15 Z -30 X 0 Y 19 30 30 250 15 Z -45 X 0 Y 20 30 30 250 10 Z -40 X 0 Y 21 30 30 250 15 Z -30 X 0 Y 22 30 30 250 10 Z -30 X 0 Y 23 30 30 250 15 Z -35 X 0 Y 24 30 30 250 20 Z -25 X 0 Y 25 30 30 250 10 Z -45 X 0 Y 26 30 30 250 10 Z -40 X 0 Y 27 30 30 250 -5 Z -45 X 0 Y 28 30 30 250 -5 Z -40 X 0 Y 29 30 30 250 20 Z -35 X 0 Y 30 30 30 250 0 Z -30 X 0 Y 31 30 30 250 5 Z -45 X 0 Y 32 30 30 250 10 Z -45 X 0 Y 33 30 30 250 10 Z -45 X 0 Y 34 30 30 250 20 Z -20 X 0 Y 35 30 30 250 20 Z -25 X 0 Y

14.4 BLOCK MODEL

The block model constructed for the Curraghinalt resource estimate utilized rectangular blocks measuring 10 m (X) by 2,000 m (Y) by 5 m (Z) in height. Although the dimension in the Y-direction is unusually long, this was selected with the understanding of the nature of the transverse sections of the mineralized zones. In order to better conform to the mineralization contacts, Datamine software uses a system of sub-blocking. The sub-blocking method used in Datamine would result in a single block in the Y-direction conforming to the width of the vein and is in line with the method of compositing. Blocks were permitted to split in the X and Z directions. This procedure minimizes the volume variance between the wireframe model and the block model. This block size was the most appropriate considering the morphology of the

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mineralization and the distribution of sample data. The parameters that describe the block model are summarized in Table 14.15.

Table 14.15 Block Model Parameters

Block Dimension Direction O rigin1 No. of Blocks No. of Sub-Blocks (m) X 256300 E 10 210 8 Y 385500 N 2,000 1 1 Z -630 5 210 8 1 Origin of X and Y axes based on Irish Transverse Mercator Grid; origin of Z axis is at -630 m, is altitude from mean sea level.

Block model grade interpolation was performed for gold using inverse distance to the power of 5 in different zones. Concentric search ellipses were used to avoid smearing of grade and for preservation of local variation. Estimations were carried out for both gold, length of the samples and linear metal accumulation. The grades were recalculated from linear metal accumulation by dividing it by the estimated length. This recalculated grade was used for resource reporting and was done so in order to carry through the procedure adopted during compositing.

14.5 BLOCK MODEL VALIDATION

Validation of the block model included the following:

1. Visual inspection, 2. Comparison of individual block grades with the de-clustered sample grades at two different block sizes and 3. Alternate estimation method.

14.5.1 Visual Inspection

The block model was compared with the drill hole intersection at uniform intervals both vertically and along cross section. No visible bias was observed. The cross-section, plan view on 170 m RL (the 170 m level, 60 m below surface) and a three-dimensional view of the block model are shown in Figure 14.14 through 14.16.

14.5.2 Comparison of individual block grades with de-clustered sample grade

The block model grade interpolation protocol was investigated by inspecting plots of block model grades versus mean borehole sample composite grades that occur within each block. Mean borehole sample composite grades were calculated for each block using the declustering technique in which the weighted average grade of composites that fall within a block is calculated. This value is compared with the grade interpolated for the block. A successful grade interpolation protocol will result in block grade estimates that demonstrate a minimum amount of bias. Declustering was carried out on block dimensions shown above in

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Table 14.15 and with a block dimension of 30 m (X) and 30 m (Z). This dimension was selected as it represents the drilling density.

Plots comparing composite grades and block model grades for all the veins are presented in Figure 14.17 and Figure 14.18. It is apparent from the plots that there is no significant bias. The de-clustering analysis demonstrates that the mineral resource model provides a reasonable estimate of Curraghinalt mineral resources.

14.5.3 Alternate Estimation Method

An experimental semi-variogram was calculated using the samples of vein 60 for linear metal accumulation. The variogram model developed, based on the experimental semi-variogram, was used to estimate all the veins using ordinary kriging (Figure 14.19). All other parameters including dimension and orientation of search ellipse and number of samples used for estimation were kept constant. A scatter plot was produced to compare the grade of two methods of estimation, as shown in Figure 14.20. The plot indicates that there may be local variation based on the method of estimation but, globally, they will produce similar results.

Based on these three validation checks, it was concluded that the resource model represents a reasonable grade estimate of the deposit when compared to the drill hole samples.

Figure 14.14 North-south Cross-section Showing Block Model Grades for Mineralized Veins for Curraghinalt Deposit (Section Looking West at X = 257437.5 m)

Vertical and horizontal scale shown by grid axes in metres.

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Figure 14.15 Plan View at 170 m RL of the Block Model Categorized into Different Grades for Curraghinalt Deposit

Vertical and horizontal scale shown by grid axes in metres.

Figure 14.16 Schematic 3D Rendering of Block Model Categorized into Different Grades for Curraghinalt Deposit (View from Northeast Underneath)

Scale shown by marked Line = 340.0 m.

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Figure 14.17 Comparison of De-clustered Drill Hole Data with Resource Model Block Grade for Curraghinalt Deposit (Block Dimension of 10 m x width of vein x 5 m)

100 ) u A

t / g (

a t a d

l 10 e d o M

k c o l B

e 1 t a m i t s E

0.1 0.1 1 10 100

Declustered Drill Hole Data (g/t Au)

Figure 14.18 Comparison of De-clustered Drill Hole Data with Resource Model Block Grade for Curraghinalt Deposit (Block Dimension of 30 m x width of vein x 30 m)

100 ) u A

t / g (

a t a D

l 10 e d o M

k c o l B

d 1 e t a m i t s E

0.1 0.1 1 10 100

Declustered Drill Hole Data (g/t Au)

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Figure 14.19 Omni-directional Semi-variogram Model Using Data from Vein 60

Figure 14.20 Comparison of Two Methods of Estimation

) 100 u A t / g ( g n i g i r K

y 10 r a n i d r O g n i s U

n 1 o i t a m i t s E e d a r

G 0.1 0.1 1 10 100

Grade Estimation Using Inverse Distance with a Power of 5 (g/t Au)

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14.6 MINERAL RESOURCE CLASSIFICATION

Mineral resources were estimated following the CIM guidelines. The following definitions were adopted for the classification of mineral resources.

Measured mineral resources are defined as those portions of the deposit which has underground channel sample within a 15-m radius.

Indicated mineral resources are defined as those portions of the deposit estimated with a drill spacing generally defined by 30 m by 30 m and high level of confidence on geological continuity of mineralization. The limit of indicated resources was considered to be 30 m from the last intersecting drill hole with the drill spacing above.

Inferred mineral resources are defined as those portions of the deposit for which grade is interpolated utilizing a wider drill spacing, at places with a 100-m radius, or fewer intersections but with a high level of confidence on the geological continuity of the mineralization. The limit of the inferred resource was considered to be 30 m from the last drill hole on the periphery.

Only Veins 40 and 60 were explored by one level of underground development. Measured mineral resources are from only these two veins. Most of the major veins have been classified as indicated and inferred resources. Veins 80, 90, D and G have been classed as inferred since there is insufficient drilling on each for them to be included in the indicated category. Sub-zones 55 and 75 have been classified as indicated and inferred resources respectively. All the other sub-zones have been classified as inferred resources. All the veinlets have been classified as inferred except Veins 61, 62, 604 and 701 where, at the upper levels, there is sufficient drilling to classify some portions as indicated resources.

A vertical longitudinal projection of Vein 60 is given in Figure 14.21 to demonstrate the method of resource classification. Figure 14.22 shows a cross-section of the resource model, colour-coded by category of resource. A plan view (170 m RL) is also presented in Figure 14.23, with classification of the mineral resource. A three-dimensional view of the model in presented in Figure 14.24.

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Figure 14.21 Vertical Longitudinal Projection of Vein 60 (Looking North-East)

Circular dots indicate the drill intersection. The red line indicates the boundary of measured resource, green line indicates indicated resource and the blue line indicates the boundary of inferred Resource. Vertical and horizontal scale shown by grid axes in metres.

Figure 14.22 North-South Cross-Section Showing Resource Classifications for Curraghinalt Deposit (Section Looking West at X = 257437.5 m)

Vertical and horizontal scale shown by grid axes in metres.

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Figure 14.23 Plan View at 170 m RL Showing Resource Classifications for the Curraghinalt Deposit

Vertical and horizontal scale shown by grid axes in metres.

Figure 14.24 Schematic 3D View of the Block Model for the Curraghinalt Deposit (View Northeast Underneath)

Scale shown by marked line = 340.0 m.

The grade-tonnage distribution of the Curraghinalt block model is presented in Figure 14.25 and Figure 14.26. It is apparent that the deposit maintains a significant profile of measured and indicated mineral resources at higher cut-off grades. The estimated mineral resources at different cut-off grades and vein thickness are shown in Table 14.16 and Table 14.17. It is evident from the grade tonnage curves and the tables that there is little resource associated with blocks of vein thickness less than 0.1 m.

163

Figure 14.25 Grade-Tonnage Distribution of the Measured and Indicated Mineral Resource at Different Thickness Cut-off

Figure 14.26 Grade-Tonnage Distribution of the Inferred Mineral Resource at Different Thickness Cut-off

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Table 14.16 Mineral Resource Estimate at Different Grade and Thickness Cut-offs O riginal Vein Thickness Cut-off

O riginal Measured Resource Indicated Resource Measured + Indicated Resource Inferred Resource Vein G rade Contained Contained Contained Metal Contained Metal Thickness Cut-off Million G rade Metal Million G rade Metal Million G rade Million G rade Cut-off (g/t Au) Tonnes (g/t Au) Million Tonnes (g/t Au) Million Tonnes (g/t Au) Million Tonnes (g/t Au) Million Tonnes Tonnes Tonnes Tonnes (m) Oz Oz Oz Oz 0.0 0.0 0.03 14.54 0.46 0.01 2.11 7.57 15.99 0.51 2.14 7.67 16.45 0.53 11.59 7.03 81.47 2.62 0.0 1.0 0.03 17.67 0.46 0.01 1.70 9.33 15.88 0.51 1.73 9.45 16.34 0.53 9.45 8.55 80.80 2.60 0.0 2.0 0.02 18.96 0.46 0.01 1.52 10.25 15.61 0.50 1.55 10.39 16.07 0.52 8.22 9.61 78.96 2.54 0.0 3.0 0.02 19.87 0.45 0.01 1.37 11.11 15.24 0.49 1.39 11.25 15.69 0.50 7.25 10.57 76.56 2.46 0.0 4.0 0.02 20.73 0.45 0.01 1.22 12.08 14.70 0.47 1.24 12.23 15.15 0.49 6.29 11.64 73.24 2.35 0.0 5.0 0.02 21.51 0.44 0.01 1.11 12.84 14.20 0.46 1.13 13.00 14.65 0.47 5.46 12.73 69.48 2.23 0.0 6.0 0.02 22.35 0.44 0.01 0.97 13.86 13.45 0.43 0.99 14.03 13.89 0.45 4.77 13.78 65.73 2.11 0.0 7.0 0.02 23.24 0.43 0.01 0.81 15.28 12.45 0.40 0.83 15.46 12.88 0.41 4.04 15.11 60.97 1.96 0.0 8.0 0.02 24.04 0.43 0.01 0.69 16.68 11.52 0.37 0.71 16.86 11.94 0.38 3.45 16.39 56.55 1.82 0.0 9.0 0.02 24.82 0.42 0.01 0.59 18.15 10.62 0.34 0.60 18.34 11.04 0.36 3.05 17.43 53.16 1.71 0.0 10.0 0.02 25.65 0.41 0.01 0.52 19.31 9.97 0.32 0.53 19.50 10.38 0.33 2.59 18.85 48.82 1.57 0.1 0.0 0.03 14.59 0.46 0.01 2.11 7.57 15.99 0.51 2.14 7.68 16.45 0.53 11.44 7.10 81.29 2.61 165 0.1 1.0 0.03 17.67 0.46 0.01 1.70 9.33 15.88 0.51 1.73 9.46 16.34 0.53 9.38 8.60 80.66 2.59

0.1 2.0 0.02 18.96 0.46 0.01 1.52 10.26 15.61 0.50 1.55 10.39 16.07 0.52 8.19 9.64 78.87 2.54 0.1 3.0 0.02 19.87 0.45 0.01 1.37 11.11 15.24 0.49 1.39 11.25 15.69 0.50 7.23 10.58 76.51 2.46 0.1 4.0 0.02 20.73 0.45 0.01 1.22 12.08 14.70 0.47 1.24 12.23 15.15 0.49 6.29 11.64 73.22 2.35 0.1 5.0 0.02 21.51 0.44 0.01 1.11 12.84 14.20 0.46 1.13 13.00 14.65 0.47 5.45 12.74 69.44 2.23 0.1 6.0 0.02 22.35 0.44 0.01 0.97 13.86 13.45 0.43 0.99 14.03 13.89 0.45 4.77 13.78 65.73 2.11 0.1 7.0 0.02 23.24 0.43 0.01 0.81 15.28 12.45 0.40 0.83 15.46 12.88 0.41 4.04 15.11 60.97 1.96 0.1 8.0 0.02 24.04 0.43 0.01 0.69 16.68 11.52 0.37 0.71 16.86 11.94 0.38 3.45 16.39 56.55 1.82 0.1 9.0 0.02 24.82 0.42 0.01 0.59 18.15 10.62 0.34 0.60 18.34 11.04 0.36 3.05 17.43 53.16 1.71 0.1 10.0 0.02 25.65 0.41 0.01 0.52 19.31 9.97 0.32 0.53 19.50 10.38 0.33 2.59 18.85 48.82 1.57 1.0 0.0 0.02 18.05 0.40 0.01 1.21 8.61 10.42 0.34 1.23 8.78 10.83 0.35 4.03 9.33 37.63 1.21 1.0 1.0 0.02 20.13 0.40 0.01 1.04 10.02 10.39 0.33 1.06 10.21 10.79 0.35 3.39 11.07 37.51 1.21 1.0 2.0 0.02 21.14 0.40 0.01 0.97 10.62 10.28 0.33 0.99 10.83 10.68 0.34 3.19 11.67 37.21 1.20 1.0 3.0 0.02 21.71 0.40 0.01 0.90 11.19 10.12 0.33 0.92 11.40 10.52 0.34 3.05 12.09 36.86 1.19 1.0 4.0 0.02 22.28 0.40 0.01 0.82 11.99 9.84 0.32 0.84 12.21 10.23 0.33 2.83 12.74 36.11 1.16 1.0 5.0 0.02 22.89 0.39 0.01 0.77 12.49 9.60 0.31 0.79 12.72 10.00 0.32 2.64 13.35 35.24 1.13 1.0 6.0 0.02 23.53 0.39 0.01 0.69 13.31 9.15 0.29 0.70 13.56 9.54 0.31 2.42 14.05 34.07 1.10 1.0 7.0 0.02 24.23 0.39 0.01 0.58 14.65 8.43 0.27 0.59 14.91 8.82 0.28 2.08 15.31 31.80 1.02 1.0 8.0 0.02 24.86 0.38 0.01 0.48 16.03 7.73 0.25 0.50 16.30 8.11 0.26 1.76 16.72 29.38 0.94 1.0 9.0 0.01 25.53 0.38 0.01 0.40 17.59 7.03 0.23 0.41 17.87 7.40 0.24 1.55 17.83 27.61 0.89 1.0 10.0 0.01 26.20 0.37 0.01 0.35 18.73 6.56 0.21 0.36 19.03 6.93 0.22 1.25 19.88 24.75 0.80

Table 14.17 Mineral Resource Estimate at 5.0 g/t Au Cut-off Grade With Different Vein Thickness

O riginal Vein Contained Metal Resource M illion G rade Thickness Cut-off Million Category Tonnes (g/t Au) Tonnes (m) Oz 0.0 0.02 21.51 0.44 0.01 Measured 0.1 0.02 21.51 0.44 0.01 1.0 0.02 22.89 0.39 0.01 0.0 1.11 12.84 14.20 0.46 Indicated 0.1 1.11 12.84 14.20 0.46 1.0 0.77 12.49 9.60 0.31 0.0 1.13 13.00 14.65 0.47 Measured + 0.1 1.13 13.00 14.65 0.47 Indicated 1.0 0.79 12.72 10.00 0.32 0.0 5.46 12.73 69.48 2.23 Inferred 0.1 5.45 12.74 69.44 2.23 1.0 2.64 13.35 35.24 1.13

14.7 MINERAL RESOURCE ESTIMATE

Micon has considered the technical and economic criteria used to estimate a reasonable gold cut-off grade for reporting of the mineral resources. Using these criteria, Micon has produced a preliminary estimate for a break-even cut-off grade for reporting mineralization at Curraghinalt considering all factors, including the accessibility and infrastructure. Original vein thickness blocks of less than 0.10 m were eliminated from the resource tabulation. There were minor differences between a 0.00-m and 0.10-m minimum thickness which disappear with rounding. For horizontal vein thicknesses of more than 0.10 m, but less than1.0 m, the grade was diluted to a minimum width of 1.0 m prior to reporting.

It was concluded that 5 g/t Au is an appropriate cut-off grade for mineral resource reporting. The mineral resource estimate was updated with the surface topography (veins close to surface were wireframed and modeled beyond the surface). There are two underground drives, one on Vein 40 and other on Vein 60. The total volume of mined out material is small compared to the volume of these veins and are well within the rounding error of estimate. The resources as a result have not been depleted for these two workings.

Based on a cut-off grade of 5 g/t Au and the considerations set out above, the measured mineral resource totaled 0.02 million tonnes at an average grade of 21.51 g/t Au and the indicated mineral resources totaled 1.11 million tonnes at an average grade of 12.84 g/t Au. The inferred mineral resources totaled 5.45 million tonnes at an average grade of 12.74 g/t Au. The mineral resources, as first published in Hennessey and Mukhopadhyay (2012), are summarized in Table 14.18 below.

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Table 14.18 Classified Mineral Resource Estimate at 5 g/t Gold Cut-off Grade and 0.1 m Minimum Width, Diluted to a Minimum Width of 1.0 m

After the public disclosure of the current Curraghinalt mineral resource estimate on November 30, 2011, and its publication in Hennessey and Mukhopadhyay (2012), silver and copper grades were estimated into the block model. This was done to allow for the contribution of silver to the cash flow for the PEA and to model FRSSHU¶VHIIHFWRQcyanide consumption. Silver and copper grades were estimated using the same techniques as described for gold above. No copper concentrate is produced and there is no copper revenue in the base case cash flow. Silver accounts for approximately 1% of revenue in the economic analysis.

The amended Curraghinalt deposit mineral resource estimate showing copper and silver grades is set out in Table 14.19 below. No changes in tonnages or gold grades have occurred. In order to make it comparable to Table 14.18, Table 14.19 is also reported at a cut-off grade of 5.0 g/t Au. No gold equivalent calculations were made.

Table 14.19 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper

Million Au Ag Cu Category Notes Tonnes (g/t) (g/t) (%) Measured 0.02 21.51 17.56 0.49 Indicated 1.11 12.84 4.05 0.17 Measured + 1.13 13.00 4.29 0.18 Indicated With Cu* 5.38 12.77 5.48 0.12 Inferred Without Cu* 0.08 10.37 2.00 - * Copper assay data were not available for the 752 and D veins and no copper grades were estimated there. 1 - Original vein thickness blocks of less than 0.10 m were eliminated from the resource tabulation. There were minor differences between a 0.00-m and 0.10-m minimum thickness estimates which disappear with rounding. For horizontal vein thicknesses of more than 0.10 m, but less than 1.0 m, the grades were diluted to a minimum horizontal width of 1.0 m at zero grade prior to reporting. 2 - Mineral resources are reported at a cut-off grade of 5 g/t Au after dilution. 3 - Mineral resources are not mineral reserves and do not have demonstrated economic viability. 4 - The mineral resource estimate was prepared by Dibya Kanti Mukhopadhyay, MAusIMM (CP), under the overall direction of B. Terrence Hennessey, P.Geo. Mr. Hennessey has taken responsibility for the estimate. The Curraghinalt resource estimate is compliant with the current standards and definitions required under NI 43-101.

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To the best knowledge of the author the stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues, unless stated elsewhere in this report. There are no known mining, metallurgical, infrastructure or other factors that materially affect this mineral resource estimate.

The mineral resource estimate presented herein was originally completed by Dibya Kanti Mukhopadhyay, MAusIMM (CP) under the overall direction of B. Terrence Hennessey, P.Geo. In the absence of Mr. Mukhopadhyay, Mr. Hennessey has accepted responsibility for the estimate for this PEA and report. It is compliant with the current standards and definitions required under NI 43-101 and is, therefore, reportable as a mineral resource by Dalradian Resources.

This mineral resource was estimated from a database which was frozen on October 10, 2011 and is current as of November 30, 2011.

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15.0 M IN E R A L R ESE R V E EST I M A T ES

As there has been no prefeasibility or feasibility study completed to date, no mineral reserve estimates have been determined for the Tyrone project or the Curraghinalt deposit at this time.

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16.0 M ININ G M E T H O DS

16.1 INTRODUCTION

The mineralized veins at the Curraghinalt deposit will be extracted using the Longitudinal Sublevel Retreat Underground Mining method with both paste and waste backfill. This backfill method was chosen in order to maximize the extraction of the high value resources at the Curraghinalt Project, reduce the surface footprint required for tailings disposal, and allow savings in the volume and transportation cost of waste rock from underground development.

The Curraghinalt deposit is comprised of a swarm of gold-bearing quartz veins. The veins range from a few centimetres wide to over 3 m wide and are aligned along a west-northwest trend. Average geological width is reported to be approximately 1.2 m. The strike length of the vein swarm at Curraghinalt has been traced by prospecting, trenching and drilling to a minimum length of 1,950 m and remains open to the east and west. The veins dip between 60º and 75º to the north and have been traced to a depth of approximately 840 m below surface (Hennessey and Mukhopadhyay, January, 2012).

The extraction of mineralized material from underground will be carried out with rubber tired mechanized equipment to maximize the production and provide flexibility to the underground operations.

16.2 MINE DESIGN PARAMETERS

The following summarizes the preliminary key design parameters and assumptions for the proposed mining method and extraction of mineralized material at Curraghinalt:

General Parameters: Pre-production period with mine development in waste and mineralized material duration is one year. Production ramp-up at 50%, 75% and 100% from year -1, 1 and 2 respectively. This is equivalent to 310,500 tonnes per year (t/y) in year -1, 465,375 t/y in year 1 followed by full production at 620,500 t/y by year 2. Total mine life (excluding pre-production, year -1) is 15.0 years. Mine production rate of 1,700 tonnes per day (t/d) at full production. 365 operating days per year (d/y). Underground operators working three 8 hour shifts per day (s/d). Specific gravity for waste material is 2.7 (assumed). Specific gravity for mineralized material is 2.85 (average).

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Underground mining: Mining Method is longitudinal sublevel retreat. Decline or ramp dimensions are 4.5 m H x 4.5 m W at 15% grade. By-passes at the decline or ramp at 4.5 m H x 4.5 m W x 3.0 m L. Safety bays along the decline and ramp at 3.0 m H x 3.0 m W x 3.0 m L. Sumps at 3.0 m H x 4.0 m W x 3.0 m L. Cross-cut dimension at 4.0 m H x 4.0 m W. Sill or development in ore at 3.0 m H x 3.0 m W. Refuge station/lunch room at 4.0 m H x 4.0 m W x 10.0 m L. Drift to ventilation shafts and remuck bays at 4.0 m H x 4.0 m W. Ventilation raise and ore or waste-pass at 3 m diameter. Level intervals at 20 vertical metres from the floor of the top sill to the floor of the bottom sill.

16.2.1 Geotechnical Parameters

In November 2011, Snowden performed a site visit to Curraghinalt to assess the geotechnical core logging process carried on the project and performed a scan line mapping of the exploration adit and drifts totaling 65 m to obtain an overall preliminary assessment the site geotechnical conditions (Snowden, January 2012).

Snowden reviewed the geotechnical data collected by Dalradian Resources from 94 drill holes (31,344 m) which included some basic geotechnical parameters.

Snowden had noted the geotechnical data collected was not fully compliant with the industry norm but does not consider this to be significantly influential at the current level of study, although this issue should be addressed for higher level of studies. Snowden also provided recommendations to Dalradian Resources on methodologies to improve the quality of the data collected and a quality assurance process prior to data analysis.

Results of the scan line mapping and review of geotechnical data concluded that:

Rock mass in the exploration adit is fresh and unweathered at greater than 100 MPa Single dominant joint set with mean orientation of 86°/49° (dip/dip direction) Five faults have similar orientation of 52°/330° and another at 75°/35° Overburden is clayey at 3 to 18 m (typically 9 m) across the site. Weathered zone - 0 to 81 m. In most drill holes extends from 25 to 50 m. The average RQD is presented in Table 16.1 for the different lithologies. The percentage of the rock mass having joint sets of: o None - 0.4% o One - 45.9% o Two - 51.4% o Three and more - 2.3%

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6QRZGHQLQGLFDWHGWKDWLQ³DEURDGVHQVHWKHUHLVDJUHHPHQWDVPXFKRIWKHFRUHLVUHFRUGHG DVEHLQJLQWHUFHSWHGE\RQO\RQHMRLQWVHWDQGWKHMRLQWFRQGLWLRQVZHUHVLPLODU´EHWZHHQWKH observation made during mapping and the core.

Table 16.1 RQD and Joint Data by Lithology

R Q D % Jointing Description Min Avg. Max J/m Psammite PS 3 72 100 1.48 Psammite Weathered PSW 0 39 84 1.55 Semi Pelite SP 0 65 100 1.94 Semi Pelite Weathered SPW 0 33 100 3.48 Pelite PE 0 53 100 1.95 Pelite Weathered PEW 0 20 38 2.64 Source: Snowden, January 2012

Observations of the geotechnical conditions of Curraghinalt Project made during 0LFRQ¶V site visit in April, 2012 are consistent with observations made by Snowden in scan line mapping. The existing underground excavations appear stable with no major falls of ground or geotechnical stability issues, despite the excavations having been made in the late 19¶V

Based on the geotechnical information presented by Snowden, and for the purpose of the mine design for the PEA, the rock mass is considered to be of fair to good quality. A preliminary assessment of the geotechnical information provided by Snowden indicated ³«the overall rock mass is fair based on the RQD with a typical rock strength of R3 (25 to 50 MPa) though the weathered material is poor. The average joint set is 0.5 m but as there is QR EHGGLQJ DQG WKH IROLDWLRQ DSSHDUV ZHOGHG WKLV LQ HIIHFW TXLWH D PDVVLYH URFN PDVV´ (Snowden, January 2012).

Micon concurs with the recommendations made by Snowden regarding future geotechnical work on data collection, analysis and testing. This additional work will be required to better determine the rock mass classes for the deposit and for mine design purposes.

Micon recommends that, during the geotechnical data collection, the rock mass should be accurately characterized and described. This allows the rock mass attributes, parameters and ratings to be assigned at a later stage enabling the classification to be made in any of the proposed rock mass classification systems (e.g., RMR, Q-system or MRMR).

16.3 CUT-OFF GRADE DETERMINATION

Within the mineral resource, the selection of material for inclusion in the mine plan for this PEA mine plan was determined by a cut-off grade. The value of the mining and milling cut- off grade for underground mining at Curraghinalt Project is at 5.0 g/t gold (Au).

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This cut-off grade was determined considering the preliminary cost estimates for mining, milling and general administration for the deposit, taking into account expected recoveries and metal price.

16.4 MINERAL RESOURCES CONSIDERED IN THE MINE PLAN

For the purposes of the PEA, the diluted and potentially extractable portions of the mineral resources that are above a 5.0 g/t Au cut-off grade have EHHQ WHUPHG ³PLQHUDO UHVRXUFH FRQVLGHUHGLQWKHPLQHSODQ´7KLVWHUPLVXVHGhere for the purposes of distinguishing the mineral resources contained within the preliminary mine design and production plan from the mineral resource estimate tabulated in Section 14.

The mineral resource estimate given in Section 14 is derived from a geological block model created in Datamine software and was used as the basis for this PEA. The Datamine models were converted to Surpac format in order to prepare the PEA mine design and plan.

The mineral resource considered in the mine plan for Curraghinalt was based on the following estimation parameters:

Tonnages above the 5.0 g/t Au cut-off grade, after applying dilution factors to the grade to account for a minimum mining width of 1.8 m for stopes and 3 m for sills.

A non-recoverable crown pillar extending 20 m below and parallel to the topography.

Exclusion from the mine plan of resources below the -490 m level.

Mining recovery of 95%.

16.4.1 Mining Recovery and Dilution

The mining recovery and dilution values for Curraghinalt deposit were estimated based on the geological characteristics of the vein system, the use of the longhole retreat mining method and the dimensions of the selected narrow-vein mining equipment.

A mining recovery of 95% is assigned to account for mineralized material losses contributed by continuity of the deposit and mining activities in general.

The dilution values for the Curraghinalt Project were estimated in two stages, during the: Mineral resource stage and, Mine design stage.

At the mineral resource estimate stage, all the mineralized blocks in the block model with vein thickness of greater than 0.10 m, but less than 1.0 m, were diluted to 1.0 m. The diluting material is conservatively assumed to have zero grades. The mineral

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resource was then reported at an economic cut-off grade of 5.0 g/t Au over the minimum thickness of 1 m (Hennessey and Mukhopadhyay, January 2012).

In the mine design stage, all the mineralized blocks in the block model with thickness of greater than 1.0 m, but less than 1.8 m in thickness, were diluted to 1.8 m. The diluted material is considered as external dilution and is conservatively assumed to have zero grade. The 1.8 m width was considered as the minimum mining width for the stopes so, except where the veins have been shown to be wider than this, 1.8 m is the final mined stope width.

The design width of the sublevel sills is 3.0 m. This is the minimum mining width for rubberized equipment entry to develop the sill, perform production drilling and blasting and for blasted material to be extracted out of the stope. Again, additional dilution at zero grade was added to the mineralized zone to bring the width of the sills less than 3.0 m to 3.0 m. This diluted material is considered as external dilution.

The overall weighted average mining dilution is 33%. This dilution is additional to that already included in the resource estimate to account for a 1.0 m minimum width. A breakdown of the percentage dilution by resource classification is presented in Table 16.2.

Table 16.2 Percentage Dilution

Percentage Percent G rade Differences Remarks/ Description Dilution Au Ag Cu Reference Measured 15% -22% -22% -21% Percentage Difference Indicated 28% -45% -45% -45% from "Mineral Resources Considered in the Mine Total Measured & Indicated 27% -45% -44% -44% Plan´ to Mineral Total Inferred 35% -61% -61% -61% Resources Considered in the Mine Plan Diluted to Overall Total Tonnage Dilution 33% 1.8 m at 95% Recovery with External Dilution"

The mining recovery and dilution values are considered to be within the range for the proposed mining method and backfill incorporated, level interval, regularity and continuity of the deposits, losses in fines during blasting and mucking, and potential material left in the stope to form pillars to stabilize the excavations.

16.4.2 Classification of Mineral Resources Considered in the Mine Plan

Table 16.3 displays the conversion of the 2012 Mineral Resources to the mineral resource considered in the mine plan diluted to a 1.8 m mining width at 95% recovery with external dilution.

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The mineral resources considered in the mine plan for Curraghinalt Project are:

Measured and Indicated resources o 1.44 Mt o 8.88 g/t Au average grade o 2.95 g/t Ag average grade o 0.12% of Cu average grade Inferred resource o 7.98 Mt o 7.91 g/t Au average grade o 3.38 g/t Ag average grade o 0.08% of Cu average grade

These values include the losses in tonnages from the crown pillar, mineralized material at depth, minimum mining widths, recoveries and dilution factors described in the previous sections.

Table 16.3 Measured, Indicated and Inferred Mineral Resources Considered in the Mine Plan

Avg. Avg. Avg. Tonnage Description Au Ag Cu Remarks/Reference (Mt) (g/t) (g/t) (%) M ineral Resource Measured 0.02 21.51 17.56 0.490 Indicated 1.11 12.84 4.05 0.170 Hennessey and Mukhopadhyay, Total Measured & Indicated 1.13 12.99 4.29 0.176 January, 2012 Total Inferred 5.46 12.73 5.43 0.118

M ineral Resource In C rown Pillar and Below -490 m Elevation Measured 0.00 20.46 17.36 0.49 Resource with thickness 0.1 m Indicated 0.08 12.64 4.01 0.17 to 1.0 m diluted to 1.0 m as Total Measured & Indicated 0.08 12.80 4.27 0.17 reported in 2012 Mineral Total Inferred 0.24 12.74 5.44 0.12 Resource

M ineral Resource Considered in the M ine Plan, prior to application of modifying factors Measured 0.02 20.10 17.29 0.49 Indicated 1.03 12.69 4.01 0.17 Resource less Tonnages in Crown Pillar and below -490 m Total Measured & Indicated 1.05 12.84 4.27 0.17 Elevation Total Inferred 5.22 12.71 5.43 0.12

M ineral Resource Considered in the M ine Plan Measured 0.02 16.46 14.19 0.40 Mineral Resources Considered Indicated 1.42 8.76 2.77 0.12 in the Mine Plan Diluted to 1.8 Total Measured & Indicated 1.44 8.88 2.95 0.12 m at 95% Recovery with Total Inferred 7.98 7.91 3.38 0.08 External Dilution

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16.5 MINING METHOD

During the conceptual study prepared by Snowden (January 2012), seven underground mining methods and an open pit mining option were evaluated. The evaluated underground mining methods range from longhole stoping with and without backfill, cut and fill mining and shrinkage mining, incorporating of several backfill types.

The proposed underground mining method for the Curraghinalt Project for this PEA is Longitudinal Sublevel Retreat with a vertical level interval of 20 m from the floor of the top level to floor of the bottom level. This mining method and the height of the vertical level were selected based on:

Preliminary geotechnical assessment of the ground conditions indicate massive rock which should be conducive to better drilling, blasting and dilution control. Availability of mechanized narrow vein mining equipment for the geometry of the deposit. Dilution control measures and knowledge in mining and milling. Increased technology and knowledge in open stoping mining methods in narrow vein deposits. Overall continuity and regularity of the veins.

Paste backfill will be incorporated into the mining method as the primary backfill type. Waste rock will also be used as backfill material in combination with the paste backfill.

A sill pillar may be left between levels in the open stope, or a sill mat can be constructed to isolate the fill material from one level to another. Depending on the regularity and rock mass condition, more than one level can be mined within the same mining sequence.

Figure 16.1 presents typical schematic of a longitudinal sublevel retreat mining method with multiple mining levels.

Micon recommends that further optimization of the mining method be performed to evaluate potential additional economic benefits to the project by considering and combining a higher selectivity non-mechanized mining method with the proposed Longitudinal Sublevel Retreat.

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Figure 16.1 Typical Schematic of Longitudinal Sublevel Retreat with Multiple Mining Levels

16.6 UNDERGROUND MINE PLAN AND DEVELOPMENT

Access to the underground mineable zones will be through a decline excavated from the surface to elevation 210 m. Mining of the underground resources will commence at elevation 250 m, followed by the remaining mineralized zones as mining progresses.

Figures 16.2, 16.3 and 16.4 display isometric and plan views of the proposed underground mine for the Curraghinalt Project.

The main access decline will be excavated at a grade of -15% with an arched roof for greater stability at 4.5 m H x 4.5 m W.

Underground ramp networks, with similar dimensions to the main decline, connect one level to another within the ramping system. The vertical distance between the levels is 20 m from the floor the top sill to the floor of the bottom sill.

The proposed ramp system excavation includes mobile equipment passing bays, a safety bay and a sump between each level. Refuge stations or lunch rooms will be located in the cross- cuts at every 40 m interval (i.e., on every second level).

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Figure 16.2 Isometric View of Curraghinalt Project Mine Layout

Cross-cuts at 4.0 m H x 4.0 m W will be excavated to connect the ramp to the mineralized zones and are configured to be approximately perpendicular to the mineralized zones.

A sump of 3.0 m H x 4.0 m W x 3.0 L will be excavated in every cross-cut to decant all the water generated from the mining activities or potentially generated from the ground prior to pumping and discharging to the surface water treatment facility.

Two drifts at 10 m each in length will be excavated to connect the cross-cuts to the ore and waste-passes. These drifts will be utilized as temporary remuck bays for waste and mineralized storage areas until the ore and waste-passes are developed.

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Figure 16.3 Isometric View of the Mine Layout with Mineralized Veins

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Figure 16.4 Typical Plan View of the Mine Layout (at 180 m Elevation)

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Mining in a production level will commence from one end of the veins retreating to the OHYHO¶V FURVVFXW 7KH DYDLODELOLW\ RI QXPHURXV SURGXFWLRQ headings when a cross-cut intercepts the veins within the level provides multiple production headings to meet the 1,700 t/d production target.

The total amount of development in Life of Mine (LOM) for the Curraghinalt Project is presented in Table 16.4.

Table 16.4 Total Underground Development for the Curraghinalt Project

Description Total Length (m) Ramp 5,325 Ramp By Passes 157 Ramp Safety Bays 120 Sump in Ramp 120 Cross Cuts (incl. Refuge Stns & Service Drifts) 16,614 Sumps in Cross Cut 137 Sill 207,420 Sill (Through Waste) 7,152 Vent, Ore & Waste Pass 2,990 Total 240,035

16.7 MINE VENTILATION REQUIREMENT

The mine fresh air requirement estimation is based on the operating mining equipment fleet, their utilization and allowances for losses due to friction and short circuits. This forms the preliminary estimated fresh air requirement of approximately 118 m3/s or 250,000 CFM for the mine (Table 16.5).

The estimate is based on the provision of 0.06 m3/s per kW of diesel powered equipment operating underground with some allowance for personnel working around some equipment for dust and fume control.

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The ventilation demand estimated by Micon is preliminary in nature and considers all the diesel power mining equipment working within a deposit or mineralized zone. However, this may not be the case during operation because ore can be extracted from various zones.

Micon recommends that a detailed evaluation of the ventilation demand be investigated in further detail as part of a pre-feasibility study to ensure that the fresh air requirement for underground mine operation at Curraghinalt meets the Northern Ireland regulation for ventilation in mining operations, as well as any other regulations related to the levels of dust and particulates, carbon monoxide and oxides of nitrogen in the vicinity of operators and in the undiluted equipment exhaust which may be applicable.

Table 16.5 Estimated Ventilation Requirement

Power Power Flow Rate Description Units Total Utilization (kW/unit) (kW) (m3/s) Stoping Drill (Elec.-hyd.) 1 45 45 54% 24.3 1.5 Sill Narrow Vein Jumbo 5 38 190 54% 102.6 6.2 Development Jumbo (Double Boom) 1 110 110 54% 59.4 3.6 Bolter 1 110 110 54% 59.4 3.6 LHD at 4 m3 1 150 150 54% 81.0 4.9 LHD at 3.0 m3 4 210 840 54% 453.4 27.2 Trucks - 30 t 3 310 930 54% 502.0 30.1 Explosive Truck 1 78 78 45% 35.1 2.1 Scissor Truck 1 96 96 45% 43.2 2.6 Fuel and Lube Truck 1 96 96 45% 43.2 2.6 Mechanic Light Pickup 1 78 78 54% 42.1 2.5 Electrician Light Pickup 1 78 78 54% 42.1 2.5 Surveying Light Pickup 1 78 78 54% 42.1 2.5 Light Pickup (Service/Supervision) 1 78 78 45% 35.1 2.1 Man Carrier 1 96 96 22% 21.6 1.3 UG Grader 1 103 103 45% 46.3 2.8 Fume Dilution and Losses 20% 19.6 Total 117.6

16.8 DRILLING AND BLASTING

Underground ramps and cross cuts will be excavated using double boomed electric-hydraulic jumbo drills.

Sill development in the mineralized areas will be excavated using single-boomed electric- hydraulic jumbo drills to maintain the mining width of 3.0 m and to reduce the amount of dilution.

Production drilling will be performed by a combination of pneumatic and electric-hydraulic longhole drills. Production drill holes in the wider sections of the veins will be drilled with 89 to 102 mm (3.5 to ´ GLDPHWHU GULOO KROHV DQG GULOOLQJ ZLOO EH SHUIRUPHG ZLWK WKH

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electric-hydraulic longhole drill. Narrow vein production drill holes will be drilled by the VPDOOHUSQHXPDWLFORQJKROHGULOO )LJXUH ZLWKGULOOKROHVRIWRPP WR´ diameter.

Figure 16.5 Production Drilling in Narrow Vein (Source: Boart Longyear)

Ammonium Nitrate Fuel Oil (ANFO) will be the bulk explosive for all underground blasting and cartridge emulsion will be used during priming in production drill holes and in areas susceptible to water infiltration such as lifters in horizontal development.

16.9 H A U L A G E O F W AST E A ND M IN E R A L I Z E D M A T E RI A L

Blasted waste and mineralized material from the underground mining activities will be transported to the designated disposal areas or processing facility by rubber tyre mechanized mining equipment.

Waste material generated during the initial stage of the underground development will be trucked to the surface until voids are available in the open stopes. Blasted waste and mineralized material from the cross-cuts and stopes will be mucked by Load-Haul-Dump (LHD) and hauled to the remuck bays or ore and waste-passes located in the cross-cuts.

The blasted muck from development faces will be cleaned with 4 m3 (5.2 yd3) LHD units to temporary re-muck bays or directly into waste-passes. Narrower development in sills will be mucked with a low-profile 3 m3 (3.9 yd3) LHD units.

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Material from the ore and waste-passes will be loaded into 30 t low-profile diesel-powered trucks to be transported to the surface.

Oversize material will be blasted during the end of the shift in secured areas before being discharged into the ore or waste-passes or used as backfill material.

An estimated 1.26 Mt (653,000 m3) of waste material will be generated during the Life of Mine (LOM) based on the mine plan and it is assumed at that 50% of this material will be utilized as backfill material underground.

16.10 B A C K F I L L SYST E M - PAST E F I L L

The placement of tailings as backfill material at the Curraghinalt Project is an important component of the underground mining cycle. Paste backfill is an engineered material, which enhances the stability of excavated underground mine openings by providing engineered supporting pillars or as a working platform.

The tailings produced during gold extraction will be used as one of the major ingredients in the backfill material. Waste rock generated from underground development will be also be used as backfill material to minimized the amount of material transportation to the surface disposal areas.

Paste backfill will be prepared with dewatered tailings from the mill, mixed with cement and a binding agent and tailings slurry or water to control the pulp density.

The effective void volume for backfill placement is estimated based on the assumption that 90% or 3.0 M m3 of the voids in the stopes or mineralized areas are fillable. A total amount of 2.7 M m3 will be filled with paste backfill and the remaining 0.33 M m3 will be filled with the waste material generated underground.

Micon recommends that further studies be performed on the backfill system for the Project and that a systematic laboratory test program be performed on the mine tailings to determine the optimum cement addition and cement type to be used as the binding agent for the paste backfill.

16.11 MINE PRE-PRODUCTION AND PRODUCTION SCHEDULE

The overall mine production schedule is shown in Table 16.6. Figure 16.6 Figure 16.7 and Figure 16.8 display the underground production tonnage ramp up and annual gold and silver grades, respectively.

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Table 16.6 Curraghinalt Project Annual Production Plan

L O M Total -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Description Tonnes/ G rades Sublevel 7,534,046 236,459 386,791 525,524 482,681 454,786 531,030 453,227 540,662 477,534 492,158 498,164 513,510 487,576 508,063 486,633 459,249 Au (g/t) 8.56 8.67 8.93 8.77 8.63 8.63 8.83 8.76 8.33 8.36 8.54 8.58 8.68 8.50 8.30 8.47 8.15 Ag (g/t) 3.52 3.76 3.82 3.65 3.51 3.48 3.59 3.58 3.38 3.40 3.41 3.44 3.44 3.50 3.51 3.65 3.43 Cu (%) 0.09 % 0.08 % 0.09 % 0.09 % 0.09 % 0.09 % 0.09 % 0.09 % 0.08 % 0.08 % 0.09 % 0.09 % 0.09 % 0.09 % 0.09 % 0.09 % 0.08 % Sill 1,890,460 73,791 78,584 94,976 137,819 165,714 89,470 167,273 79,838 142,966 128,342 122,336 106,990 132,924 112,437 133,867 123,132 Au (g/t) 6.03 6.10 6.70 6.07 5.98 6.02 6.33 6.07 5.76 5.89 5.94 6.08 6.05 6.06 5.92 5.95 5.84 Ag (g/t) 2.49 2.75 2.81 2.68 2.45 2.45 2.58 2.50 2.32 2.32 2.38 2.43 2.37 2.46 2.49 2.62 2.45 Cu (%) 0.06 % 0.06 % 0.07 % 0.07 % 0.07 % 0.07 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % 0.06 % Total Tonnage 9,424,506 310,250 465,375 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 620,500 582,381 Au (g/t) 8.06 8.11 8.55 8.36 8.04 7.93 8.47 8.03 8.00 7.79 8.01 8.09 8.23 7.98 7.87 7.92 7.66 Ag (g/t) 3.31 3.52 3.65 3.50 3.28 3.20 3.45 3.29 3.24 3.15 3.19 3.24 3.26 3.28 3.33 3.42 3.22 Cu (%) 0.08 % 0.07 % 0.09 % 0.09 % 0.09 % 0.09 % 0.09 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % 0.08 % Contained Au (t) 75.94 2.52 3.98 5.19 4.99 4.92 5.26 4.98 4.96 4.83 4.97 5.02 5.11 4.95 4.88 4.92 4.46 Contained Ag (t) 31.24 1.09 1.70 2.17 2.03 1.99 2.14 2.04 2.01 1.96 1.98 2.01 2.02 2.03 2.07 2.12 1.88 Contained Cu (t) 7,784.01 229.60 409.06 563.95 541.76 531.70 539.99 496.73 492.81 471.28 508.10 511.55 510.72 516.98 502.51 495.30 461.95

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Figure 16.6 Annual Production Tonnage

600,000

500,000

) y / t (

400,000 e g a n n

o 300,000 T

l a u n

n 200,000 A

100,000

0 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Tonnage from Sublevel 236,4 386,7 525,5 482,6 454,7 531,0 453,2 540,6 477,5 492,1 498,1 513,5 487,5 508,0 486,6 459,2 Tonnage from Sill 73,79 78,58 94,97 137,8 165,7 89,47 167,2 79,83 142,9 128,3 122,3 106,9 132,9 112,4 133,8 123,1

Figure 16.7 Annual Gold Grades

)

t 10.0 / g ( 9.0 u A

, 8.0

s e

d 7.0 a r 6.0 G

n 5.0 i n i 4.0 M

l 3.0 o G

2.0 e g

a 1.0 r e

v 0.0 A -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Avg Au Grd 8.11 8.55 8.36 8.04 7.93 8.47 8.03 8.00 7.79 8.01 8.09 8.23 7.98 7.87 7.92 7.66 Sublevel Au Grd 8.67 8.93 8.77 8.63 8.63 8.83 8.76 8.33 8.36 8.54 8.58 8.68 8.50 8.30 8.47 8.15 Sill Au Grd 6.10 6.70 6.07 5.98 6.02 6.33 6.07 5.76 5.89 5.94 6.08 6.05 6.06 5.92 5.95 5.84

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Figure 16.8 Annual Silver and Copper Grades

3.7 g/t 0.10 % 3.6 g/t 0.09 % 0.08 % 3.5 g/t 0.07 % 3.4 g/t 0.06 % 3.3 g/t 0.05 % 3.2 g/t 0.04 % 0.03 % 3.1 g/t 0.02 % Average Ag Average Ag Grades (g/t) 3.0 g/t 0.01 % Grades Cu (%) Average 2.9 g/t 0.00 % -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Avg Ag Grd 3.52 3.65 3.50 3.28 3.20 3.45 3.29 3.24 3.15 3.19 3.24 3.26 3.28 3.33 3.42 3.22 Avg Cu Grd 0.07 0.09 0.09 0.09 0.09 0.09 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08 0.08

16.12 DILUTION CONTROL

Dilution control during the mining operation will be a collaborative responsibility of all the departments on the Curraghinalt Project site.

The following highlights some of the measures which will have to be implemented during operation to reduce the amount of potential dilution during mining:

Sampling and mapping of the mineralized structures during the development of the sills by geological department. Interpretation of mineralized structures when the sills are developed by the geological department. Accuracy of the mine engineering group in the drilling and blasting design from the interpreted mineralized structures. Accuracy of the surveying personal to identify the set-up location for production drill rings. Drilling accuracy and deviation control by the drillers. Drill holes verifications - the drill holes drilled from the upper sill will allow the breakthrough location of each hole to be examined. This allows for re-drilling of any deviated drill holes before blasting to minimize unplanned dilution. Drill cutting sampling at the prescribed interval recommended by the geological department during production drill. Rapid turnaround time on sample preparation and assay reporting by the analytical team.

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Revision of the geological interpretation with the assay results from the drilling cutting and adjustment made according to the production drilling and blasting plan. Training for all the production drillers and parties involved.

Additional costs have been accounted for in the mine plan for auxiliary manpower to assist with the sampling of material underground.

Other dilution control measures can include the installation of an optical sensor-based sorting machine to separate the waste material from mineralized material prior to mineral processing.

16.13 EQUIPMENT FLEET

16.13.1 Mine Services

The underground services equipment estimate is presented in Table 16.7.

Table 16.7 Estimated Underground Service Equipment

Description Units Jackleg/Stoper 4 Compressor (2500 CFM) 1 Surveying Equipment & Software 1 Ventilation Fans - 75 HP 2 Ventilation Fans - 50 HP 2 Ventilation Fans - 30 HP 4 Pumps - 30 HP 2 Pumps - 15 HP 4 General Electrical Equipment 1 Fuel Tank (15,000 litre) & Facilities 1 Self-Contained Rescue Chamber 1

16.13.2 Underground Mobile Equipment

The basis of estimate for the mobile equipment is based on the assumption of the following productivity estimated for the proposed equipment, as shown in Table 16.8

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Table 16.8 Estimated Underground Service Equipment

Description Units Productivity Jumbo productivity in the ramp m/y 1,157 Jumbo productivity in the cross-cuts m/y 1,822 Production drill productivity t/y 110,974 Narrow vein jumbo productivity m/y 3,053 Bolter m/y 1,313 LHD at 4 m3 (ramp) t/y 90,520 LHD at 4 m3 (cross-cuts) t/y 204,400 LHD at 3.0 m3 (stope) t/y 190,530 Trucks - 30 t units Varies with depth

Table 16.9 summarizes the proposed equipment fleet required to develop and extract 1,700 t/d of mineralized material from the Curraghinalt deposit. The table presents the average number of equipment units over the life of mine. The underground mobile and auxiliary equipment estimate is EDVHGRQSURMHFW¶VGHPDQG at full capacity.

Table 16.9 Underground Mobile Equipment Fleet

16.14 MINING MANPOWER

The mine will operate at 365 days per year at 3 shift/day for 8 hours/shift with most of the major underground mining tasks be performed during the working days within a calendar week. Overtime expenses have been accounted for in the manpower estimate for additional work which has to be performed during the weekend periods.

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The workforce for the project will comprise locally and regionally trained and skilled manpower. Individuals will be trained and certified for the operation of a range of mining equipment to increase operational efficiencies.

All the mining and engineering staff will be encouraged to domicile within the surrounding communities, close to the project site.

The work schedule for geology and engineering staff will be 8 hours/day, 5 days/week. This manpower estimate depicts manpower on site and does not account for contractors, medical and general absenteeism, and vacations. The site manpower was estimated based on the project tasks and responsibilities, working hours, shift rotations, site operating equipment and operating capacity. A summary of the underground manpower requirements is presented in Table 16.10.

Table 16.10 Estimate M ine Labour Requirement

,QDGGLWLRQWRWKHDERYHWKHSURMHFW¶VJHQHUDODGPLQLVWUDWLRQDQGWHFKQLFDOPDQDJHPHQWZLOO require staff as outlined in Table 16.11.

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Table 16.11 General and Administrative Manpower

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17.0 R E C O V E R Y M E T H O DS

The gold at Curraghinalt is associated with chalcopyrite and pyrite. Generally, the gold is free milling and not refractory. The processing options considered for Curraghinalt take into account the association of gold with copper and pyrite.

The four process flowsheet options considered and economically evaluated for this PEA study are:

These options are shown schematically in Figures 17.1, 17.2, 17.3, and 17.4.

Figure 17.1 Option A Whole Ore Leach

FEED Cyanidation Tails

Refinery

Gold Bars

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Figure 17.2 Option B Bulk Flotation Concentrate Leach Whole Ore Leach

Feed Gravity Bulk Sulphide Tails Concentrator Flotation

Concentrate Cyanidation Tails

Refinery

Gold Bars

Figure 17.3 Option C Bulk Flotation Concentrate - No Leach

Feed Gravity Bulk Sulphide Tails Concentration Flotation

Concentrate Concentrate

Figure 17.4 Option D Copper Flotation and Pyrite Flotation Concentrate Leach

Copper Flotation Pyrite Flotation Float Tails Feed Circuit Tails Copper Flotation

Copper 1st Cleaner Tails Cyanidation Leach Tails

Concentrate

Refinery

Gold Bars

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Options A, and D both involve a similar set of processing steps including Crushing, Grinding, and Cyanidation. Option D includes 2 flotation steps producing a copper concentrate for sale and a pyrite concentrate for on-site cyanide leaching.

The source of assay data for this report was on-going testwork being conducted at G&T 0HWDOOXUJLFDO6HUYLFHVDQGUHFHQW/DNHILHOGWHVWZRUNFRPSOHWHGLQ ³$QLQYHVWLJDWLRQ into the recovery of gold and silver from the Tyrone Project prepared for Dalradian Resources Inc. Project 13471-)LQDO5HSRUW$SULO´SUHSDUHGE\6*6/Dkefield.).

17.1 DESIGN BASIS

A summary of the project production and process design criteria are presented in Table 17.1 and Table 17.2 respectively. The process design criteria are based on the metallurgical testwork discussed in section 13.0.

Table 17.1 Production Summary for the Dalradian Resources Tyrone Project Milling Plant

Item Option A Option D Whole O re Copper Flotation Leach and Pyrite Flotation Concentrate Leach Processed Annually, Average (t/y) 443,179 443,179 Feed G rades (L O M)

Operating Days Per Year 365 365 Mill Feed Rate, t/d 1,700 1,700 Chemical & Metallurgical Parameters

Cu Concentrate Grade, Cu % - 8.4

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Table 17.2 Process Design Criteria

OPT I O N A OPT I O N D

Copper Flotation and Whole O re General O re Characteristics Units Pyrite Flotation Leach Concentrate Leach Annual Processing Rate (nominal) t/y 621,000 621,000 Operating days/year day/year 365 365 Daily Processing Rate t/day 1,700 1,700 Average Ore Grades g/t Au 8.06 8.06 g/t Ag 3.31 3.31 % Cu 0.08 0.08 Ore Specific Gravity 2.85 2.85

C rushing

Crusher Rate - Nominal t/h 71 71 Crusher Operating Time h/day 14.4 14.4 Crusher Availability % 80 80 Crushing Rate - Design t/h 148 148 Crusher Type Jaw Jaw

Crushed Ore Product P80 mm 150 150

G rinding

Total Feed Tonnage to Grinding - Nominal t/day 1,700 1,700 Mill Availability % 92 92 Total Feed Tonnage to Grinding - Design t/day 1,848 1,848 Feed Rate to Grinding - Design t/h 77 77

SAG Mill Feed F80 mm 150 150 SAG Mill Discharge P80 micron 850 850 Abrasion Index kWh/t 0.1278 0.1278 Bond Work Index kWh/t 15.2 15.2 SAG Mill Unit Power Consumption kWh/t 11.6 11.6 SAG Mill Installed Power kW 896 896 Ball Mill Unit Power Consumption kWh/t 8.6 8.6 Ball Mill Installed Power kW 672 672

Ball Mill Discharge P80 micron 141 141

Copper Flotation

Feed Rate to Cu Flotation - Design t/h - 77 Cu Flotation Feed Density % solids - 40 Lime to Cu Flotation kg/t - 2.890 3418A Collector to Cu Flotation kg/t - 0.010 MIBC Frother to Cu Flotation kg/t - 0.047 Cu Concentrate Mass Pull % - 0.9 Cu Concentrate Production - Annual t/year - 5,585 Cu Concentrate Production - Design t/day - 17 Cu Concentrate Grade %Cu - 8 Cu Concentrate Recovery %Cu - 94 Cu Tailings - Design t/day - 1,785

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OPT I O N A OPT I O N D

Copper Flotation and Whole O re General O re Characteristics Units Pyrite Flotation Leach Concentrate Leach Concentrate Dewatering

Cu Concentrate Solids Feed Rate t/h - 0.7 Cu Concentrate Thickener % Solids % - 70 Cu Concentrate Filter Press % Solids % - 90 Flocculant Addition kg/t - 0.025

Bulk Sulphide Flotation Feed rate to Bulk Sulphide Flotation- t/day - 1,785 Design Mass Pull to Flotation Product % of feed - 6 Bulk Sulphide Concentrate - Design t/day - 111 Bulk sulphide production annual t/year - 40,565 Bulk sulphide tailings annual t/year - 562,173 3418A Collector Addition kg/t - 0.060

Cyanide Leaching and C IP

Feed rate to Cyanide Leaching - Nominal t/h 71 4.6 Cyanide Leaching Availability % 90 90 Feed rate to Cyanide Leaching - Design t/h 79 5.1 Leach Slurry Density % solids 40 40 Leach Slurry Flow rate m3/h 144 9.3 Leach Residence Time h 24 24 Number of Leach Tanks # 6 6 Cyanide Concentration g/L 2 2 Cyanide Consumption kg/t 0.6 0.6 Lime Consumption kg/t 1.84 0.6 Lead Nitrate Consumption kg/t 0.5 0.5

Effluent T reatment

Feed Rate to Cyanide Destruction - t/h 71 4.6 Nominal Effluent Treatment Availability % 90 90 Feed Rate to Cyanide Destruction - Design t/h 79 5.1 Slurry Density % solids 40 40 Slurry Flow Rate m3/h 79 9.3 Cyanide Destruction Residence Time h 2 2 No. Cyanide Destruction Reactor Tanks 3 3

Tailings Thickener % Solids % 60 60

17.2 PROCESS DESCRIPTION

17.2.1 Crushing (Options A and D):

The ore is delivered by mine trucks to the plant and fed over a grizzly with 400 mm openings to the Run Of Mine (ROM) ore feed hopper. The ore is extracted from the feed hopper by an

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apron feeder and fed to a jaw crusher. The crushed ore is discharged onto a conveyor which will transport the ore to the plant. The crusher product is 80% passing 150 mm.

17.2.2 Grinding (Options A and D):

The crushed ore reports to the grinding circuit. The primary grinding SAG mill discharge product typically has a size passing of 850 µm.

The secondary grinding ball mill will operate in closed circuit with a cyclone, with the target cyclone overflow 80% size passing 141 µm.

17.2.3 Copper Flotation (Option D):

The ball mill ore slurry discharge, at 50% solids, is fed to a conditioning tank, which then overflows to the copper flotation circuit. The conditioning tank also acts to buffer flotation from upsets in the grinding areas.

The copper concentrate is produced in a conventional copper flotation circuit. The copper FLUFXLW¶VILUVWFOHDQHUWDLOLQJVUHSRUWWRWKHFyanide leach circuit. Tails from the circuit are pumped to the bulk flotation circuit.

Aerophine Collector 3418A, MIBC frother and lime (used as a pH modifier) are added to the copper flotation circuit.

The concentrate from the rougher cells is fed to the cleaner cells from which the final copper concentrate is sent to a pressure filter for dewatering.

17.2.4 Copper Concentrate Dewatering (Option D)

Copper concentrate is thickened to 70% solids in conventional thickener. Copper concentrate thickener overflow is recycled to the water reclaim tank. This reclaim water tank can serve as a headtank for the reclaim water pumps provide the bulk of process water to the mill. It also serves as the water supply for the emergency diesel powered fire pump. Copper concentrates are filtered using pressure filters. This product is then sold to a copper smelter.

17.2.5 Bulk Sulphide Flotation (Option D)

For Option D, tails from copper flotation are subjected to a bulk sulphide flotation to recover all remaining sulphides, pyrite, etc. as well as gold and silver bearing minerals.

The flotation concentrate is thickened and pumped to cyanidation circuit for recovery of gold, and silver.

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17.2.6 Pre-treatment and Cyanidation (Options A and D)

A six stage CIL circuit is installed based on a retention time of 24 hours and a new carbon concentration of 20g/L. Slurry pH is adjusted to 10.5 with the addition of lime. The first leach tank is aerated, and lead nitrate is added. The oxidation of sulphides in this step reduces cyanide consumption considerably as well as enhances leaching kinetics. The leach circuit feed is pumped to the train of six CIL tanks in series where it is contacted with activated carbon. Each of the tanks is provided with an in-tank screen to allow the slurry to flow by gravity through the CIL train, while retaining the carbon in the tanks. The carbon is transported counter-currently to the slurry flow by means of carbon advance pumps. The loaded carbon from the first CIL tank is collected on a screen and is sent to the carbon stripping circuit. The slurry leaving the last CIL tank is passed over a safety screen to capture any carbon particles before it is sent to the cyanide destruction circuit.

17.2.7 Carbon Stripping/Carbon Reactivation/Gold Room (Options A and D)

A carbon stripping/carbon reactivation circuit is installed to treat the loaded carbon to recover the gold and to recycle the carbon to the CIL circuit.

The loaded carbon is acid washed and then it is fed to a ZADRA elution circuit to strip the gold from the carbon. The pregnant solution is then sent to two electrowinning cells to recover the gold from solution. The carbon from the stripping circuit is fed to a diesel-fired rotary kiln to reactivate the carbon which is passed over a sizing screen prior to being recycled to the CIL circuit.

The stainless steel wool cathodes carrying the gold from the electrowinning cells are fed to an induction furnace to produce gold doré bars for sale.

17.2.8 Cyanide Destruction (Options A and D)

A cyanide destruction circuit is installed to treat the slurry discharge from the CIL circuit.

17.3 MANPOWER

Manpower requirements are based on a continuous operation, 7 days per week and 365 days per year.

Table 17.3 shows the manpower requirements for the base case (Option A) and Option D.

Option D includes the process operations of Option A, but also includes two additional flotation circuits. Thus, Option D will require nine more personnel than Option A.

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Table 17.3 Summary of Estimated Processing Plant Operations Personnel

Option A: Option D: Copper Float Whole O re Leach and Pyrite Flotation Description Concentrate Leach Plant Superintendent 1 1 Senior Metallurgist 1 1 Metallurgist 0 1 Chief Chemist 1 1 Sample Prep. Technician 4 4 Assay Technician 2 2 Plant Shift Supervisor 4 4 Crusher Operator 4 4 Plant Operator 4 4 Control Room Operator 0 4 Operator Helper 4 4 Tailings Operator 2 2 Load out Operator 0 4 Cyanidation Plant Operator 4 4 Day Labour Crew 4 4 Maintenance Superintendent 1 1 Maintenance Supervisor 1 1 Instrument Technician 2 2 Electrician 4 4 Plant Mechanic 4 4 Security Personnel 5 5 Senior Accountant 1 1 Accountants/buyer 1 1 HR and Community Relations 1 1 Environmental Technician 1 1 Supervisor 1 1 Bus Driver 1 1 Secretaries/Admin Assistant 2 2 Site Surface Maintenance Crew 4 4 Equipment Operator 2 2 Site Administrator 1 1 Total 67 76

17.4 PROCESSING CONCLUSIONS

The capital and operating costs for the concentrator have been estimated on the basis of available data. The PEA considers four possible process flowsheets for the extraction of gold from the Curraghinalt resource. Of these, the whole-ore leach (Option A) was considered the most effective and has been used as the base case for project evaluation. It is recommended that this option should be explored further in future testwork, including optimizing grind, reagent strengths and retention times, and pilot studies are required before flowsheet development and design specifications can be finalized

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Micon considers the relatively low grade of the copper concentrate produced in Option D will present a challenge with regard to shipping and marketing. Dalradian Resources will need to confirm off-take terms with a smelter, and also investigate any deleterious content of the copper concentrate.

The PEA has been prepared using the metallurgical testwork results available at the time. Testwork is ongoing that may validate producing a more valuable, higher grade, copper concentrate, based on a lower weight recovery than was considered for flowsheet Option D during the PEA evaluation. Accordingly, Micon is of the opinion that during further development of the project this option should continue be considered along with the base case flowsheet A. It is of critical importance that representative samples of the copper mineralization are used for future testwork

Micon also suggests that alternative flotation reagents should be investigated in future copper flotation testwork, since availability of the collector used in the G&T copper flotation testwork is subject to manufacturing capacity constraints.

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18.0 PR OJE C T IN FR AST RU C T UR E

18.1 SITE ACCESS ROAD

The property is readily accessible by a network of existing all-weather paved highways and local roads. These include more specifically the two lane paved road B48 from Omagh to the village of Gortin and from there the B46 road to Greencastle.

The existing adit is accessed from a narrow road on the north side of the Curraghinalt deposit. Nevertheless, Micon considers that ultimately the property should be accessed from the south (off the B46) which most notably will have the benefit of reducing the potential for congestion associated with the northern access.

18.2 SITE ROADS

A site road system will be built to connect the onsite facilities: gate-house, administration complex, process plant, mine dry, warehouse, ore and waste storage pads, tailings storage facility, etc.

18.3 POWER SUPPLY

The base case total power requirement is 70 kWh/t which results in a maximum demand of approximately 6 MW for a 1,700 t/d operation. Power will be supplied by connecting to the national high voltage electricity grid.

Due to the required demand, a predominantly high capacity 33kV overhead transmission line will be built from the 110/33kV substation located at Strabane, for approximately 27 km to the site. It is expected that it should follow the existing right of way (ROW) to Plumbridge and then parallel the existing 11kV line to Gortin and from there directly to the site. There is another 110/kV substation located at Omagh which is equidistant but, according to Northern Ireland Electricity Ltd. (NIE), does not have enough capacity. In addition, NIE has also indicated that it is planning to install a new 33kV switchboard at the Strabane site within the next 18 months. Another option could be to tie-in to the existing 110kV transmission line located between Strabane and Omagh and building a 33kV substation at the site itself. It would take Northern Island Electricity Ltd. approximately two years to complete the power line construction, once it receives a firm commitment from Dalradian.

A substation will be built onsite which will transform the electrical power from 33kV to 11kV for distribution around the site.

In case of interruption of the main power supply, a 1.12 MW, 400 V standby diesel generator located on site will automatically be activated through the use of a motor control centre (MCC) in order to supply emergency power.

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18.4 WATER SUPPLY

Micon considers that options for make-up water for the process plant are fairly limited. Effectively, although there is a river (Owenkillew River) which is located near the edge of the property on the North side, the fact that it is a protected habitat with Areas of Special Scientific Interest (ASSI) makes it usage unlikely or, if not, would then require a long lead time and substantial efforts in order to obtain the necessary permits from the governmental agencies.

A hydrogeology survey report produced by SLR indicates that the usage of water wells on the property is not a likely option either. The report states that according to the Geological Survey of Northern Ireland (GSNI) the meta-sediments of the Mullaghcarn, Glenawne and Glanelly Formations are all considered to have limited potential for containing significant bedrock aquifer. Most supply wells in these meta-sediments are known to have only a low yield and the occurrence of moderate yields are unusual. However, it is important to mention that SLR recommends undertaking a local wells survey around the adit area.

Micon considers that, at this point in time, the most likely viable option will be to supply the make-up water for the initial process plant start-up with the construction of a rain water catchment system. The SLR report states that the rainfall in the area is in the order of 1,300 mm/y to 1,400 mm/y with a considerable net residual residue for run-off water in the order of probably 1,000 mm/y. Micon proposes utilizing the proposed 25 ha tailings storage facility (TSF) to capture the rain water and possibly store pumped run-off water if necessary.

In Micon's opinion, further work is required in order to properly assess water supply sources and recommends that further investigations be carried out during further development of the project.

18.5 BUILDINGS

As previously referred to, the site will accommodate the construction of a process plant, a two storey administration complex which will also house the technical department offering a total available space of 1,060 m , laboratory, a single-storey mine dry (change-house) complex incorporating mine supervisors offices covering an area of 816 m , a maintenance shop for underground and surface equipment with an area of 896 m , a 756 m heated warehouse and a gatehouse.

18.6 ANCILLARY FACILITIES

Water tanks for industrial use will be built including one for fire suppression water. The fire suppression water will be distributed to the protected area through an underground water pipes network.

A fuel tank farm with adjacent fuelling station will primarily supply underground equipment needs.

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Sewage will be collected from the various facilities and pumped into a sewage treatment plant.

Communication facilities will be comprised of a redundant fibre communication backbone system which will link and manage the data transmission of the distributed control system (DCS), third party programmable logic controllers (PLCs), motor controls, fire detection system, and computers around the mine site. It is expected that the site will likely be connected to the local phone system network. In the case that this would not be possible, then a Voice over Internet Protocol (VoIP) telephone system would be required. In any case, a transceiver or cellular radio tower will be installed in order to optimize the utilization of cellular telephones, as the current coverage in the area is not adequate.

For the duration of the construction work, portable temporary office and facility trailers will be used.

18.7 CONSTRUCTION FILL MATERIALS

In order to reduce the capital costs and land disturbance, Micon recommends utilizing the non-PAG waste rock which will be produced by further underground exploration and mine development activities as bulk fill materials to meet the construction needs, where appropriate. The structural fill material will be imported from local suppliers as required.

18.8 TAILINGS STORAGE FACILITY

In 2011, Golder Associates UK (Golder) was retained by Dalradian to carry out a Tailings Management Facility Site Selection Study aimed at identifying potential sites to store approximately 2.92 million tonnes of tailings. Golder, in its December 2011 report, identified seven potential sites and recommended three of them (in order of preference, Sites 3, 7 and 6) for further investigation. Golder also recommended the use of conventional tailings disposal technology and to carry out the construction work in two separate phases (Phase I and Phase II).

18.9 SCHEMATIC SITE LAYOUT

At this stage of project development, there is a high degree of flexibility in the siting of the main ramp portal, ventilation shaft collars, process plant and other surface infrastructure. Micon has not, therefore, made specific recommendations as to location of these works pending discussion with affected landowners. Accordingly, the layout shown in Figure 18.1 is schematic only, in the expectation that negotiation with landowners will take place as the project moves forward so that a site-specific design can be developed during the next stages of project engineering.

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Figure 18.1 Schematic Surface Layout

Tailings Storage Facility (250,000 m2) 500 m

Reclaim Water Pond

Reclaim Water line

Waste Rock Stockpile Process Water Pond

Crushed Ore Main Decline Crusher ROM Stockpile/ Crusher Pad Stockpile Conveyor (Portal)

Powerline Sub- Reclaim/ C Tailings Process Changehouse station Recycling Area (Mine Dry) V line Water line

Process Plant

Maintenance Warehouse Workshop

Construction Laydown Area

Fuel Process Plant Storage

Admin Access Bldg

Parking Area

0 25 50 75 100 metres

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19.0 M A R K E T ST UDI ES AND C O N T R A C TS

In evaluating the project, Micon made assumptions about the terms on which products from the project might be sold which, based on its experience on similar projects elsewhere, it believes to be reasonable. Micon is not aware of any project-specific contract or off-take terms having been negotiated for sales from the Curraghinalt project.

Significant contracts that will be required prior to entering production will be for:

the supply of electrical power to the site. the supply of fuels, explosives, cement, lime, sodium cyanide and other reagents and consumables for its mining and processing activities. Micon has used publicly available information to derive its estimate of the cost of power required by the project from the relevant utilities, and has used knowledge and experience gained on other projects to derive other significant input unit costs. Micon is not aware of any project-specific contract or terms having been negotiated for supplies to the Curraghinalt project.

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20.0 E N V IR O N M E N T A L ST UDI ES, PE R M I T T IN G A ND SO C I A L O R C O M M UNI T Y I MPA C T

20.1 ENVIRONMENTAL STUDIES AND ISSUES

20.1.1 Introduction

Micon has reviewed available environmental studies completed by SLR Consulting (SLR) and Golder. This section provides a summary of this work as well as an indication of the permitting and management plans needed to take the project into production. Micon completed a site visit, but the work does not constitute a liability assessment of the property. Similarly, existing permit information has been provided by Dalradian Resources but this does not constitute a legal opinion on the status of existing permits.

20.1.1.1 Landscape and Ecology

This area of the Sperrin Mountains is designated an Area of Outstanding Natural Beauty. There are also several protected and special interest areas around the project (Figure 20.1).

Figure 20.1 Ecological and Landscape Designations

Source: SLR, August, 2012

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The nearest Areas of Special Scientific Interest (ASSIs) to the project include Mullaghcarn/ Mountfield Quarry, Murrins, Cashel Rock, Boorin Wood and Black Bog. The nearest Special Areas of Conservation (SACs) include Drumlea and Mullan Woods, Owenkillew River, and Black Bog.

Within the Owenkillew River SAC are four Annex II listed species: Freshwater Pearl Mussel (Margaritifera margaritifera), River Otter (Lutra lutra), Brook Lamprey (Lampetra planeri) and Atlantic Salmon (Salmo salar).

20.1.1.2 Environmental Baseline Studies

Environmental baseline studies were initiated by SLR for Dalradian Resources in June 2010 and have included collecting data on meteorology, hydrology, hydrogeology, water quality, sediment quality, acid rock generation potential of the mineral and waste rock, flora, terrestrial and aquatic fauna, air quality, visual resources, cultural heritage resources, and the local socio-economy. Initial monitoring locations for physical and biological components are shown in Figure 20.2.

Figure 20.2 Environmental Monitoring Locations

Source: SLR, August, 2012

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20.1.1.3 Water Quality

Surface water quality is monitored at fifteen sites. Water quality is good for aquatic life; there are naturally occurring elevated concentrations of copper, iron, zinc, as well as of ammonia, nitrite, and phosphorous, likely associated with runoff from local farming activity. The marginally elevated concentrations of heavy metals in surface water are likely a result of mineralization in this region. The background copper concentration detected to date in surface water bodies near the adit range from 2.3 micrograms per litre in the Owenkillew River, to 12.8 micrograms per litre in nearby tributaries of the Owenkillew River; the maximum background copper concentration in drainage from the existing adit is 9.0 micrograms per litre.

20.1.1.4 Soil Quality

Soils in the project area range from predominantly blanket peats on upper slopes to humic gleys on lower slopes in the valleys and podzols on the valley floors (SLR 2011a). Heathland grasses and shrubs are the dominant vegetation at the upper elevations; the lower elevations within the project area are mainly grasslands used for farming with taller shrubs and forested areas bordering farms and along the river valley floor (Figure 20.3).

Figure 20.3 Landscape of the Project Area

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20.2 WASTE AND WATER MANAGEMENT

20.2.1 Waste Characterization and Waste Management

Based on the ABA laboratory results, it was concluded that 5 samples were considered to be acid generating, 23 were classified as non-acid generating and the results of 7 samples were considered to be inconclusive. Rock types that were potentially acid generating included the following: oxidized semi-pelites with quartz veins and chlorite or pyrite; semipelite/psammite with chlorite stubs >2mm; and psammite with predominant quartz vein (SLR 2012b).

Leach tests indicated that the potential contaminants of concern are aluminium, arsenic, iron, manganese, copper, iron, molybdenum, nickel and lead. The rock type that was most problematic was the semi-pelite/pelite with potassium feldspar and hematite veinlets under pH3 conditions (SLR 2012b). Additional testwork on waste rock and tailings conducted during subsequent stages will provide additional information to assist with segregation and management of the various rock types during operations.

An estimated 40% of tailings will be backfilled underground. The remaining tailings will be disposed of on surface in a conventional tailings impoundment, which will be lined with synthetic and/or impermeable material. It is recommended that the project consider paste or dry stack tailings to minimize the footprint and increase the available options for siting the tailings and progressive reclamation to minimize visual impacts.

Potentially acid generating waste rock will be backfilled underground and ultimately flooded after closure which will minimize future oxidation and metal release. The remaining waste rock will be stored in surface dumps, which will be lined if required, contoured and eventually capped and revegetated in character with the surrounding landscape.

20.2.2 Water Management

Temporary abstraction permits granted by the Northern Ireland Environmental Agency (NIEA) were in place for the exploration drilling program to provide for drilling makeup water. There is currently an approved discharge from activities associated with the exploration adit.

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There are not any groundwater abstraction licences in the immediate project area. A discharge consent has been obtained from the NIEA for site drainage from the existing exploration adit during any future underground exploration works.

Mine water will be collected in underground sumps and pumped to surface for use as makeup water in the process plant. This water may be treated prior to use in the process, depending on the concentrations of suspended sediments and blasting residues.

The process plant is designed to include an INCO SO2/air cyanide destruction circuit to meet the limit of 10 ppm weak acid dissociable (WAD) cyanide at the point of discharge to the tailings pond as required under Directive 2006/21/EC, Article 13(6). Cyanide concentrations will then degrade and attenuate within the tailings pond through various processes including complexation, precipitation, adsorption, oxidation, reaction with sulphur, volatilization, biodegradation and hydrolysis (www.cyanidecode.org). Slurry tailings will be pumped to the tailings impoundment and pond water will be recycled back to the process plant via a reclaim barge and pipeline. The impoundment will be designed to minimize seepages. Monitoring wells will be placed hydraulically down-gradient of the tailings impoundment to monitor and collect seepage, which would be pumped back to the pond if necessary. The tailings impoundment will be fenced to prevent livestock access and other measures may need to taken to deter birds, depending on the pond water quality.

Any additional makeup water will be obtained from abstraction of groundwater or surface water, depending on results from more water balance studies that will be completed during the subsequent stages of the project.

20.2.3 Social and Environmental Management

There are a number of specific social and environmental design criteria that should to be implemented to meet the conservation and visual landscapes objectives of the surrounding area. The following recommendations were made by SLR based on their Landscape and Visual Baseline Study (SLR 2011b):

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Design buildings in colours, styles and size similar to those in the area and use earth sheltered buildings if the building will exceed the size of a typical farm shed; Use earth berms and stonewalls to screen development and maintain the character of the surrounding landscape; Haul roads should be designed to follow existing field boundaries and landscape contours; Retain existing vegetation in accordance with recommendation in BS5837:2005, Trees in Relation to Construction.

Following good industry practice, it is assumed a social, environmental, health and safety PDQDJHPHQWV\VWHPZLOOEHLPSOHPHQWHGWRPHHWWKHFRPSDQ\¶VFRPPLWPHQWVWRSURWHFWWKH environment and the health and safety of the workers and the surrounding communities. The management system and plans should be designed to monitor and maintain permit compliance and a social licence to operate.

20.3 PERMITTING REQUIREMENTS

20.3.1 Permits for Exploration

DGL has an option agreement to allowing for prospecting for gold and silver on DG1, DG2 DG3 and DG4. In addition DGL also have four mineral prospecting licenses issued by the Department of Trade and Industry (DETI) - DG1/08, DG2/08, DG3/08 and DG4/08 to prospect for other minerals in County Tyrone and Londonderry. The Licence and Prospecting Licences permit DGL to undertake exploration activities subject to a number of conditions.

Exploration activities within the project area are generally permitted development and do not require planning permission. However, DGL is required to notify the Department of the Environment (DOE) Strategic Planning Division in writing prior to exploration works taking place; DOE must be informed of project details, including the location of boreholes, the target minerals, details of all plant and operations, the anticipated timescale and confirmation that the works will not be undertaken within a designated area such as an ASSI. Appropriate notifications have been sent to the DOE for exploration works undertaken by DGL.

In addition, the Health and Safety Executive Northern Ireland (HSENI) is also provided with notification of exploration boreholes, in accordance with the 1995 Borehole Regulations. Appropriate notifications have been sent to the HSENI for exploration works undertaken by DGL.

The proposed extension of the existing exploration adit at Curraghinalt does not qualify as permitted development owing to the scale, time frame and nature of the proposed works. Planning permission is required from the DOE Strategic Planning Division for these works.

Dalradian Resources reports that, to date, it has negotiated, or is in the process of negotiating, land access and compensation agreements with all the landowners for the present drilling

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program. Additional access agreements will be required when drilling to the east and west of the known mineralization takes place. Dalradian Resources indicates that it expects to be able to successfully negotiate access to the required land to undertake any future drill programs at the Curraghinalt project.

20.3.2 Permit Requirements for Development

The project is subject to legislative requirements from the European Union, England (UK) and Northern Ireland. Applicable environmental legislation is listed in Table 20.1. In addition, there are a number of international conventions that will apply to this project and need to be considered in further planning.

Table 20.1 Key Applicable Environmental Legislation

An initial meeting with the Planning Service was held February 22, 2011 to introduce the project and the project Scoping Report to the authorities and to initiate the above-noted environmental baseline studies. The project requires consultation and preparation of an Environmental Statement (ES) since it could potentially have significant socio-environmental effects. Baseline information and project engineering is being completed and will be used for the preparation of the ES. At this time it is uncertain how long it will take to permit the project.

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Land acquisition will also need to be negotiated once decisions are made on the final location of infrastructure and facilities.

20.4 SOCIAL AND COMMUNITY ASPECTS, STAKEHOLDER CONSULTATION

Curraghinalt is located on the south side of the Owenkillew River, approximately 7.5 km east of Gortin. There are a number of groups of residential houses and farm holdings along the B46 road south of the Crocknamoghil hill, including Rousky, Casorna and Teebane West. There are also rural residences and farms north of the project along Camcosy Road, south of the Owenkillew River and along Gorticashel road, north of the river. The predominant land use in the local area is agriculture.

Cultural heritage resources will need to be considered in the final siting of project components. There are a number of sites of cultural heritage designation around the project area including two Industrial Heritage Sites (river crossings) and a site monument located in the near vicinity of the exploration adit. There are many other sites in the surrounding area that are generally located along roads, valley bottoms or viewpoints (Figure 20.4; SLR 2011a).

Figure 20.4 Cultural Heritage Sites

Source: SLR, August, 2012

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The largest urban centre is Omagh southwest of the project site. The socio-economic impact assessment is not yet complete, but for the purposes of the PEA, it is anticipated that Omagh would be the closest centre for obtaining minor supplies and services for the project and Belfast would likely be the major supply and services centre. It has also been assumed that there will be sufficient local housing for the majority of operations workforce.

Dalradian Resources is consulting with various agencies and associations listed below and community members as part of the environmental assessment and planning process. Initial stakeholders include:

The Crown Estate Commissioners; Department of the Environment, including The Planning Service; Northern Ireland Environment Agency (NIEA); Rivers Agency; Omagh District Council and neighbouring local authorities, including Environmental Health Officers; Department of Agriculture and Rural Development (DARD), including Rivers, Quality, Veterinary, Forest and Countryside; Department for Regional Development (DRD) - Roads Service; Department of Enterprise, Trade and Investment (DETI) - Northern Ireland Tourist Board (NITB), Geological Survey of Northern Ireland (GSNI), Health and Safety Executive NI (HSE NI); Department of Health, Social Services and Public Safety (DHSSPS); Department of Culture, Arts and Leisure (DCAL); Northern Ireland Electricity (NIE); Northern Ireland Water; Northern Ireland Housing Executive; Fisheries Conservancy Board for Northern Ireland (FCBNI); Loughs Agency; and Industrial Pollution and Radiochemical Inspectorate (IPRI).

20.5 MINE CLOSURE REQUIREMENTS

The objective of the mine closure plan is to remove and close down activities in a manner that ensures public safety and to reclaim the land to a usable state consistent with the surrounding land use objectives. The visual landscape objective is to minimize disruption of the outstanding natural beauty of the area during operations and restore this value at closure.

Closure will consist of plugging and securing underground openings, removing to licensed facilities any hazardous and contaminated materials from site, decommissioning and demolition of facilities and buildings, and reclaiming waste rock and tailings facilities. It has been assumed that the electrical substation is an infrastructure asset that will be left in place post-closure and owned by the utility.

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An adit plug will be engineered and constructed at closure to prevent access to the underground workings and allow the underground workings to flood. Mine flooding will slow down the oxidation of remaining sulphides exposed within the drifts and so reduce acid production and help prevent contaminants in discharges.

Unused fuel and reagents will be removed from site and either sold or disposed of at an approved hazardous waste facility. Used oil and oil filters, fuel storage tanks, and contaminated soil will be removed from site and disposed of at an approved hazardous waste facility.

The tailings facility seepage and surface waters will be monitored. Seepages will be collected and pumped back to the pond if the quality is not acceptable for release. Tailings pond water will be monitored and pumped back to the plant for treatment if necessary until the pond water is acceptable for discharge to the environment. Once acceptable, a spillway will be constructed in the dam and the dams and beach areas capped with overburden and revegetated. Similarly, any non-acid generating waste rock stored on surface will be capped with overburden and revegetated. It is assumed that any potentially acid generating (PAG) waste rock will be used in backfill and flooded after mine closure to minimize oxidation.

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21.0 C API T A L A ND OPE R A T IN G C OSTS

21.1 CAPITAL COSTS

Capital expenditures and capitalized development costs for the base case are summarized as initial and sustaining costs in Table 21.1. The estimates are expressed in second quarter 2012 Canadian dollars, without escalation, unless otherwise noted. The expected accuracy of the estimates is ± 30%.

Table 21.1: Capital Cost Summary

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Capitalized Development 15,745 64,925 Mining Equipment 14,770 33,058 Processing 50,743 - Infrastructure 49,833 16,301 Indirect Costs 12,927 - 2ZQHU¶V&RVWV 17,386 - Contingency 38,341 - Total 199,745 114,284

Expressed as US dollars, initial and sustaining capital amounts to US$192.1 million and US$109.9 million, respectively.

21.1.1 Mine Development

The initial capital cost estimate for mine development comprises haulage access, cross-cuts and ventilation systems development, as well as pre-production operating costs which are capitalized. Sustaining capital costs include only direct costs for haulage access, cross-cuts and ventilation raises. Table 21.2 shows a breakdown of these development costs.

Table 21.2 Capital Cost Summary - Mine Development

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Ramp 846 16,085 Ramp's By Passes 17 320 Ramp's Safety Bays 13 245 Sumps in Ramp 13 245 Cross Cuts 2,185 45,341 Sumps in Cross Cuts 21 247 Vent, Ore and Waste Raise/Pass 21 2,444 Other pre-production mining 12,630 - Total 15,745 64,925

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21.1.2 Mining Equipment

The initial capital cost estimate for mining comprises fleet purchase costs, as shown in Table 21.3. Sustaining capital comprises the replacement of this equipment at appropriate intervals over the life of the mine.

Table 21.3 Capital Estimate - Mining Equipment

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Drilling equipment, incl. jumbos 4,798 11,208 LHD (3 and 4 m3) 2,229 7,212 Truck (30t payload) 1,864 4,660 Ancillary vehicles 1,120 1,830 Pickup trucks 270 810 Fans 270 203 Pumps 90 68 Other equipment 944 4,817 Portal construction 185 0 Pastefill plant 3,000 2,250 Total 14,770 33,058

21.1.3 Processing Plant

The capital estimate for the processing plant breaks down as shown in Table 21.4. Ongoing maintenance is covered by operating costs, so there is no sustaining capital forecast.

Table 21.4 Process Plant Capital Estimate

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Crushing plant 7,382 - Grinding, incl. Conc. grinding 19,572 - Dewatering 5,368 - Common systems 13,421 - Leach/Goldwin Area 5,000 - Total 50,743 -

21.1.4 Infrastructure

7KHSURMHFW¶VLQLWLDOLQIUDVWUXFWXUDOUHTXLUHPHQWVDUHHVWLPDWHGWREHDVVKRZQLQTable 21.5. Capital for the establishment of the tailings storage facility is an average of the estimated cost for the three most highly recommended sites. Sustaining capital comprises a phased expansion of this tailings storage facility, in line with the recommendation by the consultant.

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Table 21.5 Infrastructure Capital Estimate

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Access Road 200 - Site Preparation 2,000 - Power Supply 4,053 - Power Distribution 5,270 - Power - Emergency Supply 673 - Water Supply 2,500 - Administration Complex 1,280 - Mine Dry Complex 2,552 - Maintenance Shop 3,157 - Warehouse & Laydown Area 1,760 - Fuel Tank Farm & Fuelling Station 850 - Sewage System 680 - Fire Water System 560 - Communications 290 - Guard House 250 - Fencing 190 - Solid Waste Disposal & Recycling 40 - Tailings Storage Facility 23,528 16,301 Total 49,833 16,301

21.1.5 ,QGLUHFW&DSLWDO2ZQHU¶V&RVWDQG&RQWLQJHQF\

The indirect capital cost estimate is shown in Table 21.6.

Table 21.6 Indirect Capital Cost Estimate

Initial Capital Sustaining Capital A rea Cost ($ 000) Cost ($ 000) Vendor Engineering & Representatives 1,015 - Construction Equipment 1,522 - EPCM 7,612 - Equipment Spares 1,510 - General Site Costs 1,269 - Sub-total Indirect Costs 12,927 -

General (incl. First Fills) 6,180 - Insurance 507 - Commissioning/Training 2,537 - Owners Site costs 2,768 - Mine Rehabilitation and Closure 5,393 - Sub-WRWDO2ZQHU¶V&RVWV 17,386 -

Contingency 38,341 - Total 68,654 -

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Indirect costs include EPCM costs, estimated to be 15% of the direct capital cost estimate for the process plant. 3URYLVLRQ LV DOVR PDGHIRUYHQGRU¶V UHSUHVHQWDWLYHVGXULQJFRQVWUXFWLRQ and commissioning, as well as the temporary hire of construction equipment (e.g., mobile cranes) and general site costs.

2ZQHU¶V FRVWV LQFOXGH ILUVW ILOOV RI UHDJHQWV DQG FRQVXPDEOHV LQ WKH SODQW FRQVWUXFWLRQ LQVXUDQFH FRPPLVVLRQLQJ UHFUXLWPHQWWUDLQLQJ FRVWV DQG RZQHU¶V VLWH FRVWV LQFOXGLQJ VLWH management and supervision).

Mine rehabilitation costs are estimated to be $7.5 million, incurred following closure of the mine in Year 16. This cost is discounted back to the present at 2% annually to find the amount of the environmental bonding expected to be required for the project. The discounted amount of $5.4 million is then reflected in the cash flow as a cost incurred in two tranches, 30% prior to construction and 70% prior to production start-up.

A total contingency of $38.3 million includes $7.6 million for mining, $15.2 million in the plant, $12.5 million for infrastructure, and $3.0 million for indirect costs. Overall, the contingency equates to approximately 24% of the initial capital cost including indirects.

21.2 OPERATING COSTS

21.2.1 Mine Operating Costs

Mine operating costs are inclusive of supervision, operating and maintenance labour, spares DQGFRQVXPDEOHVIRUWKHRZQHU¶VIOHHWRIPRELOHSURGXFWLRQKDXODJHDQGVXSSRUWHTXLSPHQW ventilation, dewatering and backfill. Provision is made for rehandling of stockpiled ore, where appropriate.

Development carried out in waste rock for haulage ramps, ventilation raises and cross-cuts will utilize the same fleet of equipment. However, the direct costs of this work are treated as sustaining capital expenditure and so are excluded from operating expenses.

21.2.2 Processing Operating Costs

Processing costs include crushing ROM material, milling, leaching of the pulp, carbon adsorption, stripping, electrowinning and smelting of gold doré, cyanide destruction and disposal of barren tailings. Reagents and process consumables, power, operating and maintenance labour and spare parts, and laboratory costs are considered.

21.2.3 General and Administration Costs

General and Administration costs include site management, supervisory and technical staff, office running costs, environmental management and other costs, including overheads.

Estimated cash operating costs over the life of the project are summarized in Table 21.7.

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Table 21.7 Summary of Life-of-Mine Operating Costs

Labour - Metallurgy 5,180 0.53 - Laboratory 6,963 0.71 - Production 27,417 2.81 - Maintenance 14,037 1.44 Power 35,561 3.65 Maintenance 18,346 1.88 Crushing 759 0.08 Grinding 33,760 3.46 Reagents 34,835 3.57 Miscellaneous 13,376 1.37 Sub-total Processing 190,235 20.19

Labour 21,903 2.25 Equipment Maintenance 6,646 0.68 Mobile Equipment Operation 1,472 0.15 Environmental & Social 7,594 0.78 G&A (other) 43,770 4.49 Sub-total General and Administrative 81,385 8.64

Total Operating Costs 1,021,431 108.38

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22.0 E C O N O M I C A N A L YSIS

22.1 BASIS OF EVALUATION

Micon has prepared its assessment of the Project on the basis of a discounted cash flow model, from which Net Present Value (NPV), Internal Rate of Return (IRR), payback and other measures of project viability can be determined. Assessments of NPV are generally accepted within the mining industry as representing the economic value of a project after allowing for the cost of capital invested.

22.2 MACRO-ECONOMIC ASSUMPTIONS

22.2.1 Exchange Rate and Inflation

Unless otherwise stated, all results are expressed in Canadian dollars. Cost estimates and other inputs to the cash flow model for the project have been prepared using constant, second quarter 2012 money terms, i.e., without provision for escalation or inflation. Using trailing 36-month averages, exchange rates of CDN$1.04/US$ and CDN$1.62/GBP are applied in the base case.

22.2.2 Weighted Average Cost of Capital

In order to find the NPV of the cash flows forecast for the project, an appropriate discount factor must be applied which represents the weighted average cost of capital (WACC) imposed on the project by the capital markets. The cash flow projections used for the valuation have been prepared on an all-equity basis. This being the case, WACC is equal to the market cost of equity, and can be determined using the Capital Asset Pricing Model (CAPM):

where E(Ri) is the expected return, or the cost of equity. Rf is the risk-free rate (usually taken to be the real rate on long-term government bonds), E(Rm)-Rf is the market premium for HTXLW\ FRPPRQO\HVWLPDWHGWREHDURXQG DQGEHWD ȕ LVWKHYRODWLOLW\RIWKHUHWXUQVIRU the relevant sector of the market compared to the market as a whole.

Figure 22.1 illustrates the real return on Canadian long bonds computed by the Bank of Canada, taken as a proxy for the risk-free interest rate. Recently, this has dropped from around 2.0% to less than 0.5%. Nevertheless, it is generally accepted that using a long-term average rate will give a more reliable estimate of the cost of equity. Micon has therefore

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used a value of 2.0% for the base case risk free rate, close to the real rate of return averaged over 10-years.

Figure 22.1 Real Return on Canadian Long Bonds (Source: Bank of Canada)

Taking beta across this sector of the equity market to be in the range 0.7 (for instance, some large gold producers) to 1.7 (for diversified base metal groups) with a mid-range of 1.2 (typical for the mining sector), CAPM gives an estimated cost of equity for the Curraghinalt project of between 5% and 11%, as shown in Table 22.1. Micon has taken a figure of 8% (i.e., in the middle of this range) as its base case, and provides the results at alternative rates of discount for comparative purposes.

Table 22.1 Estimated Cost of Equity

Range Lower M iddle Upper Risk Free Rate (%) 1.5 2.0 2.5 Market Premium for equity (%) 5.0 5.0 5.0 Beta 0.7 1.2 1.7 Cost of equity (%) 5.0 8.0 11.0

22.2.3 Expected Metal Prices

Figure 22.2 shows the monthly average gold and silver prices over the past six years, together with the 3-year trailing averages. At the end of June, 2012, the three-year trailing averages for each metal were US$1378/oz gold and US$26.28/oz silver, and these metal prices were selected for the base case. These prices were applied consistently throughout the operating period.

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Silver contributes approximately 0.6% of the projected total revenue for the base case, so the impact of changing the silver price forecast is minimal.

Figure 22.2 Monthly Average Gold and Silver Prices since July 2006 (source: Kitco.com)

For comparison, Micon also evaluated the sensitivity of the project to using recent (1 month), and 1, 2, 3, 5 and 10-year price averages. The prices used in each of these cases are shown in Table 22.2. As part of its sensitivity analysis, Micon also tested a range of prices 30% above and below base case values.

Table 22.2 Metal Price Averages

22.2.4 Taxation Regime

Depreciation allowances of 18% are assumed to be taken annually on all project capital, on a declining balance basis.

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22.2.5 Royalty

A royalty of 6% of NSR value has been provided for in the cash flow model.

22.2.6 Selling Expenses

22.3 TECHNICAL ASSUMPTIONS

22.3.1 Mine Production Schedule

Figure 22.3 shows the annual tonnage of development and stoping material mined, as well as the waste rock tonnage, all of which are held reasonably constant over the LOM period.

Figure 22.3 Annual Mining Schedule

In Figure 22.4, the head grades for gold and silver in the mill feed are shown to remain within very narrow ranges over the LOM period. Tonnages milled reflect the drawdown of stockpiled material during the mill startup and again at the end of the LOM period.

As a consequence of steady tonnage, grade and recovery from mill feed, annual production of gold and silver remain steady over the LOM period (Figure 22.5). This chart also shows the annual total cash cost per ounce of gold sales. Total cash costs, including refining charges

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and royalties and net of silver credits, average US$532/oz gold over the LOM period. In Years 5 and 7, costs fall below this level owing to a reduction in development costs in those periods.

Figure 22.4 Annual Processing Schedule

Figure 22.5 Annual Production Schedule

22.3.2 Operating Costs

Direct operating costs average $108.38/t milled over the LOM period, comprised of $79.56/t mining, $20.19/t processing, and $8.64/t general and administrative costs. Figure 22.6 shows these expenditures over the LOM period, compared to the net sales revenue, showing the strong margin maintained over the LOM period.

225

Figure 22.6 Direct Operating Costs

22.3.3 Capital Costs

Pre-production capital expenditures are estimated to total $199.75 million, including $30.5 million for mining equipment and pre-production development, $50.7 million processing, $49.8 million infrastructure, $12.9 million indirect costs, $12.0 million in RZQHU¶V FRVWV D PLOOLRQ SURYLVLRQ IRU FORVXUH DQG UHKDELOLWDWLRQ DQG FRQWLQJHQFLHV totalling $38.3 million.

Figure 22.7 compares annual capital expenditures over the preproduction and LOM periods ZLWKWKHSURMHFW¶VFDVKRSHUDWLQJPDUJLQ

Figure 22.7 Capital Expenditures

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22.3.4 Base Case Cash Flow

The LOM base case project cash flow is presented in Table 22.3 and Figure 22.8.

Table 22.3 Life-of-Mine Cash Flow Summary

L O M total $/t US$/oz ($ 000) M illed Gold Revenue 3,186,233 338.08 1378.03

Figure 22.8 Life-of-Mine Cash Flows

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The project demonstrates an undiscounted pay back of 2.1 years, or approximately 2.4 years when discounted at 8.0%, leaving a production tail of just over 12 years.

Over the LOM period, gross revenues for gold and silver totalling $3,205.8 million are distributed as shown in Figure 22.9. This diagram illustrates the robust nature of the net cash flows forecast to be generated by the project, both on a before- and after-tax basis.

Figure 22.9 Distribution of Gold and Silver Revenues

Annual cash flows are presented in Table 22.4 (over).

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Table 22.4 Base Case Life of Mine Annual Cash Flow

Operating Margin 207.57 846.05 1,956,209 - (23,919) 151,563 142,805 128,968 119,548 149,016 122,615 138,540 118,385 125,025 131,214 135,460 124,762 125,102 119,491 114,127 33,508 -

Change in Working Cap - - - - - 22,153 323 (159) 330 (208) 367 (1,457) 815 330 (158) 224 (126) (494) 709 (513) (3,400) (18,735)

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22.3.5 Base Case Evaluation

The base case evaluates to an IRR of 51.7% before taxes and 41.9% after tax. At a discount rate of 8.0%, the net present value (NPV8) of the cash flow is $664.6 million (US$ 639 million) before tax and $485.3 million (US$ 467 million) after tax.

Table 22.5 presents the results in Canadian dollars at comparative annual discount rates of 5%, 8% and 11%. Stated in US dollars, at annual discount rates of 5% and 11%, the after-tax NPV of the project is US$ 655 million and US$336 million, respectively.

Table 22.5 Base Case Cash Flow Evaluation

22.4 SENSITIVITY STUDY

22.4.1 Capital, Operating Costs and Revenue Sensitivity

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,Q0LFRQ¶VDQDO\VLVDSSO\LQJDQLQFUHDVHRIPRUHWKDQLQERWKFDSLWDODQGRSHUDWLQJFRVWV simultaneously would be required to reduce NPV8 to near zero, further demonstrating the robust nature of the project economics. Figure 22.10 shows the results of changes in each factor separately.

Figure 22.10 Sensitivity Diagram

22.4.2 Metal Price Sensitivity

Figure 22.11 Sensitivity to Metal Prices

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Table 22.6 Sensitivity to Metal Prices

Pricing Period 1-month 1-year 2-year 3-year 5-year 10-year average average average average average average Pre-tax IRR % 60.1 63.1 57.1 51.7 41.1 24.9 After-tax IRR % 48.6 51.0 46.2 41.9 33.4 20.2

Pre-tax NPV 8 C DN$ 000 818,095 874,215 762,681 664,648 478,875 218,744

After-tax NPV 8 601,704 644,264 559,678 485,330 344,441 147,160

Pre-tax NPV8 US$ 000 786,630 840,591 733,348 639,084 460,456 210,330

After-tax NPV8 578,561 619,485 538,152 466,664 331,194 141,500

At US$1,200/oz gold, the project has an IRR of 39.1% pre-tax and 31.8% after tax.

22.4.3 Sensitivity to Process Flowsheet

Micon conducted a trade-off study to determine the optimum process flowsheet for the Curraghinalt deposit. The four flowsheets considered were:

A comparison of the cash flows for each of these options suggests that Option A, Milling/CIL provides the best overall economic return, and so this was selected as the base case for this study. Figure 22.12 shows the NPV and IRR for each of the four options.

232

Figure 22.12 Sensitivity to Process Flowsheet

22.5 CONCLUSION

Micon concludes that this study demonstrates the viability of the project as proposed, and that further development is warranted.

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23.0 A DJA C E N T PR OPE R T I ES

Tully (2005) lists the following adjacent properties to the Curraghinalt deposit, Cavanacaw, Golan Burn and Glenlark.

The Cavanacaw deposit is located on the 189 km2 licence OM 1/09 held by Galantas Gold Corporation (Galantas). The processing plant was reported by Galantas as commissioned in January, 2007. An NI 43-101-compliant resource estimate was prepared by ACA Howe International Limited dated May 28, 2008 and filed on SEDAR on July 18, 2008. The reported contained gold by Galantas was: Measured - 16,000 oz Au; Indicated - 88,000 oz Au and 295,800 oz Au Inferred, as outlined in Table 16.1.

Table 23.1 Galantas Mineral Resource Estimate

Measured Indicated Inferred Au Contained Au Contained Au Contained

Tonnes G rade Au Tonnes G rade Au Tonnes G rade Au (g/t) (oz) (g/t) (oz) (g/t) (oz) Kearney 78,000 6.35 16,000 350,000 6.74 76,000 730,000 9.27 218,000 Elkin¶s 113,000 3.3 12,000 29,000 3.82 3,600 Kerr 60,000 4.03 7,800 Joshua¶s 160,000 3.96 20,400 Gormley 115,000 6.57 24,300 Garry¶s 40,000 1.27 1,600 Princes 10,000 38.93 12,500 Sammy¶s 30,000 4.26 4,100 Kearney Nth 55,000 1.97 3,500 Total 78,000 6.35 16,000 463,000 5.90 88,000 1,229,000 7.47 295,800

The property is producing at a small scale with production in the second quarter of 2011 reported as 2,500 oz Au, 5,586 oz Ag and 120.6 t Pb.

Micon has not verified this information and it is not necessarily indicative of the mineralization on the Dalradian Resources¶ property.

The Golan Burn property is located on the northwest corner of DG1. Tully (2005) describes the Golan Burn property as follows:

³$W least four, thin auriferous base-metal sulphide veins are present at Golan Burn, 4-km northeast of Gortin, County Tyrone. The veins cut the Neoproterozoic Dalradian Glenelly Formation (South Highland Group) that is composed mainly of calcareous, pale-green to white tourmaline-rich schistose semi-pelite and medium- to coarse-grained psammite. The veins are west-northwest trending and are traceable 600 m along strike. The strike of the vein system projects ESE, directly towards Curraghinalt. Typically the quartz-carbonate-sulphide veins do not exceed 0.5 m in width, however where the vein is present within fault (shear?) zones mineralized intervals can attain widths up to 3.3 m. The best drill intersection at Golan Burn was 3.3 m @ 5.9 g/t Au (internal reSRUW&HOWLF*ROGSOF ´

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Approximately 3 km southeast of Curraghinalt, a vein of mineralization has been identified in the floor of the Alwories Quarry. The vein is up to 1 m wide and grades are reported at between 6 g/t Au and 39.4 g/t Au. Alwories Quarry is located within DG1.

Also within DG1 are the Glenlark and Coneyglen prospects, both of which host stratabound pyrite with zinc, lead and gold.

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24.0 O T H E R R E L E V A N T D A T A A ND IN F O R M A T I O N

$UHSRUWZDVSUHSDUHGDV³DQLQWHUQDOUHYLHZQRWH´E\0-'RQRJKXHRQEHKDOIRI7RXUQLJDQ dated November, 2003. In addition to providing a review of previous metallurgical testwork (see Section 16, above), the report also provided a review of prior economic evaluation data. General economic and metal market conditions at the end of 2003 were significantly different to those pertaining in late 2007, late 2009/2010 or 2011.

Micon is not aware of any additional information or explanation necessary in order to make this Technical Report understandable and not misleading.

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25.0 IN T E RPR E T A T I O N A ND C O N C L USI O NS

Based on its economic evaluation of the base case and sensitivity studies, Micon concludes that this PEA demonstrates the viability of the Curraghinalt project as proposed, and that further development is warranted. The mineral resource estimate on which the PEA is based is open along strike and at depth, and there is good potential for further exploration to expand the resource and to increase confidence in the existing resource.

The key findings of the PEA are:

Pre-tax Internal Rate of Return ("IRR") of 51.7% (after-tax 41.9%) based on a 36- month trailing average gold price of US$1,378 per ounce (after-tax IRR of 31.8% based on US$1,200/oz gold price); Project payback of 2 years from first gold production; After-tax Net Present Value ("NPV") of US$467 million based on a 8% discount rate and a realized gold price of US$1,378 per ounce (US$655 million using a 5% discount rate); Initial capital expenditures of approximately US$192 million prior to production start- up (including contingencies of $36.9 million), with sustaining capital of US$110 million for a total Life of Mine ("LOM") capital spend of US$302 million; A 15 year mine life with average LOM total cash costs of US$532 per ounce, or US$125 per tonne milled, including royalties, refining costs and by-product credits of US$8.24/oz gold; LOM gold production of 2.223 million ounces; Average mined grade of 8.1 g/t gold. Processing at a rate of 1,700 tonnes per day and producing approximately 145,000 ounces gold per year using a conventional flowsheet of crushing, grinding, cyanidation and conventional tailings disposal; Underground mining using mechanized longhole methods with ramp access and truck haulage; and, The mine plan considers 89% of the November, 2011 Micon resource estimate, of which 83% is inferred.

The PEA is preliminary in nature. It includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the results of the PEA will be realized.

25.1 MINERAL RESOURCES

The PEA is based on the proposed mining and processing of the combined measured, indicated and inferred mineral resources as defined in Section 14. Mineral resources that are not mineral reserves do not have demonstrated economic viability. CIM (2010) defines a mineral reserve as the economically mineable part of a Measured or Indicated mineral resource demonstrated by at least a preliminary feasibility study. No mineral reserves have been estimated for the Curraghinalt deposit.

237

The Curraghinalt deposit represents typical mesothermal gold mineralization with several phases of mineralized fluids being injected through shear structures. The mineralization occurs in several parallel veins and sub-veins or veinlets which have been traced by trenching and drilling over a strike length of approximately 2 km.

As part of the PEA, the November 30, 2011 mineral resource estimate was amended to estimate silver and copper grades into the block model. This was done to allow for the contribution of silver to the cash flow for the PEA and to model cRSSHU¶V HIIHFW RQ F\DQLGH FRQVXPSWLRQ Silver and copper grades were estimated using the same techniques as described for gold above. No copper concentrate is produced and no revenue from copper is in the base case cash flow. Silver accounts for approximately 1% of revenue in the economic analysis.

The amended Curraghinalt deposit mineral resource estimate showing copper and silver grades is set out in Table 25.1 below. No changes in tonnages or gold grades have occurred. No gold equivalent calculations were made.

Table 25.1 Curraghinalt Mineral Resource Estimate Amended to Include Silver and Copper

The mineral resource estimate is compliant with the current CIM standards and definitions as required under NI 43-101 and is, therefore, reportable as a mineral resource by Dalradian Resources. However, the reader should be cautioned that mineral resources are not mineral reserves and do not have demonstrated economic viability.

The stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues,

238

unless stated in this report. There are no known mining, metallurgical, infrastructure or other factors that materially affect this mineral resource estimate.

The mineral resource reported above has been estimated largely within the central part of the Curraghinalt deposit. There are sufficient indications, based on mapping, recent drilling and geochemical sampling, of exploration potential in either direction, along strike from the currently defined deposit. Regional geochemical and structural studies clearly indicate that there are parallel trends of mineralization in conformity with the Curraghinalt trend.

There is potential to further increase the resource tonnage by stepping out from the presently identified mineralized zones and in the evaluation of the other possible deposits within the surrounding area. Exploration conducted so far in the adjoining areas has provided positive results. The Curraghinalt deposit is located on the DG1 mineral licence but the potential for discovery of new, similar mineralization may extend as far as the DG3 and DG4 licences (see Figure 4.1)

25.2 MINING

25.3 MINERAL PROCESSING

The PEA considers four possible process flowsheets for the extraction of gold from the Curraghinalt resource. Of these, Option A, which consists of crushing, grinding, whole-ore cyanidation and conventional tailings disposal, was considered the most effective and has been used as the base case for project evaluation. Nevertheless, Micon considers additional testwork, including optimizing grind, reagent strengths and retention times are required, and pilot studies are required before flowsheet development and design specifications can be finalized.

Total gold recoveries, based on existing metallurgical test work, are expected to be approximately 92%.

239

25.4 INFRASTRUCTURE AND CAPITAL COSTS

Infrastructure required for the project has been identified and is provided for in the evaluation. A site-specific layout has not been developed, though, pending discussion with affected landowners over the siting of these works.

Initial capital expenditures total approximately US$192 million, inclusive of a US$37 million contingency as set out in Table 25.2. LOM sustaining capital totals approximately US$110 million. Sustaining capital consists of capitalized waste development after the initial production start-up, major equipment replacement and tailings expansions. Mining sub-level development cost is included in the operating cost.

Table 25.2 Curraghinalt Initial Capital Expenditures

Year Start Up Start Up Start Up (US$ 000) Minus 2 Minus 1 Capital Preproduction - 15,139 15,139 Mining Equipment - 14,202 14,202 Processing Capital 14,638 34,154 48,792 Infrastructure 10,029 37,887 47,916 Indirect Capital 16,517 49,497 66,014 Total 41,184 150,879 192,063

A small proportion of the waste rock samples tested were acid-generating, so it will be necessary for such material to be segregated and stowed as underground backfill. On closure of the mine, the mine entrance will be plugged and flooding of the workings will then minimize further oxidation.

240

26.0 R E C O M M E ND A T I O NS

26.1 PROPOSED EXPLORATION AND DEVELOPMENT PROGRAM

Dalradian Resources has proposed a multifaceted exploration and development program for advancing the Curraghinalt deposit. The proposed work is intended to further explore the extensions of the mineralization, infill gaps in the current resource, continue development work to increase confidence in the resource, test proposed mining methods, and further refine the process flowsheet. Additional exploration work is planned to advance other targets in the Dalradian Supergroup but is not considered in the proposed budget, presented here.

26.1.1 Curraghinalt Resource Drilling and Update

The Curraghinalt deposit is open both along and across strike and to depth. There is scope for identifying additional veins both to the north and south of the known mineralization, as well as extending the known gold-bearing veins along strike to the east and west and to depth. An approximately 8,100 m drill program that continues to test the extensions of the Curraghinalt mineralization, where high grades are considered most likely to be found, is currently underway.

An additional 8,200 m of infill drilling is planned which will fill in several gaps in the resource that currently exist, and reduce the spacing between drill holes used in the resource. Following completion of this program it is planned to update the mineral resource estimate for the Curraghinalt deposit.

26.1.2 Curraghinalt Development

With the completion of this PEA, work will focus on preparation of a planning application to extend the underground workings. In the timeframe covered by the current budget 1,000 m of underground development is proposed. Dalradian Resources plans eventually to complete some 2,000 m of underground drifting on the 106-16, Bend and Crow veins and extend the No 1 and T-17 veins to the east of the current adit cross cut. The purpose of this work will be to confirm the grade and tonnage and continuity of the veins in an underground setting, further evaluate underground ground conditions, test proposed mining methods, and access additional material for further metallurgical testing. Dalradian Resources also plans to continue with the metallurgical testing of the Option A and Option D flowsheets and evaluate the tailings properties of those two flowsheets.

Work will continue in the gathering of environmental baseline data to support an ES.

26.1.2.1 Geotechnical and Mine Design

241

accurately characterized and described. This allows the rock mass attributes, parameters and ratings to be assigned at a later stage enabling the classification to be made in any of the proposed rock mass classification systems (e.g., RMR, Q-system or MRMR).

Regarding development of the mine design, Micon recommends that:

26.1.2.2 Metallurgical

Future metallurgical testwork should be considered upon review of the final testwork report currently being prepared by G&T.

Based on a review of reports leading up to the current testwork underway, it is recommended that future testwork should include the following:

Pre-aeration and lead nitrate addition should be included in the future leach tests to determine the effect on leach residence time and cyanide consumption.

Tests for graphitic carbon in the ore should be investigated.

Additional Bond ball mill work index tests should be conducted

242

Alternative flotation reagents should be investigated in future copper flotation testwork, since the vendor of the 3418A collector used in the G&T copper flotation testwork is not accepting fresh orders for this product until their new facility opens in 2014.

Carbon in Leach (CIL) testwork should be performed.

Testwork with higher cyanide concentrations should be conducted.

Confirmatory testwork on a variety of representative feed samples once a preferred flowsheet is established

26.1.2.3 Tailings and Water

26.1.2.4 Other Items

In this PEA, the estimated unit cost for electrical power is based on best available information. As the project develops, Dalradian Resources should obtain project-specific pricing.

26.1.3 Budget

0LFRQ KDV UHYLHZHG 'DOUDGLDQ 5HVRXUFHV¶ SURSRVHG EXGJHW IRU WKH &XUUDJKLQDOW GHSRVLW WR complete the programs described above (Table 26.1) and finds it to be reasonable and warranted in light of the data presented and observations made in this report. Micon recommends that Dalradian Resources proceed with the program.

243

Table 26.1 Exploration and Development Program Summary Budget

Cost Activity Metres ($ millions) Exploration Resource Drilling and Update 16,300 6.59

Development Program Engineering 0.83 Underground Development 1,000 6.41 Environmental 0.54 Permitting 0.92 CSR 0.26 Health and Safety & G&A 1.43 Subtotal 10.39 Total 16.98 5% Contingency 0.85 G rand Total 17.83 The budget presented in Table 26.1 is a base case scenario and may be modified. This budget does not include additional exploration programs planned for Northern Ireland and elsewhere. Nor does it include general corporate costs.

7KLV UHSRUW WLWOHG ³A Preliminary Economic Assessment for the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´, with an effective date of July 25, 2012, and prepared for Dalradian Resources, was completed by the following authors:

MICON INTERNATIONAL LIMITED

B. T. Hennessey {signed and sealed}

B. Terrence Hennessey, P.Geo. Vice President 6th September, 2012

B. Foo {signed and sealed}

B. Foo, P.Eng. Senior Mining Engineer 6th September, 2012

244

%'DPMDQRYLü^VLJQHGDQGVHDOHG`

%'DPMDQRYLü3(QJ Senior Metallurgist 6th September, 2012

A. Villeneuve {signed and sealed}

A. Villeneuve, P.Eng. Senior Assoc. Mining Engineer 6th September, 2012

C. Jacobs {signed and sealed}

C. Jacobs CEng MIMMM Vice President 6th September, 2012

245

27.0 R E F E R E N C ES

Archibald, S. M. And Williams, V. (2009). An Overview of the VMS Potential of the Tyrone Volcanic Group. An internal PowerPoint presentation to Dalradian Resources.

Boland, M. Curraghinalt, 1997. Gold Deposit Geology and Resource Status May 1997 - Updated Nov. 1997, Unpublished Report for Ulster Minerals Limited, Geological Survey of Northern Ireland.

Coller, D., 2007. Review of the Structure of the Curraghinalt Gold Vein System, prepared for Aurum Exploration Services, June, 2007.

CSA, 1997. Polygonal Resource Calculation of a 1.25 m Minimum Mining Width at the Curraghinalt Deposit, prepared for Ennex International, May, 1997.

Department of Enterprise, Trade and Investment, Minerals and Petroleum Exploration and Development in Northern Ireland, 1997-2000.

Donoghue, M. J., November, 2003. Curraghinalt Gold Deposit Review Report, prepared for Tournigan Gold.

* 7³$Q,QYHVWLJDWLRQLQWRWKH5HFRYHU\RI*ROGDQG6LOYHUIURPWKH7\URQH Project prepared for Dalradian Resources Inc. Project 13471-)LQDO5HSRUW$SULO´

Golder Associates UK, ³&XUUDJKLQDOW*ROG3URMHFW- Tailings Management Facility Site 6HOHFWLRQ6WXG\´5HSRUW1R$'HFHPEHU

Hennessey, B. T. and Mukhopadhyay, D. K., 2010. A Mineral Resource Estimate for the Curraghinalt Gold Deposit and a Review of a Proposed Exploration Program for the Tyrone Project, County Tyrone and County Londonderry, Northern Ireland. An NI 43-101 Technical Report for Dalradian Resources Inc., 122 pp. Filed on SEDAR, www.sedar.com.

Hennessey, B. T. and Mukhopadhyay, D. K. 2012. An Updated Mineral Resource Estimate For The Curraghinalt Gold Deposit, Tyrone Project, County Tyrone And County Londonderry, Northern Ireland. An NI 43-101 Technical Report for Dalradian Resources Inc., 134 pp. Filed on SEDAR, www.sedar.com.

,QWHUQDWLRQDO 0HWDOOXUJLFDO DQG (QYLURQPHQWDO ,QF ³6SHUULQ 0Runtain Gold Project, /DERUDWRU\7HVW:RUN5HVXOWV´

/DNHILHOG5HVHDUFK³$QLQYHVWLJDWLRQRIWKHUHFRYHU\RIJROGDQGVLOYHUIURPWKHWKUHH Sperrin Mountain gold ore samples submitted by Ennex International PLC (per Richard F. Down) Progress Report 1R´3URMHFW1R/5GDWHG)HEUXDU\

246

/DNHILHOG5HVHDUFKD³$QLQYHVWLJDWLRQRIWKHUHFRYHU\RIJROGDQGVLOYHUIURP6SHUULQ Mountain gold ore samples submitted by Ennex International PLC (per Richard F. Down), 3URJUHVV5HSRUW1R´, Project No. LR 2936, dated January 30, 1986.

/DNHILHOG5HVHDUFKE³$QLQYHVWLJDWLRQRIWKHUHFRYHU\RIJROGDQGVLOYHUIURP6SHUULQ Mountain gold ore samples submitted by Ennex International PLC (per Richard F. Down), 3URJUHVV5HSRUW1R´3URMHct No. LR 2936, dated March 6, 1986.

/DNHILHOG5HVHDUFK³$QLQYHVWLJDWLRQRIWKHUHFRYHU\RIJROGIURP&XUUDJKLQDOWSURMHFW samples submitted by Ennex International plc per R.F. Down, Progress Report No.1 Project No.LR 3588, Lakefield Research, dated April 17, 1989.

Mukhopadhyay, D. K., 2007. Technical Report on the Curraghinalt Property, County Tyrone, Northern Ireland, An NI 43-101 Technical Report for Tournigan Gold Corporation (now Tournigan Energy Ltd.), 87pp. Filed on SEDAR, www.sedar.com.

Peatfield, G.R., 20, January, 2003. Updated Technical Review Report on the Tyrone Mineral Exploration Property (Prospecting Licences UM 11/96 and UM 12/96), County Tyrone, Northern Ireland, prepared for Tournigan Gold Corporation.

SLR, 2011a. Environmental Baseline Study (EBS) Scoping Report. Gold Deposits at Curraghinalt, near Gortin, Co. Tyrone. Prepared for Dalradian Gold Ltd. SLR Ref: 501.000138.00037Rev1.

SLR, 2011b. Landscape and Visual Baseline Report. Curraghinalt Gold Deposit Curraghinalt, Gortin, Omagh, Co. Tyrone. Draft report.

SLR, 2012a. Environmental Baseline Hydrogeology: Interim Report. Prepared for Dalradian Gold Ltd. SLR Ref: 120125.501.0138.00037.04.Rev02.

SLR, 2012b. Waste Rock Characterisation. Curraghinalt Gold Project, Co. Tyrone. Draft report.

Snowden, 2012. Dalradian Resources Inc.: Curraghinalt Gold Deposit, Conceptual Study. Project No. J2056, January, 2012.

Strong D. F. (1981). Notes on Ore Mineralogy of the Buchans District. In The Buchans Orebodies: Fifty Years of Geology and Mining, The Geological Association of Canada Special Paper Number 22, E. A. Swanson, D. F. Strong and J. G. Thurlow editors.

The Claim Group Inc., Base Line Land Audit, Northern Ireland, prepared for Tournigan Gold Corporation, March, 2007 (updated by Aurum Exploration Limited, May, 2007).

247

Tully, John V., January 27, 2005. Technical Review Report on the Curraghinalt Gold Property, Exploration License UM-1/02 (UM-11/96), County Tyrone, Northern Ireland, prepared for Tournigan Gold Corporation.

Thurlow, J. G. and Swanson, E. A., 1981. Geology and Ore Deposits of the Buchans Area, Central Newfoundland. In The Buchans Orebodies: Fifty Years of Geology and Mining, The Geological Association of Canada Special Paper Number 22, E. A. Swanson, D. F. Strong and J. G. Thurlow editors.

US EPA 2007. Aquatic Life Ambient Freshwater Quality Criteria - Copper. 2007 Revision. CAS Registry Number 7440-50-8.

Williams, V., July, 2007. Tournigan Gold Projects Co. Tyrone, Technical Presentation.

Williams, V., (2008). Volume 1, Renewal Report, Licence UM-1/02 - Curraghinalt Licence, Period - 1st January, 2006 to 31st December, 2007. A report issued to the Department of Enterprise, Trade and Investment & Crown Mineral Agent for work completed by Dalradian Gold Limited /Aurum Exploration Ltd.

Williams, V., (2009). Technical Report (Renewal), Licence TG-3/03 (Newtownstewart East), Licence TG-4/03 (Sawel-Dart). A Report issued to the Department of Enterprise, Trade and Investment and the Crown Mineral Agent for work completed by Dalradian Gold Limited /Aurum Exploration Ltd.

Williams, V. and Archibald, S. M., (2009). Proposed Work Programme and Budget for VMS Exploration on DG2. An internal PowerPoint presentation to Dalradian Resources.

White, G., Bennett, J. and Holloway, N. (2008). Technical Report On The Omagh Gold Project, Counties Tyrone And Fermanagh, Northern Ireland. An NI 43-101 Technical Report by ACA Howe International Ltd. for Galantas Gold Corporation. Filed on SEDAR, www.sedar.com.

248

28.0 C E R T I FI C A T ES

249

C E R T I FI C A T E

B. T E RR E N C E H E NNESSE Y

As co-author of this report on certain mineral properties of Dalradian Resources Inc. in Northern Ireland, I, B. Terrence Hennessey, P.Geo., do hereby certify that:

1. I am employed by, and carried out this assignment for:

Micon International Limited Suite 900, 390 Bay Street Toronto, Ontario M5H 2Y2

tel. (416) 362-5135 fax (416) 362-5763 e-mail: [email protected]

2. I hold the following academic qualifications:

B.Sc. (Geology) McMaster University 1978

3. I am a registered Professional Geoscientist with the Association of Professional Geoscientists of Ontario (membership # 0038); as well, I am a member in good standing of several other technical associations and societies, including:

The Australasian Institute of Mining and Metallurgy (Member) The Canadian Institute of Mining, Metallurgy and Petroleum (Member).

4. I have worked as a geologist in the minerals industry for over 30 years.

5. I do, by reason of education, experience and professional registration, fulfill the requirements of a Qualified Person as defined in NI 43-101. My work experience includes 7 years as an exploration geologist looking for iron ore, gold, base metal and tin deposits, more than 11 years as a mine geologist in both open pit and underground mines and over 15 years as a consulting geologist working in precious, ferrous and base metals as well as industrial minerals.

6. I visited Northern Ireland and the Tyrone property during the period November 6 and 7, 2009 and October 3 to 5, 2010 to review exploration results and examine drill core. The property had not previously been visited by me.

7. I am responsible for the preparation of Sections 2 to 12, 14, and the portions of Sections 1, 25 and 26 summarized therefrom, of the Technical Report titled ³A

250

Preliminary Economic Assessment of the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´ and dated 6th September, 2012.

8. I am independent of the parties involved in the transaction for which this report is required, as defined in Section 1.5 of NI 43-101.

9. I have had no prior involvement with the mineral properties in question.

10. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared in compliance with the instrument.

11. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make this report not misleading.

Effective date: July 25, 2012

Dated this 6th day of September, 2012

³%7HUUHQFH+HQQHVVH\´^VLJQHGDQGVHDOHG`

B. Terrence Hennessey, P.Geo. Vice President

251

Certificate of Author

Barnard Foo

As co-author of this report entitled ³A Preliminary Economic Assessment of the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´ZLWKDQHIIHFWLYHGDWHRI-XO\ WKH³7HFKQLFDO 5HSRUW´ I, Barnard Foo, do hereby certify that: 1. I am employed as a senior mining engineer by, and carried out this assignment for: Micon International Limited, Suite 900 - 390 Bay Street, Toronto, ON, M5H 2Y2 Tel. (416) 362-5135 Email: [email protected] 2. I hold the following academic qualifications: Laurentian University, B.Eng., Mining Engineering 1998 University of British Columbia, M. Eng., Rock Mechanics 2007 University of Northern British Columbia, Executive MBA 2010 3. I am a registered Professional Engineers of Ontario (Membership # 100052925); 4. I have worked as a mining engineer in the minerals industry for 14 years; 5. I am familiar with NI 43-101 and, by reason of education, experience and professional registration; I fulfill the requirements of a Qualified Person as defined in NI 43-101. My work experience includes 4 years as an mining engineer in cassiterite, base and precious metal deposits, 5 years in underground and open pit geotechnical engineering and 5 years with in mine design and mining project evaluations for the mineral industry; 6. I visited the Curraghinalt property on April 18-19, 2012; 7. I am responsible for the preparation of Sections 15, 16, 21.1.1, 21.1.2, 21.2.1 and the portions of Sections 1, 25 and 26 summarized therefrom, of the Technical Report. 8. I am independent of Dalradian Resources Inc., 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 in compliance with the instrument; 11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make this report not misleading.

Dated this 6th day of September, 2012

³%DUQDUG)RR´^VLJQHGDQGVHDOHG`

Barnard Foo, M.Eng., P.Eng., MBA. Senior Mining Engineer

252

Certificate of Author

%RJGDQ'DPMDQRYLü

As co-author of this report entitled ³A Preliminary Economic Assessment of the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´ZLWKDQHIIHFWLYHGDWHRI-XO\ WKH³7HFKQLFDO 5HSRUW´ I, Bogdan Damjanoviü, do hereby certify that: 1. I am employed as a senior metallurgist by, and carried out this assignment for: Micon International Limited, Suite 900 - 390 Bay Street, Toronto, ON, M5H 2Y2 tel. (416) 362-5135 email: [email protected] 2. I hold the following academic qualifications: B.A.Sc. Engineering, University of Toronto, 1992. 3. I am a Professional Engineer registered with the Professional Engineers Ontario. (registration number 90420456); Also, I am a professional member in good standing of: The Canadian Institute of Mining, Metallurgy and Petroleum (Member); 4. I have worked in the minerals industry for 20 years; my work experience includes 8 years as a metallurgist on gold, copper/nickel and lead/zinc/gold deposits; and the remainder as an independent consultant when I have worked on a variety of precious and base metal deposits; 5. I do, by reason of education, experience and professional registration, fulfill the requirements of a Qualified Person as defined in NI 43-101; 6. I did not visit the Curraghinalt property; 7. I am responsible for the preparation of Section 13, 17, 21.1.3, 21.2.2 and the portions of Sections 1, 25 and 26 summarized therefrom, of the Technical Report. 8. I am independent of Dalradian Resources Inc., 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 in compliance with the instrument; 11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make this report not misleading.

Dated this 6th day of September, 2012

³%RJGDQ'DPMDQRYLü´^VLJQHGDQGVHDOHG`

%RJGDQ'DPMDQRYLü, P.Eng. Senior Metallurgist

253

C E R T I FI C A T E O F A U T H O R

André Villeneuve

As co-author of this report entitled ³A Preliminary Economic Assessment of the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´ZLWKDQHIIHFWLYHGDWHRI-XO\ WKH³7HFKQLFDO 5HSRUW´ I, Andre Villeneuve, do hereby certify that: 1. I am an Associate Mining Engineer with the firm of Micon International Limited, with offices at Suite 205, 700 West Pender Street, Vancouver, British Columbia; 2. I hold the following academic qualifications: B.Sc. in Mining Engineering , University of Montreal 1983; 3. I am a Professional Engineer registered with the APEG of British Columbia (registration number 19287); 4. I have worked in the minerals industry for 29 years; I have extensive senior level management and executive experience with all the various phases of a mineral development project from advanced exploration stage through mine development and construction to mine start-up and production for gold, silver, base metals and coal. I have been involved since 1989 on a consulting basis on numerous mining projects and operations in the Americas, Africa and Australasia.; 5. I do, by reason of education, experience and professional registration, fulfill the requirements of a Qualified Person as defined in NI 43-101; 6. I visited the Curraghinalt property on April 18-19, 2012; 7. I am responsible for the preparation of Sections 18, 21.1.4 and the portions of Sections 1, 25 and 26 summarized therefrom, of the Technical Report. 8. I am independent of Dalradian Resources Inc., 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 in compliance with the instrument; 11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make this report not misleading.

Dated this 6th day of September, 2012

³$QGUp9LOOHQHXYH´^VLJQHGDQGVHDOHG`

André Villeneuve, P.Eng. Senior Associate Mining Engineer

254

Certificate of Author

Christopher Jacobs

As co-author of this report entitled ³A Preliminary Economic Assessment of the Curraghinalt Gold Deposit, Tyrone Project, Northern Ireland´ZLWKDQHIIHFWLYHGDWHRI-XO\ WKH³7HFKQLFDO 5HSRUW´ I, Christopher Jacobs, do hereby certify that: 1. I am employed as a mineral economist by, and carried out this assignment for: Micon International Limited, Suite 900 - 390 Bay Street, Toronto, ON, M5H 2Y2 tel. (416) 362-5135 email: [email protected] 2. I hold the following academic qualifications: B.Sc. (Hons) Geochemistry, University of Reading, 1980; M.B.A., Gordon Institute of Business Science, University of Pretoria, 2004. 3. I am a Chartered Engineer registered with the Engineering Council of the U.K. (registration number 369178); Also, I am a professional member in good standing of: The Institute of Materials, Minerals and Mining; and The Canadian Institute of Mining, Metallurgy and Petroleum (Member); 4. I have worked in the minerals industry for 30 years; my work experience includes 10 years as an exploration and mining geologist on gold, platinum, copper/nickel and chromite deposits; 10 years as a technical/operations manager in both open pit and underground mines; 3 years as strategic (mine) planning manager and the remainder as an independent consultant when I have worked on a variety of precious and base metal deposits; 5. I do, by reason of education, experience and professional registration, fulfill the requirements of a Qualified Person as defined in NI 43-101; 6. I visited the Curraghinalt property on April 18-19, 2012; 7. I am responsible for the preparation of Sections 19, 20, 21.1.5, 21.2.3 and 22, and the portions of Sections 1, 25 and 26 summarized therefrom, of the Technical Report. 8. I am independent of Dalradian Resources Inc., 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 in compliance with the instrument; 11. As of the date of this certificate to the best of my knowledge, information and belief, the sections of this Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make this report not misleading.

Dated this 6th day of September, 2012

³&KULVWRSKHU-DFREV´^VLJQHGDQGVHDOHG`

Christopher Jacobs, CEng MIMMM Vice President

255

29.0 APPE NDI X 1 U LST E R M IN E R A LS/E NN E X/ST R O N GB O W DRI L L H O L ES

256

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 5170-1 257085.2 386356.0 171.0 225.0 3 54.0 5170-10 257088.0 386319.0 171.0 45.0 5 86.0 5170-11 257006.5 386358.0 171.0 225.0 3 18.0 5170-12 257003.0 386358.2 171.0 235.0 3 19.0 5170-13 257090.5 386326.5 171.0 35.0 3 20.0 5170-14 257092.0 386308.0 171.0 40.0 3 45.5 5170-15 257093.6 386288.3 171.0 45.0 3 67.0 5170-16 256946.0 386369.6 171.0 270.0 3 12.0 5170-17 256940.4 386373.2 171.0 230.0 3 12.5 5170-18 256942.5 386375.0 171.0 50.0 3 8.5 5170-19 256902.7 386380.2 171.0 340.0 3 10.8 5170-2 257086.5 386342.8 171.0 230.0 3 15.0 5170-20 256888.0 386380.4 171.0 245.0 3 12.5 5170-21 256889.8 386383.0 171.0 20.0 3 2.5 5170-22 256916.0 386337.6 171.0 180.0 3 7.0 5170-23 256916.2 386381.4 171.0 0.0 3 25.0 5170-24 256941.4 386372.5 171.0 225.0 3 7.5 5170-25 256916.0 386377.6 171.0 230.0 3 10.5 5170-3 257091.0 386284.8 171.0 219.0 5 29.0 5170-4 257094.5 386250.0 171.0 326.0 45 22.9 5170-5 257094.5 386248.4 171.0 209.0 4 18.3 5170-6 257096.0 386250.0 171.0 46.0 4 85.5 5170-7 257092.3 386285.5 171.0 176.0 -45 29.0 5170-8 257054.5 386347.0 171.0 26.0 3 5.1 5170-9 257049.5 386343.5 171.0 268.0 5 10.6 90-01 257183.0 386515.0 191.0 210.0 -45 35.4 90-02 257181.0 386536.0 189.2 200.0 -45 52.4 90-03 257168.0 386543.0 187.0 218.0 -45 50.0 90-04 257176.0 386573.0 183.0 205.0 -45 79.3 90-05 257244.0 386417.0 209.0 180.0 -45 100.3 90-06 257151.0 386589.0 180.0 205.0 -45 110.8 90-07 257059.0 386642.0 171.0 205.0 -45 63.4 90-08 257132.0 386554.0 183.0 205.0 -45 54.3 90-09 257070.0 386666.0 168.0 205.0 -45 83.8 90-10 257168.0 386620.0 178.0 205.0 -45 125.6 90-11 257105.0 386568.0 183.0 205.0 -45 59.7 90-12 257056.2 386446.9 198.6 180.0 -45 244.5 90-13 256990.5 386455.3 198.6 180.0 -45 145.6 90-14 256939.9 386409.9 198.6 180.0 -45 135.5 90-15 256890.0 386465.0 192.0 180.0 -45 115.2 90-16 256939.6 386459.5 193.8 180.0 -45 96.0 90-17 256748.6 386480.6 183.5 180.0 -45 117.6 90-18 256618.0 386509.0 184.5 180.0 -45 155.1 90-19 256462.3 386446.0 204.8 180.0 -45 132.0 90-20 256750.1 386399.9 197.0 180.0 -45 90.8 90-21 256619.4 386402.7 198.6 180.0 -45 117.7 90-22 256456.3 386532.6 189.9 180.0 -45 142.3 90-23 256314.6 386551.0 191.9 180.0 -45 159.4 90-24 256463.4 386412.0 209.8 180.0 -45 120.9 90-25 256398.8 386456.6 207.8 180.0 -45 133.8 90-26 256514.9 386390.7 208.2 180.0 -45 105.5

257

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 90-27 256513.8 386430.6 203.3 180.0 -45 151.6 90-28 256564.5 386333.0 210.7 180.0 -45 71.0 90-29 256565.5 386369.3 207.2 180.0 -45 89.0 90-30 256998.6 386520.9 188.5 180.0 -45 211.6 90-31 256899.5 386533.7 179.5 180.0 -45 234.5 90-32 256566.4 386406.9 203.5 180.0 -45 132.0 90-33 256348.2 386414.9 216.5 180.0 -45 46.0 90-34 256347.4 386452.6 210.8 180.0 -45 86.3 90-35 256344.4 386489.2 204.5 180.0 -45 140.0 90-36 256399.6 386419.2 212.8 180.0 -45 80.2 90-37 256963.6 386443.6 199.2 183.0 -45 90.3 90-37(24) 257094.5 386440.0 200.0 180.0 -45 123.8 90-38 256913.7 386459.9 193.4 182.0 -45 114.2 90-38(25) 256852.0 386420.0 191.0 180.0 -45 6.0 90-39 256487.4 386366.6 212.3 184.0 -45 62.5 90-39(26) 256850.0 386476.0 190.0 180.0 -45 0.0 90-40 256488.1 386402.1 209.0 180.0 -45 110.0 90-40(27) 257014.5 386452.0 199.0 180.0 -45 109.7 90-41 257036.1 386412.7 205.2 184.0 -45 209.8 90-41(28) 257167.0 386438.0 206.0 180.0 -45 72.9 90-42 256488.5 386437.0 204.7 180.0 -45 149.1 90-43 256949.2 386386.3 207.0 186.0 -45 195.8 90-44 256537.5 386345.4 211.7 182.0 -45 50.0 90-45 256538.3 386382.4 207.7 184.0 -45 90.1 90-46 256539.3 386420.3 203.0 184.0 -45 127.6 90-47 257038.3 386459.9 198.4 184.0 -45 140.7 90-48 256489.2 386471.3 200.6 186.0 -45 164.1 90-49 256998.9 386486.7 194.8 184.0 -45 150.8 90-50 256963.9 386482.8 193.1 182.0 -45 121.9 90-51 256540.9 386451.9 199.9 184.0 -45 173.0 90-52 256431.5 386416.3 211.6 180.0 -45 83.3 90-53 256434.8 386452.9 206.4 182.0 -45 148.5 90-54 256373.4 386441.5 211.1 182.0 -45 88.0 90-55 256373.6 386473.7 206.6 182.0 -45 126.6 90-56 256346.5 386435.9 213.5 180.0 -45 80.8 106-1 257375.0 386330.0 225.0 176.0 -45 100.6 106-10 257146.0 386301.0 220.0 180.0 -45 141.1 106-100 257201.3 386276.0 221.2 180.0 -45 75.1 106-101 257074.3 386189.1 233.1 180.0 -45 72.5 106-102 257281.2 386220.1 227.0 182.0 -45 124.9 106-103 257123.2 386179.5 232.9 180.0 -45 76.9 106-104 257370.7 386123.4 239.3 180.0 -45 116.2 106-105 257172.6 386170.0 232.9 180.0 -45 50.4 106-106 257173.4 386213.3 227.9 180.0 -45 96.1 106-107 257124.5 386214.6 229.5 180.0 -45 105.8 106-108 257076.3 386235.4 226.1 180.0 -45 117.7 106-109 257224.9 386164.7 229.5 180.0 -45 77.3 106-11 257189.0 386341.0 214.6 180.0 -45 162.2 106-12 257265.0 386285.2 221.6 180.0 -45 117.7 106-13 256991.1 386347.3 214.9 180.0 -45 117.7 106-14 257474.5 386205.5 245.9 180.0 -45 189.0 106-15 257375.6 386195.9 236.6 180.0 -45 140.2 106-16 256990.2 386406.4 204.8 180.0 -45 190.5

258

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 106-17 257587.5 386145.2 256.0 180.0 -45 149.7 106-18 257476.7 386150.6 248.3 180.0 -45 125.0 106-19 257807.0 385959.9 271.7 180.0 -45 112.2 106-2 257374.0 386366.0 223.0 180.0 -45 84.1 106-20 257187.2 385980.1 242.1 180.0 -45 77.4 106-21 257186.9 386062.9 236.9 180.0 -45 173.7 106-22 256891.5 386363.5 209.4 180.0 -45 131.8 106-23 257052.6 386275.8 224.3 180.0 -45 69.3 106-24 256890.4 386413.6 201.0 180.0 -45 182.4 106-25 257049.1 386327.6 219.0 180.0 -45 106.7 106-26 257019.9 386338.6 217.2 180.0 -45 112.5 106-27 257035.8 386376.6 210.6 180.0 -45 160.7 106-28 257013.6 386375.5 210.2 180.0 -45 139.9 106-29 256949.9 386349.8 213.6 180.0 -45 137.5 106-3 257241.0 386351.0 215.0 180.0 -45 46.3 106-30 256515.0 386345.0 212.0 180.0 -45 45.7 106-31 256471.0 386372.3 212.0 180.0 -45 56.4 106-32 256426.7 386373.6 216.0 180.0 -45 54.9 106-33 256404.4 386379.3 216.8 180.0 -45 36.9 106-34 256390.2 386395.0 215.7 180.0 -45 31.1 106-35 256369.6 386246.7 236.9 180.0 -45 37.2 106-36 256336.5 386257.0 238.3 180.0 -45 32.3 106-37 256306.6 386258.3 240.0 180.0 -45 34.1 106-38 256282.4 386271.0 240.9 180.0 -45 45.7 106-39 256260.4 386275.9 240.6 180.0 -45 19.8 106-4 257243.0 386375.0 212.0 180.0 -45 50.3 106-40 257004.2 386262.1 226.2 180.0 -45 107.6 106-41 256986.9 386379.2 208.4 180.0 -45 89.1 106-42 256970.0 386397.0 203.5 180.0 -49 64.0 106-43 256970.0 386384.5 206.0 180.0 -40 40.0 106-44 256970.0 386366.8 211.0 180.0 -50 31.0 106-45 257048.5 386372.8 211.1 215.0 -45 50.9 106-46 257043.0 386360.0 213.0 215.0 -45 42.7 106-47 257035.5 386350.0 214.5 215.0 -45 32.9 106-48 257130.0 386420.0 200.0 180.0 -45 135.7 106-49 257121.0 386395.0 206.0 180.0 -45 108.9 106-5 257092.0 386283.0 222.0 180.0 -45 61.0 106-50 257050.0 386360.0 212.0 215.0 -45 58.8 106-51 256890.0 386388.0 205.0 180.0 -45 50.0 106-52 256848.7 386392.6 202.8 180.0 -45 79.3 106-53 257116.1 386315.5 219.2 180.0 -45 110.3 106-54 257115.7 386359.7 214.4 180.0 -45 141.5 106-55 257154.6 386360.2 214.0 180.0 -45 158.8 106-56 256797.7 386402.0 199.7 180.0 -45 110.4 106-57 257494.6 385935.1 262.5 180.0 -45 86.9 106-58 256797.5 386436.0 194.5 180.0 -45 145.4 106-59 257494.6 385944.9 261.6 180.0 -45 78.0 106-6 257091.0 386332.0 217.0 180.0 -45 103.8 106-60 256887.1 386284.3 221.6 180.0 -45 92.0 106-61 257579.1 386054.0 261.0 180.0 -45 64.9 106-62 256887.0 386323.8 215.3 180.0 -45 111.0 106-63 257578.1 386108.3 257.4 180.0 -45 101.7 106-64 256941.8 386258.7 226.1 180.0 -45 63.4

259

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 106-65 257689.4 386104.4 261.2 180.0 -45 68.8 106-66 256942.7 386301.0 220.8 180.0 -45 103.9 106-67 257801.4 385925.3 275.6 180.0 -45 55.2 106-68 256990.6 386300.5 221.3 180.0 -45 118.9 106-69 258004.7 385884.7 267.1 180.0 -45 57.4 106-7 257149.0 386250.0 225.0 180.0 -45 64.0 106-70 256863.0 386295.6 219.2 180.0 -45 83.5 106-71 258007.7 385924.3 264.4 180.0 -45 89.9 106-72 256863.4 386332.9 212.2 180.0 -45 109.8 106-73 256988.7 386249.1 228.2 180.0 -45 75.9 106-74 257019.9 386231.7 229.4 180.0 -45 75.9 106-75 257255.3 386029.5 239.0 192.0 -45 124.2 106-76 257226.3 385951.7 243.5 192.0 -45 47.2 106-77 256600.8 386321.4 211.0 180.0 -45 56.1 106-78 256602.3 386364.7 205.0 180.0 -45 86.9 106-79 256639.8 386371.5 202.1 180.0 -45 47.0 106-8 257090.0 386383.0 211.0 180.0 -45 146.9 106-80 256886.9 386253.6 226.5 180.0 -45 47.5 106-81 256920.1 386242.2 228.5 180.0 -45 38.1 106-82 256919.3 386283.8 222.1 180.0 -45 77.0 106-83 256955.5 386223.9 230.8 180.0 -45 39.0 106-84 256987.8 386212.1 231.8 180.0 -45 40.5 106-85 256919.6 386317.9 218.0 180.0 -45 112.3 106-86 257116.5 386277.6 223.0 180.0 -45 71.5 106-87 257070.7 386361.4 214.4 180.0 -45 120.0 106-88 256866.8 386254.1 225.6 180.0 -45 50.3 106-89 256965.2 386319.8 218.9 185.0 -45 115.9 106-9 257265.0 386235.0 223.0 184.0 -45 71.0 106-90 257089.3 386411.7 206.2 180.0 -45 167.2 106-91 256917.5 386363.0 210.3 183.0 -45 145.3 106-92 257189.4 386218.8 226.8 180.0 -45 51.9 106-93 257217.0 386207.3 226.7 180.0 -45 50.1 106-94 257372.8 386163.1 238.0 180.0 -45 116.2 106-95 257376.2 386231.9 235.1 182.0 -45 182.7 106-96 257324.9 386234.4 230.7 184.0 -45 160.9 106-97 257328.2 386274.7 227.9 184.0 -45 179.8 106-98 257321.7 386194.9 232.3 184.0 -45 114.3 106-99 257261.2 386260.7 222.3 182.0 -45 141.8 106-SH1 258086.7 385574.6 279.5 180.0 -45 40.9 106-SH10 258096.5 385773.7 266.1 180.0 -45 40.5 106-SH11 258097.9 385798.6 264.8 180.0 -45 65.0 106-SH12 258099.1 385823.9 263.4 180.0 -45 60.1 106-SH13 258100.6 385850.3 262.1 180.0 -45 63.0 106-SH14 257802.9 385702.4 289.9 180.0 -45 39.1 106-SH15 257802.7 385728.4 288.4 180.0 -45 60.2 106-SH16 257802.4 385753.1 286.2 180.0 -45 40.3 106-SH17 257802.3 385777.6 285.1 180.0 -45 39.8 106-SH2 258087.9 385599.7 277.5 180.0 -45 62.7 106-SH3 258089.2 385625.0 275.7 180.0 -45 63.9 106-SH4 258090.4 385649.6 274.0 180.0 -45 35.2 106-SH5 258091.7 385673.9 271.9 180.0 -45 34.2 106-SH6 258085.4 385550.2 282.4 180.0 -45 40.1 106-SH7 258092.8 385700.1 270.3 180.0 -45 40.5

260

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 106-SH8 258094.3 385725.6 268.3 180.0 -45 41.0 106-SH9 258095.4 385748.6 267.1 180.0 -45 39.6 03-CT-01 257409.6 386242.8 238.0 180.0 -45 203.5 03-CT-02 257536.6 386129.4 253.2 180.0 -45 150.5 03-CT-03 257616.7 386138.4 256.5 180.0 -45 171.5 03-CT-04 257695.7 386101.0 261.1 180.0 -45 173.0 03-CT-05 257069.1 386422.8 203.4 180.0 -45 162.0 03-CT-06 257080.0 386478.0 195.0 175.6 -48 173.5 03-CT-07 257057.3 386421.2 204.4 181.3 -45 165.0 03-CT-08 257057.4 386485.0 194.3 179.4 -42 204.5 03-CT-09 257058.4 386512.3 191.0 186.2 -43 172.3 03-CT-10 257070.5 386387.3 210.3 178.6 -43 125.5 03-CT-11 257106.8 386343.6 215.9 168.7 -44 187.5 04-CT-12 257134.8 386310.8 219.5 181.7 -44 160.6 04-CT-13 257065.9 386369.7 213.6 188.6 -44 54.0 04-CT-14 257066.2 386371.8 213.3 187.2 -49 53.6 04-CT-15 257047.7 386349.2 216.2 186.5 -45 154.2 04-CT-16 257043.7 386414.9 204.5 199.9 -44 147.0 04-CT-17 257014.1 386480.4 195.4 184.5 -45 222.5 04-CT-18 256950.0 386435.3 199.3 187.3 -45 86.0 04-CT-19 256948.9 386477.5 192.6 175.5 -54 115.6 04-CT-20 256877.6 386446.6 195.1 180.7 -45 91.0 04-CT-21 256876.2 386184.8 234.4 181.2 -44 49.0 04-CT-22 257045.6 386366.8 213.5 198.8 -44 39.6 04-CT-23 257143.5 386468.4 195.3 183.6 -44 284.9 04-CT-24 257137.5 386559.8 181.7 197.4 -61 351.0 04-CT-25 256755.2 386568.5 165.6 192.4 -52 216.0 04-CT-26 257046.4 386772.9 156.7 201.0 -50 474.8 05-CT-27 257135.5 385981.8 245.2 182.9 -50 84.0 05-CT-28 257553.5 385953.0 266.6 176.6 -47 99.0 06-CT-29 257377.5 386301.1 230.4 182.7 -45 240.0 06-CT-30 257436.9 386227.6 241.1 189.3 -47 172.1 06-CT-31 257398.6 386172.2 239.8 185.5 -45 150.0 06-CT-32 257443.5 386147.2 245.0 186.8 -47 120.0 06-CT-33 257404.0 386126.0 242.0 188.8 -49 149.0 06-CT-34 257435.1 386097.8 245.7 190.9 -47 72.0 06-CT-35 257441.3 386304.6 237.0 183.4 -47 263.0 06-CT-36 257507.1 386280.8 243.7 189.3 -47 228.0 06-CT-37 257559.1 386256.7 247.9 190.2 -45 264.0 06-CT-38 257506.6 386209.1 247.3 186.5 -46 180.0 06-CT-39 257503.2 386153.1 250.1 191.8 -47 146.0 06-CT-40 257514.0 386090.0 254.0 193.8 -44 233.8 06-CT-41 257560.4 386196.2 251.9 189.2 -46 207.0 06-CT-42 257600.1 386208.0 252.3 186.3 -45 171.0 06-CT-43 257600.3 386274.6 248.2 199.7 -46 231.0 07-CT-44 257600.0 386136.9 256.1 192.7 -44 131.3 07-CT-45 257558.6 386096.2 256.7 186.7 -45 207.2 07-CT-46 257600.0 386074.9 260.2 195.4 -47 75.0 07-CT-47 257649.6 386216.4 252.9 194.6 -46 230.3 07-CT-48 257650.3 386152.2 256.6 190.1 -45 195.0 07-CT-49 257650.0 386099.0 260.6 197.7 -45 290.0 07-CT-50 257375.1 386536.8 202.7 191.3 -44 398.0 07-CT-51 257342.0 386778.6 156.3 107.3 -46 185.5

261

Drill Hole X Y Z Collar Hole Dip Hole Depth Number Coordinate Coordinate Coordinate Azimuth (degrees) (m) 07-CT-52 257381.4 386942.1 123.9 194.7 -46 182.0 07-CT-53 256936.6 386810.4 147.5 189.5 -45 501.1 08-CT-54 257050.3 386800.2 153.0 195.3 -71 130.0 08-CT-54a 257043.4 386800.5 152.9 210.8 -71 545.5 08-CT-55 256764.5 386753.6 131.5 208.8 -69 633.2 08-CT-56 256907.8 386866.5 135.4 209.7 -71 138.0 08-CT-56a 256907.8 386866.5 135.4 218.4 -72 663.1 08-CT-57 256845.5 386825.3 136.4 213.3 -71 661.0

262

30.0 APPE NDI X 2 D A L R A DI A N R ESO UR C ES DRI L L H O L E IN T E RSE C T I O NS USE D IN 2011 M IN E R A L R ESO UR C E EST I M A T E

263

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 5170-10 55.40 56.06 29.1 Pre-Dalradian 5170-13 1.50 1.60 108.3 Pre-Dalradian 5170-14 20.57 21.00 27.4 Pre-Dalradian 5170-15 40.50 40.74 17.5 Pre-Dalradian 5170-17 6.90 7.00 23.8 Pre-Dalradian 5170-6 74.00 74.40 34.5 Pre-Dalradian 90-02 33.72 34.25 31.8 Pre-Dalradian 90-03 36.12 37.43 31.4 Pre-Dalradian 90-04 66.26 66.45 32.5 Pre-Dalradian 90-05 67.82 68.34 3.6 Pre-Dalradian 90-05 93.45 93.84 27.9 Pre-Dalradian 90-06 28.38 28.49 5.1 Pre-Dalradian 90-06 64.66 64.71 29.3 Pre-Dalradian 90-06 67.14 67.19 23.3 Pre-Dalradian 90-06 75.71 75.96 2.6 Pre-Dalradian 90-06 86.62 86.80 1.4 Pre-Dalradian 90-07 34.30 34.66 3.1 Pre-Dalradian 90-08 32.54 33.54 35.2 Pre-Dalradian 90-08 36.62 36.71 22.3 Pre-Dalradian 90-09 54.87 55.27 18.7 Pre-Dalradian 90-10 113.00 114.12 2.5 Pre-Dalradian 90-10 121.51 121.57 29.8 Pre-Dalradian 90-11 33.00 33.26 2.5 Pre-Dalradian 90-11 33.48 33.82 13.9 Pre-Dalradian 90-12 102.18 103.59 33.0 Pre-Dalradian 90-12 168.95 169.77 28.0 Pre-Dalradian 90-12 171.71 172.00 7.3 Pre-Dalradian 90-13 33.20 33.45 15.1 Pre-Dalradian 90-13 145.00 145.59 5.1 Pre-Dalradian 90-14 99.53 100.11 3.0 Pre-Dalradian 90-14 110.49 111.31 14.1 Pre-Dalradian 90-14 122.05 123.72 9.6 Pre-Dalradian 90-15 82.16 82.30 18.6 Pre-Dalradian 90-16 90.50 91.55 32.2 Pre-Dalradian 90-18 86.90 87.12 1.2 Pre-Dalradian 90-18 98.44 99.61 5.5 Pre-Dalradian 90-19 123.06 124.30 13.9 Pre-Dalradian 90-20 56.93 57.30 54.6 Pre-Dalradian 90-20 79.36 80.38 9.5 Pre-Dalradian 90-21 15.80 15.98 21.6 Pre-Dalradian 90-24 74.45 74.74 8.8 Pre-Dalradian 90-24 84.94 86.00 11.8 Pre-Dalradian 90-25 26.64 26.84 3.1 Pre-Dalradian 90-25 33.79 34.89 9.7 Pre-Dalradian 90-25 97.26 97.54 12.9 Pre-Dalradian 90-25 107.00 108.41 19.8 Pre-Dalradian 90-25 128.35 128.46 42.8 Pre-Dalradian 90-26 79.25 79.48 53.2 Pre-Dalradian 90-26 92.80 93.09 64.3 Pre-Dalradian 90-27 121.64 123.20 31.5 Pre-Dalradian 90-27 126.92 127.00 63.7 Pre-Dalradian 90-27 142.87 143.04 18.6 Pre-Dalradian 90-28 39.58 39.72 12.0 Pre-Dalradian

264

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 90-29 23.10 23.25 8.7 Pre-Dalradian 90-29 68.80 69.15 12.2 Pre-Dalradian 90-29 74.75 75.29 54.6 Pre-Dalradian 90-30 24.71 25.61 19.1 Pre-Dalradian 90-30 163.66 163.80 1.5 Pre-Dalradian 90-30 164.27 166.69 17.2 Pre-Dalradian 90-31 33.20 33.45 15.1 Pre-Dalradian 90-31 145.24 145.59 7.4 Pre-Dalradian 90-32 109.65 110.67 38.2 Pre-Dalradian 90-32 114.60 114.80 28.2 Pre-Dalradian 90-33 37.30 38.50 3.0 Pre-Dalradian 90-34 36.60 36.80 1.9 Pre-Dalradian 90-34 75.25 75.30 25.0 Pre-Dalradian 90-34 77.75 77.88 15.7 Pre-Dalradian 90-35 48.70 48.80 3.5 Pre-Dalradian 90-35 129.10 129.30 145.8 Pre-Dalradian 90-36 54.57 55.80 9.2 Pre-Dalradian 90-36 66.61 66.84 5.1 Pre-Dalradian 90-37 81.84 82.46 7.4 Pre-Dalradian 90-38 74.14 74.20 13.0 Pre-Dalradian 90-39 56.10 58.74 5.6 Pre-Dalradian 90-40 79.01 79.80 18.9 Pre-Dalradian 90-40 98.22 98.90 12.9 Pre-Dalradian 90-41 67.10 67.32 5.5 Pre-Dalradian 90-41 71.40 72.15 24.4 Pre-Dalradian 90-41 127.75 128.33 2.9 Pre-Dalradian 90-41 139.63 139.78 42.8 Pre-Dalradian 90-41 194.65 195.20 7.8 Pre-Dalradian 90-41 199.02 199.55 10.6 Pre-Dalradian 90-42 113.90 115.20 4.5 Pre-Dalradian 90-43 55.52 57.10 14.1 Pre-Dalradian 90-43 62.00 62.75 27.0 Pre-Dalradian 90-43 87.00 90.00 17.3 Pre-Dalradian 90-43 103.41 103.54 14.5 Pre-Dalradian 90-43 138.39 139.03 4.3 Pre-Dalradian 90-44 29.50 30.36 6.6 Pre-Dalradian 90-44 40.66 40.84 27.5 Pre-Dalradian 90-45 73.08 74.50 14.3 Pre-Dalradian 90-45 83.00 83.50 22.4 Pre-Dalradian 90-46 116.55 117.20 11.4 Pre-Dalradian 90-47 101.55 102.20 9.4 Pre-Dalradian 90-48 147.20 147.70 30.3 Pre-Dalradian 90-48 160.75 161.09 12.3 Pre-Dalradian 90-49 124.83 125.50 4.6 Pre-Dalradian 90-50 111.05 112.15 3.9 Pre-Dalradian 90-51 159.78 160.59 23.8 Pre-Dalradian 90-52 63.00 63.43 19.4 Pre-Dalradian 90-52 74.55 74.90 56.8 Pre-Dalradian 90-53 140.20 140.80 5.4 Pre-Dalradian 90-54 76.70 76.93 9.7 Pre-Dalradian 90-54 82.20 82.50 43.3 Pre-Dalradian 90-56 53.85 54.86 15.0 Pre-Dalradian 90-56 71.80 72.04 3.4 Pre-Dalradian

265

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-1 13.55 13.65 5.0 Pre-Dalradian 106-1 23.04 25.25 6.2 Pre-Dalradian 106-1 35.66 35.81 7.2 Pre-Dalradian 106-1 39.30 39.35 3.6 Pre-Dalradian 106-1 39.58 39.61 2.2 Pre-Dalradian 106-10 18.38 18.54 13.6 Pre-Dalradian 106-10 46.60 46.69 26.5 Pre-Dalradian 106-10 59.49 61.39 22.7 Pre-Dalradian 106-10 85.20 85.30 23.0 Pre-Dalradian 106-100 61.98 63.35 9.8 Pre-Dalradian 106-101 22.95 23.15 30.5 Pre-Dalradian 106-101 40.55 41.00 11.7 Pre-Dalradian 106-101 58.65 58.80 11.5 Pre-Dalradian 106-102 43.36 43.50 111.8 Pre-Dalradian 106-102 101.95 102.56 23.8 Pre-Dalradian 106-102 107.38 107.82 69.0 Pre-Dalradian 106-103 61.00 61.50 9.6 Pre-Dalradian 106-105 42.46 42.86 34.8 Pre-Dalradian 106-106 21.15 21.56 24.0 Pre-Dalradian 106-106 76.85 78.91 25.7 Pre-Dalradian 106-107 67.15 67.43 39.0 Pre-Dalradian 106-107 81.81 82.05 9.7 Pre-Dalradian 106-108 59.20 64.00 13.8 Pre-Dalradian 106-108 71.55 72.00 5.3 Pre-Dalradian 106-108 83.35 83.54 9.6 Pre-Dalradian 106-108 106.57 107.40 29.9 Pre-Dalradian 106-11 8.15 8.79 6.7 Pre-Dalradian 106-11 42.82 43.16 15.0 Pre-Dalradian 106-11 73.10 73.25 21.2 Pre-Dalradian 106-11 75.17 75.30 12.1 Pre-Dalradian 106-11 88.62 88.78 6.3 Pre-Dalradian 106-11 111.33 113.00 8.1 Pre-Dalradian 106-11 117.02 117.14 4.1 Pre-Dalradian 106-11 138.48 138.60 2.1 Pre-Dalradian 106-12 67.60 68.41 5.8 Pre-Dalradian 106-12 85.04 85.55 19.8 Pre-Dalradian 106-12 90.83 91.17 18.2 Pre-Dalradian 106-12 96.08 96.55 28.7 Pre-Dalradian 106-13 87.50 88.12 57.2 Pre-Dalradian 106-13 92.64 92.82 15.3 Pre-Dalradian 106-13 107.21 107.64 15.8 Pre-Dalradian 106-14 85.56 85.80 5.0 Pre-Dalradian 106-14 107.90 108.07 3.7 Pre-Dalradian 106-14 111.25 111.36 5.3 Pre-Dalradian 106-14 136.43 137.53 2.3 Pre-Dalradian 106-14 138.76 140.19 2.5 Pre-Dalradian 106-15 67.15 67.60 7.5 Pre-Dalradian 106-15 75.10 75.82 5.7 Pre-Dalradian 106-15 76.94 76.96 23.1 Pre-Dalradian 106-15 80.50 82.11 13.6 Pre-Dalradian 106-15 89.69 90.93 24.1 Pre-Dalradian 106-15 107.43 108.50 109.0 Pre-Dalradian 106-15 111.15 111.25 6.3 Pre-Dalradian

266

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-15 117.80 119.70 6.7 Pre-Dalradian 106-15 122.22 123.48 9.2 Pre-Dalradian 106-15 136.28 136.53 1.6 Pre-Dalradian 106-16 57.28 57.87 95.0 Pre-Dalradian 106-16 116.60 116.88 22.4 Pre-Dalradian 106-16 127.31 127.94 90.2 Pre-Dalradian 106-16 152.36 152.61 3.5 Pre-Dalradian 106-16 174.63 175.49 9.8 Pre-Dalradian 106-16 180.16 181.71 11.2 Pre-Dalradian 106-17 45.70 47.54 6.8 Pre-Dalradian 106-17 57.05 57.12 1.4 Pre-Dalradian 106-17 85.61 87.09 14.9 Pre-Dalradian 106-17 100.51 101.26 1.5 Pre-Dalradian 106-17 113.04 113.89 16.9 Pre-Dalradian 106-17 117.17 117.28 7.8 Pre-Dalradian 106-17 119.64 119.70 1.4 Pre-Dalradian 106-18 40.49 41.63 5.6 Pre-Dalradian 106-18 68.08 69.74 10.9 Pre-Dalradian 106-18 91.18 92.23 4.3 Pre-Dalradian 106-18 106.91 107.07 6.1 Pre-Dalradian 106-19 41.50 43.52 24.0 Pre-Dalradian 106-19 47.55 48.22 1.6 Pre-Dalradian 106-19 59.96 60.09 4.3 Pre-Dalradian 106-2 68.69 70.00 13.7 Pre-Dalradian 106-20 48.95 49.16 5.0 Pre-Dalradian 106-20 51.56 51.78 4.6 Pre-Dalradian 106-21 26.64 26.98 5.5 Pre-Dalradian 106-21 44.53 44.81 6.2 Pre-Dalradian 106-21 45.43 45.61 1.5 Pre-Dalradian 106-21 47.71 47.80 6.1 Pre-Dalradian 106-21 49.38 49.52 74.4 Pre-Dalradian 106-21 82.00 82.15 17.5 Pre-Dalradian 106-21 114.73 115.45 18.8 Pre-Dalradian 106-21 141.00 141.42 47.0 Pre-Dalradian 106-22 41.45 42.92 15.5 Pre-Dalradian 106-22 67.15 67.32 33.8 Pre-Dalradian 106-22 118.48 119.53 6.2 Pre-Dalradian 106-22 121.41 121.59 30.9 Pre-Dalradian 106-23 4.90 5.45 10.6 Pre-Dalradian 106-23 26.50 26.57 1.6 Pre-Dalradian 106-23 37.76 38.90 18.6 Pre-Dalradian 106-24 42.15 44.32 28.0 Pre-Dalradian 106-24 46.62 47.15 11.3 Pre-Dalradian 106-24 163.12 163.61 13.0 Pre-Dalradian 106-25 24.53 25.70 5.6 Pre-Dalradian 106-25 69.74 69.85 5.9 Pre-Dalradian 106-25 70.71 71.17 1.8 Pre-Dalradian 106-25 91.67 91.92 19.3 Pre-Dalradian 106-26 17.36 17.93 12.7 Pre-Dalradian 106-26 72.66 72.91 2.0 Pre-Dalradian 106-26 104.23 104.43 10.8 Pre-Dalradian 106-27 37.79 47.92 28.9 Pre-Dalradian 106-27 68.22 68.60 4.0 Pre-Dalradian

267

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-27 75.29 75.41 17.4 Pre-Dalradian 106-27 113.42 113.69 14.2 Pre-Dalradian 106-28 34.38 36.60 31.2 Pre-Dalradian 106-28 97.00 97.90 3.3 Pre-Dalradian 106-28 107.14 107.47 29.2 Pre-Dalradian 106-29 23.14 25.86 9.7 Pre-Dalradian 106-29 64.52 64.93 14.9 Pre-Dalradian 106-29 73.15 73.40 2.4 Pre-Dalradian 106-29 74.76 74.90 13.9 Pre-Dalradian 106-29 127.84 129.42 18.6 Pre-Dalradian 106-29 130.81 131.06 1.3 Pre-Dalradian 106-3 13.77 13.96 3.5 Pre-Dalradian 106-3 16.03 16.55 2.1 Pre-Dalradian 106-3 36.96 37.03 30.7 Pre-Dalradian 106-30 28.55 29.78 5.6 Pre-Dalradian 106-30 30.78 31.78 2.5 Pre-Dalradian 106-30 35.35 36.08 14.9 Pre-Dalradian 106-31 18.08 18.88 27.0 Pre-Dalradian 106-31 37.75 38.62 17.9 Pre-Dalradian 106-32 22.78 22.94 9.5 Pre-Dalradian 106-32 31.87 35.14 13.5 Pre-Dalradian 106-33 23.77 24.38 1.6 Pre-Dalradian 106-33 28.80 31.39 9.2 Pre-Dalradian 106-34 22.75 24.00 8.6 Pre-Dalradian 106-34 29.45 29.59 7.7 Pre-Dalradian 106-35 10.88 11.26 29.7 Pre-Dalradian 106-36 12.70 12.92 21.5 Pre-Dalradian 106-4 17.78 18.29 1.0 Pre-Dalradian 106-4 23.74 24.08 30.0 Pre-Dalradian 106-4 33.10 33.30 18.7 Pre-Dalradian 106-42 46.80 47.27 49.1 Pre-Dalradian 106-43 35.00 35.38 1.6 Pre-Dalradian 106-44 23.80 24.25 93.1 Pre-Dalradian 106-45 45.41 47.09 44.4 Pre-Dalradian 106-46 24.67 25.67 110.0 Pre-Dalradian 106-47 14.78 15.42 1.9 Pre-Dalradian 106-48 129.06 129.23 85.8 Pre-Dalradian 106-49 94.49 94.69 8.1 Pre-Dalradian 106-5 47.62 48.27 20.8 Pre-Dalradian 106-50 25.18 26.95 94.0 Pre-Dalradian 106-52 24.80 24.90 2.4 Pre-Dalradian 106-52 57.64 57.74 3.7 Pre-Dalradian 106-52 70.52 71.02 11.1 Pre-Dalradian 106-53 21.10 21.40 45.2 Pre-Dalradian 106-53 36.66 36.69 21.3 Pre-Dalradian 106-53 76.95 78.96 48.7 Pre-Dalradian 106-53 89.94 90.30 14.9 Pre-Dalradian 106-54 61.89 62.28 21.4 Pre-Dalradian 106-54 109.57 110.17 9.0 Pre-Dalradian 106-54 120.02 120.14 19.8 Pre-Dalradian 106-55 69.87 70.04 19.8 Pre-Dalradian 106-55 113.38 113.84 24.3 Pre-Dalradian 106-58 19.35 20.05 2.3 Pre-Dalradian

268

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-58 42.50 42.60 33.5 Pre-Dalradian 106-59 19.74 19.80 19.3 Pre-Dalradian 106-59 33.44 33.59 60.8 Pre-Dalradian 106-59 66.97 68.00 18.2 Pre-Dalradian 106-6 10.70 10.81 4.1 Pre-Dalradian 106-6 27.85 28.25 57.5 Pre-Dalradian 106-6 48.58 49.21 41.2 Pre-Dalradian 106-6 81.40 83.20 12.2 Pre-Dalradian 106-6 85.22 85.46 3.0 Pre-Dalradian 106-60 50.80 50.90 9.6 Pre-Dalradian 106-60 65.10 65.90 83.3 Pre-Dalradian 106-61 22.90 23.12 4.8 Pre-Dalradian 106-61 58.00 58.25 12.7 Pre-Dalradian 106-62 53.36 53.51 119.8 Pre-Dalradian 106-62 57.08 57.67 3.6 Pre-Dalradian 106-62 86.36 86.93 42.4 Pre-Dalradian 106-63 23.10 23.80 25.3 Pre-Dalradian 106-63 63.00 63.30 7.1 Pre-Dalradian 106-63 70.31 70.56 2.2 Pre-Dalradian 106-64 54.00 54.55 16.1 Pre-Dalradian 106-65 26.47 26.65 7.9 Pre-Dalradian 106-66 38.28 38.33 13.0 Pre-Dalradian 106-66 47.90 48.00 1.4 Pre-Dalradian 106-66 87.20 88.37 32.9 Pre-Dalradian 106-67 21.25 21.50 20.6 Pre-Dalradian 106-68 47.03 48.38 8.5 Pre-Dalradian 106-68 97.55 97.65 2.0 Pre-Dalradian 106-68 101.58 101.97 9.4 Pre-Dalradian 106-68 107.28 107.57 5.5 Pre-Dalradian 106-68 108.20 108.73 1.2 Pre-Dalradian 106-70 67.35 67.69 27.6 Pre-Dalradian 106-71 43.50 43.67 65.0 Pre-Dalradian 106-71 54.10 54.64 32.4 Pre-Dalradian 106-72 89.17 89.32 6.4 Pre-Dalradian 106-73 68.52 69.82 7.7 Pre-Dalradian 106-74 52.40 52.60 17.0 Pre-Dalradian 106-74 56.93 57.14 17.5 Pre-Dalradian 106-75 35.90 36.03 26.6 Pre-Dalradian 106-75 38.07 38.18 26.0 Pre-Dalradian 106-76 44.25 44.47 7.9 Pre-Dalradian 106-78 36.12 36.93 25.5 Pre-Dalradian 106-78 62.00 62.50 7.4 Pre-Dalradian 106-8 67.12 69.87 50.9 Pre-Dalradian 106-8 91.65 91.87 18.2 Pre-Dalradian 106-8 123.71 124.77 14.5 Pre-Dalradian 106-8 128.23 128.83 6.7 Pre-Dalradian 106-80 43.74 44.36 19.0 Pre-Dalradian 106-81 34.32 34.62 68.0 Pre-Dalradian 106-82 16.30 16.60 11.1 Pre-Dalradian 106-82 68.00 68.67 58.9 Pre-Dalradian 106-83 32.90 34.75 9.7 Pre-Dalradian 106-84 32.00 33.00 2.6 Pre-Dalradian 106-85 53.86 54.19 6.0 Pre-Dalradian

269

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-85 90.85 90.92 7.3 Pre-Dalradian 106-85 92.35 93.32 16.7 Pre-Dalradian 106-87 46.13 46.95 7.0 Pre-Dalradian 106-87 66.43 66.70 18.5 Pre-Dalradian 106-87 99.10 99.30 5.3 Pre-Dalradian 106-88 43.02 43.35 122.5 Pre-Dalradian 106-89 109.80 112.20 46.4 Pre-Dalradian 106-9 19.12 19.72 49.5 Pre-Dalradian 106-9 21.53 21.73 4.5 Pre-Dalradian 106-9 49.87 50.03 27.8 Pre-Dalradian 106-9 50.19 50.28 1.0 Pre-Dalradian 106-90 89.00 89.50 19.7 Pre-Dalradian 106-90 94.80 95.07 14.0 Pre-Dalradian 106-90 101.33 101.47 5.5 Pre-Dalradian 106-90 145.42 145.56 13.5 Pre-Dalradian 106-91 33.87 34.04 6.4 Pre-Dalradian 106-91 34.80 34.90 4.0 Pre-Dalradian 106-91 89.00 89.10 20.2 Pre-Dalradian 106-91 93.60 93.73 141.3 Pre-Dalradian 106-91 124.35 125.40 7.5 Pre-Dalradian 106-92 9.18 9.36 38.3 Pre-Dalradian 106-93 12.90 13.60 46.1 Pre-Dalradian 106-93 20.50 20.75 14.6 Pre-Dalradian 106-94 21.40 21.77 29.5 Pre-Dalradian 106-94 33.60 33.68 54.0 Pre-Dalradian 106-94 35.57 35.70 53.5 Pre-Dalradian 106-94 37.74 37.83 26.0 Pre-Dalradian 106-94 43.84 43.96 11.5 Pre-Dalradian 106-94 44.94 45.00 2.6 Pre-Dalradian 106-94 55.07 55.21 12.5 Pre-Dalradian 106-94 61.30 61.46 115.0 Pre-Dalradian 106-94 108.50 109.60 27.1 Pre-Dalradian 106-95 65.70 66.05 57.0 Pre-Dalradian 106-95 91.50 91.72 9.2 Pre-Dalradian 106-95 93.55 94.06 99.0 Pre-Dalradian 106-95 99.80 100.20 4.0 Pre-Dalradian 106-95 102.24 102.38 15.3 Pre-Dalradian 106-95 107.58 110.06 11.1 Pre-Dalradian 106-95 139.20 140.05 54.6 Pre-Dalradian 106-95 144.25 146.53 12.8 Pre-Dalradian 106-95 149.20 150.50 3.5 Pre-Dalradian 106-95 155.55 155.73 25.0 Pre-Dalradian 106-96 48.50 48.57 257.0 Pre-Dalradian 106-96 50.70 50.81 1.2 Pre-Dalradian 106-96 67.20 67.54 59.8 Pre-Dalradian 106-96 77.22 77.38 13.2 Pre-Dalradian 106-96 119.24 119.84 41.0 Pre-Dalradian 106-96 124.64 126.21 4.5 Pre-Dalradian 106-97 79.70 80.73 12.2 Pre-Dalradian 106-97 101.68 101.96 19.9 Pre-Dalradian 106-97 106.59 106.93 15.2 Pre-Dalradian 106-97 158.80 159.60 15.8 Pre-Dalradian 106-98 32.45 32.95 25.8 Pre-Dalradian

270

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 106-98 92.10 92.90 24.2 Pre-Dalradian 106-99 43.22 43.45 113.8 Pre-Dalradian 106-99 64.31 66.78 14.5 Pre-Dalradian 106-99 126.90 127.12 17.8 Pre-Dalradian 106-99 129.60 130.47 22.0 Pre-Dalradian 106-SH1 35.50 35.98 9.8 Pre-Dalradian 106-SH10 36.70 36.79 19.5 Pre-Dalradian 106-SH11 15.15 15.30 72.0 Pre-Dalradian 106-SH13 57.11 58.00 9.1 Pre-Dalradian 106-SH-14 34.32 34.68 26.7 Pre-Dalradian 106-SH2 49.20 49.30 3.8 Pre-Dalradian 03-CT-01 30.80 32.60 3.1 Old Resource 03-CT-01 87.90 89.16 5.0 Old Resource 03-CT-01 120.87 121.05 28.2 Old Resource 03-CT-01 161.70 163.25 7.3 Old Resource 03-CT-01 177.28 180.22 7.7 Old Resource 03-CT-01 186.78 188.28 5.0 Old Resource 03-CT-01 188.70 189.05 17.8 Old Resource 03-CT-02 44.80 45.00 6.3 Old Resource 03-CT-02 46.85 46.93 17.2 Old Resource 03-CT-02 52.50 53.70 12.6 Old Resource 03-CT-02 64.59 67.24 3.4 Old Resource 03-CT-02 95.28 95.57 7.7 Old Resource 03-CT-02 120.37 120.47 15.4 Old Resource 03-CT-03 48.35 48.80 12.5 Old Resource 03-CT-03 64.60 65.60 8.1 Old Resource 03-CT-03 112.32 115.15 12.4 Old Resource 03-CT-03 121.36 122.56 9.8 Old Resource 03-CT-04 62.30 62.95 12.2 Old Resource 03-CT-04 89.53 90.14 7.5 Old Resource 03-CT-04 102.65 102.83 11.4 Old Resource 03-CT-04 119.51 119.75 1.0 Old Resource 03-CT-04 167.46 168.72 3.6 Old Resource 03-CT-05 87.01 88.85 5.0 Old Resource 03-CT-05 101.47 101.53 20.4 Old Resource 03-CT-05 106.97 108.44 27.0 Old Resource 03-CT-05 113.68 114.12 5.8 Old Resource 03-CT-05 123.75 123.98 4.3 Old Resource 03-CT-05 127.20 127.90 1.6 Old Resource 03-CT-05 146.07 147.74 22.0 Old Resource 03-CT-06 24.90 25.00 4.7 Old Resource 03-CT-06 39.60 40.37 16.7 Old Resource 03-CT-06 146.41 147.80 9.2 Old Resource 03-CT-06 172.59 172.99 1.5 Old Resource 03-CT-07 78.73 81.74 6.1 Old Resource 03-CT-07 93.70 94.10 4.1 Old Resource 03-CT-07 103.62 104.40 12.3 Old Resource 03-CT-08 21.13 21.84 9.3 Old Resource 03-CT-08 40.95 41.28 1.4 Old Resource 03-CT-08 123.24 126.88 2.5 Old Resource 03-CT-08 186.69 187.86 9.7 Old Resource 03-CT-09 42.92 43.25 3.0 Old Resource 03-CT-09 46.06 47.51 2.3 Old Resource

271

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 03-CT-09 157.07 158.05 3.1 Old Resource 03-CT-10 61.50 62.77 37.7 Old Resource 03-CT-10 87.75 87.90 6.1 Old Resource 03-CT-10 91.18 91.53 3.6 Old Resource 03-CT-10 115.72 117.14 8.4 Old Resource 03-CT-11 51.59 52.42 5.5 Old Resource 03-CT-11 84.46 85.53 5.1 Old Resource 03-CT-11 99.78 100.95 23.1 Old Resource 03-CT-11 164.83 165.17 2.2 Old Resource 03-CT-11 167.08 168.44 6.5 Old Resource 04-CT-12 20.37 22.29 57.2 Old Resource 04-CT-12 59.59 60.67 13.7 Old Resource 04-CT-12 78.27 79.00 6.6 Old Resource 04-CT-12 144.30 145.63 11.4 Old Resource 04-CT-13 39.59 43.02 68.1 Old Resource 04-CT-14 42.94 47.00 42.7 Old Resource 04-CT-15 10.65 11.70 46.6 Old Resource 04-CT-15 19.30 22.62 4.3 Old Resource 04-CT-15 47.05 49.41 15.6 Old Resource 04-CT-15 52.35 53.21 2.4 Old Resource 04-CT-15 80.23 81.25 12.0 Old Resource 04-CT-15 102.82 103.82 9.3 Old Resource 04-CT-15 114.70 114.84 5.3 Old Resource 04-CT-15 145.88 150.40 8.6 Old Resource 04-CT-16 74.75 77.90 8.3 Old Resource 04-CT-16 126.65 128.01 9.2 Old Resource 04-CT-16 137.45 138.32 9.3 Old Resource 04-CT-17 13.63 14.78 2.3 Old Resource 04-CT-17 38.06 38.68 5.6 Old Resource 04-CT-17 42.37 42.78 4.2 Old Resource 04-CT-17 133.05 133.55 4.2 Old Resource 04-CT-17 203.36 204.68 14.4 Old Resource 04-CT-17 209.57 209.89 4.6 Old Resource 04-CT-17 211.24 212.15 5.8 Old Resource 04-CT-19 109.93 110.65 2.1 Old Resource 04-CT-20 61.30 62.35 19.4 Old Resource 04-CT-20 64.80 64.97 1.3 Old Resource 04-CT-20 78.01 78.38 6.3 Old Resource 04-CT-21 31.95 32.10 5.6 Old Resource 04-CT-22 30.50 33.32 52.8 Old Resource 04-CT-23 62.24 62.67 2.3 Old Resource 04-CT-23 75.90 76.52 2.3 Old Resource 04-CT-23 145.25 151.40 3.4 Old Resource 04-CT-23 170.06 170.78 5.8 Old Resource 04-CT-23 192.10 193.28 10.4 Old Resource 04-CT-23 200.89 202.03 12.5 Old Resource 04-CT-23 269.57 270.67 6.2 Old Resource 04-CT-24 49.87 50.12 12.9 Old Resource 04-CT-24 58.72 59.82 19.4 Old Resource 04-CT-24 131.02 131.37 5.6 Old Resource 04-CT-24 156.69 156.94 4.6 Old Resource 04-CT-24 235.68 236.68 2.3 Old Resource 04-CT-24 297.88 298.84 3.6 Old Resource

272

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 04-CT-24 305.00 306.00 1.9 Old Resource 04-CT-24 341.73 342.78 2.8 Old Resource 04-CT-25 40.57 41.25 9.2 Old Resource 04-CT-25 176.43 177.62 1.0 Old Resource 04-CT-25 206.17 206.42 1.1 Old Resource 04-CT-26 149.62 150.74 29.3 Old Resource 04-CT-26 181.50 182.00 8.1 Old Resource 04-CT-26 195.05 195.30 17.1 Old Resource 04-CT-26 210.18 210.43 6.9 Old Resource 04-CT-26 214.88 215.92 23.2 Old Resource 04-CT-26 262.49 262.74 9.7 Old Resource 04-CT-26 287.56 287.81 21.8 Old Resource 04-CT-26 365.92 366.12 10.4 Old Resource 04-CT-26 377.60 378.71 5.2 Old Resource 04-CT-26 407.86 408.02 6.3 Old Resource 04-CT-26 416.16 417.73 56.3 Old Resource 04-CT-26 442.51 443.31 12.8 Old Resource 05-CT-27 35.58 36.00 4.7 Old Resource 05-CT-28 84.52 85.84 17.6 Old Resource 06-CT-29 108.41 108.71 3.9 Old Resource 06-CT-29 146.26 146.35 10.3 Old Resource 06-CT-29 153.04 153.70 4.1 Old Resource 06-CT-29 159.80 161.08 65.7 Old Resource 06-CT-29 168.91 169.15 2.9 Old Resource 06-CT-29 185.73 186.98 8.4 Old Resource 06-CT-29 195.42 197.54 2.2 Old Resource 06-CT-29 213.59 214.59 6.5 Old Resource 06-CT-30 84.07 84.57 1.3 Old Resource 06-CT-30 114.61 114.71 21.6 Old Resource 06-CT-30 119.19 119.57 19.5 Old Resource 06-CT-30 124.56 124.92 1.6 Old Resource 06-CT-30 151.91 152.52 6.6 Old Resource 06-CT-30 154.16 155.54 6.5 Old Resource 06-CT-30 167.97 168.15 11.4 Old Resource 06-CT-31 23.77 25.02 7.5 Old Resource 06-CT-31 48.67 48.79 2.4 Old Resource 06-CT-31 51.32 51.57 15.7 Old Resource 06-CT-31 64.94 66.35 5.6 Old Resource 06-CT-31 72.22 72.30 11.0 Old Resource 06-CT-31 76.42 76.52 2.6 Old Resource 06-CT-31 102.48 102.79 3.5 Old Resource 06-CT-31 106.22 108.99 6.1 Old Resource 06-CT-31 121.57 122.45 5.1 Old Resource 06-CT-32 53.85 53.97 22.2 Old Resource 06-CT-32 58.78 58.88 29.1 Old Resource 06-CT-32 62.86 65.41 5.8 Old Resource 06-CT-32 80.80 81.00 4.0 Old Resource 06-CT-32 99.93 100.11 12.5 Old Resource 06-CT-32 106.88 108.12 10.6 Old Resource 06-CT-33 39.87 40.17 15.4 Old Resource 06-CT-33 58.50 58.62 3.1 Old Resource 06-CT-33 61.71 62.71 15.4 Old Resource 06-CT-33 70.53 71.19 10.9 Old Resource

273

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 06-CT-33 79.29 79.87 7.8 Old Resource 06-CT-33 84.97 85.58 1.0 Old Resource 06-CT-33 86.33 86.48 10.9 Old Resource 06-CT-34 18.07 18.22 22.5 Old Resource 06-CT-34 21.60 23.66 23.0 Old Resource 06-CT-34 28.76 28.93 28.9 Old Resource 06-CT-34 33.48 33.58 3.9 Old Resource 06-CT-34 45.88 46.27 7.1 Old Resource 06-CT-34 61.89 62.05 38.6 Old Resource 06-CT-35 20.61 20.84 4.5 Old Resource 06-CT-35 48.48 49.11 7.6 Old Resource 06-CT-35 49.91 51.16 19.6 Old Resource 06-CT-35 54.71 55.09 16.0 Old Resource 06-CT-35 102.43 103.13 8.0 Old Resource 06-CT-35 187.15 187.34 13.4 Old Resource 06-CT-35 192.75 192.82 1.8 Old Resource 06-CT-35 232.56 232.63 8.5 Old Resource 06-CT-36 87.57 87.96 5.0 Old Resource 06-CT-36 90.63 90.85 14.2 Old Resource 06-CT-36 101.69 101.85 8.6 Old Resource 06-CT-36 107.54 107.93 6.8 Old Resource 06-CT-36 141.18 141.50 4.3 Old Resource 06-CT-36 148.27 148.91 4.7 Old Resource 06-CT-36 156.96 157.99 16.7 Old Resource 06-CT-36 167.35 168.35 7.6 Old Resource 06-CT-36 180.21 180.60 3.7 Old Resource 06-CT-36 199.00 199.74 7.1 Old Resource 06-CT-37 70.32 70.98 8.4 Old Resource 06-CT-37 110.26 111.22 8.2 Old Resource 06-CT-37 137.98 138.98 9.3 Old Resource 06-CT-37 172.08 173.06 4.3 Old Resource 06-CT-37 185.52 186.17 2.7 Old Resource 06-CT-37 212.80 212.87 2.9 Old Resource 06-CT-37 218.17 218.47 2.4 Old Resource 06-CT-37 233.08 234.10 14.0 Old Resource 06-CT-37 247.10 247.16 17.3 Old Resource 06-CT-37 249.90 250.00 6.0 Old Resource 06-CT-38 18.75 18.82 14.1 Old Resource 06-CT-38 48.03 48.38 3.1 Old Resource 06-CT-38 81.00 81.09 3.2 Old Resource 06-CT-38 90.44 91.86 12.8 Old Resource 06-CT-38 116.43 117.18 5.3 Old Resource 06-CT-38 120.89 121.56 7.5 Old Resource 06-CT-38 141.81 142.84 10.1 Old Resource 06-CT-38 145.03 145.74 1.6 Old Resource 06-CT-39 82.42 83.47 9.1 Old Resource 06-CT-39 87.00 87.14 4.1 Old Resource 06-CT-39 96.09 96.21 11.1 Old Resource 06-CT-39 130.83 131.82 3.6 Old Resource 06-CT-39 133.00 133.98 4.7 Old Resource 06-CT-40 17.15 17.82 1.1 Old Resource 06-CT-40 35.02 36.27 7.3 Old Resource 06-CT-40 42.47 43.48 5.5 Old Resource

274

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 06-CT-40 55.88 56.30 2.7 Old Resource 06-CT-40 63.38 64.94 1.1 Old Resource 06-CT-40 89.89 91.64 11.6 Old Resource 06-CT-40 105.12 106.07 5.3 Old Resource 06-CT-40 122.05 123.13 10.0 Old Resource 06-CT-40 164.87 165.29 6.4 Old Resource 06-CT-40 172.98 173.75 2.2 Old Resource 06-CT-41 25.10 25.30 5.0 Old Resource 06-CT-41 65.83 66.96 11.9 Old Resource 06-CT-41 87.40 88.82 2.1 Old Resource 06-CT-41 99.90 100.80 4.8 Old Resource 06-CT-41 108.62 109.64 20.4 Old Resource 06-CT-41 126.16 127.35 1.0 Old Resource 06-CT-41 145.71 146.78 3.2 Old Resource 06-CT-42 31.81 32.85 2.3 Old Resource 06-CT-42 50.78 52.36 1.6 Old Resource 06-CT-42 65.17 65.37 3.7 Old Resource 06-CT-42 77.16 77.55 1.0 Old Resource 06-CT-42 87.99 89.43 2.7 Old Resource 06-CT-42 105.12 106.07 5.3 Old Resource 06-CT-42 111.33 112.33 1.3 Old Resource 06-CT-42 123.62 123.73 6.7 Old Resource 06-CT-42 131.56 131.77 2.7 Old Resource 06-CT-42 134.04 134.85 1.2 Old Resource 06-CT-42 137.91 138.91 6.2 Old Resource 06-CT-43 54.73 56.50 5.0 Old Resource 06-CT-43 108.00 108.48 2.8 Old Resource 06-CT-43 134.45 136.07 1.4 Old Resource 06-CT-43 166.23 166.92 5.4 Old Resource 06-CT-43 196.30 197.96 9.0 Old Resource 07-CT-44 41.99 43.24 7.3 Old Resource 07-CT-44 47.02 47.47 4.2 Old Resource 07-CT-44 80.39 81.00 1.5 Old Resource 07-CT-44 101.48 101.95 5.6 Old Resource 07-CT-44 111.20 111.28 23.7 Old Resource 07-CT-45 21.23 22.25 6.3 Old Resource 07-CT-45 42.17 42.27 4.1 Old Resource 07-CT-45 48.84 50.19 6.9 Old Resource 07-CT-45 78.65 79.16 8.1 Old Resource 07-CT-45 81.09 81.21 4.6 Old Resource 07-CT-45 167.95 168.69 4.8 Old Resource 07-CT-45 175.27 177.59 20.7 Old Resource 07-CT-45 180.55 181.37 4.0 Old Resource 07-CT-45 189.32 189.51 8.8 Old Resource 07-CT-46 37.90 38.95 1.7 Old Resource 07-CT-46 69.70 70.35 1.2 Old Resource 07-CT-47 77.40 77.53 24.6 Old Resource 07-CT-47 78.93 79.67 5.3 Old Resource 07-CT-47 92.49 92.68 7.7 Old Resource 07-CT-47 104.98 105.83 2.9 Old Resource 07-CT-47 130.95 132.01 5.5 Old Resource 07-CT-47 146.66 146.86 6.5 Old Resource 07-CT-47 178.00 179.86 6.3 Old Resource

275

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 07-CT-48 80.09 81.09 5.3 Old Resource 07-CT-48 107.20 107.33 8.5 Old Resource 07-CT-48 125.80 126.00 1.7 Old Resource 07-CT-48 126.22 127.46 3.5 Old Resource 07-CT-48 132.29 132.67 5.9 Old Resource 07-CT-49 29.60 30.19 3.7 Old Resource 07-CT-49 61.36 62.40 8.3 Old Resource 07-CT-49 76.63 78.00 13.8 Old Resource 07-CT-49 82.45 83.45 11.6 Old Resource 07-CT-49 89.41 90.11 2.5 Old Resource 07-CT-49 94.04 96.00 8.0 Old Resource 07-CT-49 97.49 97.63 1.8 Old Resource 07-CT-49 112.28 112.35 12.6 Old Resource 07-CT-49 114.06 114.22 19.8 Old Resource 07-CT-49 149.90 150.00 7.1 Old Resource 07-CT-49 173.52 174.88 5.3 Old Resource 07-CT-49 272.40 273.20 1.6 Old Resource 07-CT-50 68.30 68.59 5.2 Old Resource 07-CT-50 150.14 151.19 41.6 Old Resource 07-CT-50 168.63 168.67 18.4 Old Resource 07-CT-50 197.30 197.35 34.1 Old Resource 07-CT-50 212.04 212.30 13.9 Old Resource 07-CT-50 226.56 227.05 2.5 Old Resource 07-CT-50 234.00 235.07 2.3 Old Resource 07-CT-50 287.12 288.16 7.7 Old Resource 07-CT-50 303.00 303.75 3.5 Old Resource 07-CT-50 310.67 311.42 1.9 Old Resource 07-CT-50 338.42 339.14 6.4 Old Resource 07-CT-50 345.55 345.74 9.0 Old Resource 07-CT-50 346.52 348.09 4.8 Old Resource 07-CT-53 57.25 58.00 12.0 Old Resource 07-CT-53 78.63 78.86 41.9 Old Resource 07-CT-53 107.78 108.03 3.4 Old Resource 07-CT-53 116.98 117.72 7.5 Old Resource 07-CT-53 141.49 142.74 3.1 Old Resource 07-CT-53 165.64 166.10 39.3 Old Resource 07-CT-53 167.64 169.45 3.9 Old Resource 07-CT-53 175.58 177.11 7.0 Old Resource 07-CT-53 276.49 276.93 4.8 Old Resource 07-CT-53 370.04 370.73 20.9 Old Resource 07-CT-53 375.85 376.65 7.2 Old Resource 07-CT-53 395.94 396.19 8.2 Old Resource 07-CT-53 430.07 431.12 66.9 Old Resource 07-CT-53 444.79 445.79 27.4 Old Resource 08-CT-54a 185.75 188.00 6.2 Old Resource 08-CT-54a 198.73 200.08 6.1 Old Resource 08-CT-54a 218.46 218.86 14.8 Old Resource 08-CT-54a 449.78 451.98 3.2 Old Resource 08-CT-54a 465.52 466.32 5.7 Old Resource 08-CT-54a 538.63 539.63 2.1 Old Resource 08-CT-54a 542.13 543.13 14.6 Old Resource 08-CT-55 17.68 17.93 1.7 Old Resource 08-CT-55 22.37 22.57 2.3 Old Resource

276

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 08-CT-55 78.55 80.30 3.5 Old Resource 08-CT-55 81.55 83.55 4.3 Old Resource 08-CT-55 137.21 137.90 4.3 Old Resource 08-CT-55 175.88 176.03 17.4 Old Resource 08-CT-55 275.36 276.29 7.4 Old Resource 08-CT-55 350.15 351.25 4.7 Old Resource 08-CT-55 426.00 426.15 13.8 Old Resource 08-CT-55 491.46 491.56 8.4 Old Resource 08-CT-55 519.25 519.37 7.7 Old Resource 08-CT-55 532.22 532.47 3.4 Old Resource 08-CT-55 540.01 540.26 3.5 Old Resource 08-CT-55 565.47 566.67 4.0 Old Resource 08-CT-55 615.67 618.92 17.2 Old Resource 08-CT-55 620.67 621.17 2.6 Old Resource 08-CT-56a 203.17 203.42 1.5 Old Resource 08-CT-56a 247.55 248.55 6.2 Old Resource 08-CT-56a 274.27 275.02 25.8 Old Resource 08-CT-56a 281.60 281.75 1.2 Old Resource 08-CT-56a 444.19 444.81 6.6 Old Resource 08-CT-56a 522.50 524.00 7.8 Old Resource 08-CT-56a 531.58 531.87 8.6 Old Resource 08-CT-56a 632.75 633.30 7.8 Old Resource 08-CT-56a 645.38 646.38 1.1 Old Resource 08-CT-57 349.25 349.60 2.0 Old Resource 08-CT-57 372.50 372.90 1.8 Old Resource 08-CT-57 443.63 444.34 2.4 Old Resource 08-CT-57 447.95 448.45 6.0 Old Resource 08-CT-57 542.20 542.70 1.2 Old Resource 08-CT-57 570.07 570.32 5.1 Old Resource 08-CT-57 630.06 630.31 20.2 Old Resource 10-CT-58 18.62 19.22 16.3 Dalradian New Resource 10-CT-58 24.25 24.50 53.2 Dalradian New Resource 10-CT-58 38.36 38.61 20.7 Dalradian New Resource 10-CT-58 53.00 53.25 1.0 Dalradian New Resource 10-CT-58 163.94 164.69 2.9 Dalradian New Resource 10-CT-58 189.70 190.70 3.2 Dalradian New Resource 10-CT-58 225.75 226.75 20.0 Dalradian New Resource 10-CT-58 228.00 228.50 1.4 Dalradian New Resource 10-CT-59 22.10 22.40 38.9 Dalradian New Resource 10-CT-59 24.85 25.05 98.9 Dalradian New Resource 10-CT-59 34.28 34.53 1.6 Dalradian New Resource 10-CT-59 107.54 107.94 26.6 Dalradian New Resource 10-CT-59 113.22 113.47 17.4 Dalradian New Resource 10-CT-59 168.15 168.42 25.5 Dalradian New Resource 10-CT-59 178.39 180.06 5.6 Dalradian New Resource 10-CT-60 26.00 28.30 9.3 Dalradian New Resource 10-CT-60 58.66 58.91 44.9 Dalradian New Resource 10-CT-60 60.66 61.16 7.4 Dalradian New Resource 10-CT-60 111.70 112.50 26.4 Dalradian New Resource 10-CT-60 156.36 157.36 6.9 Dalradian New Resource 10-CT-60 199.32 199.82 3.7 Dalradian New Resource 10-CT-60 201.08 201.78 6.9 Dalradian New Resource 10-CT-61 51.43 52.43 6.7 Dalradian New Resource

277

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 10-CT-61 109.22 109.47 6.4 Dalradian New Resource 10-CT-61 135.82 135.92 10.4 Dalradian New Resource 10-CT-62 169.32 169.44 4.9 Dalradian New Resource 10-CT-63 45.56 50.43 5.5 Dalradian New Resource 10-CT-63 52.61 53.36 20.2 Dalradian New Resource 10-CT-63 104.20 104.32 10.3 Dalradian New Resource 10-CT-63 118.25 118.50 1.9 Dalradian New Resource 10-CT-63 123.94 124.04 4.3 Dalradian New Resource 10-CT-63 126.87 128.12 16.6 Dalradian New Resource 10-CT-63 133.37 134.12 9.5 Dalradian New Resource 10-CT-63 151.63 151.83 19.0 Dalradian New Resource 10-CT-63 172.80 173.15 4.4 Dalradian New Resource 10-CT-63 209.80 210.40 28.0 Dalradian New Resource 10-CT-63 227.70 227.83 5.7 Dalradian New Resource 10-CT-63 265.12 265.39 5.0 Dalradian New Resource 10-CT-63 284.58 284.88 17.6 Dalradian New Resource 10-CT-63 285.77 286.02 3.5 Dalradian New Resource 10-CT-63 309.29 310.04 1.4 Dalradian New Resource 10-CT-64 31.30 31.40 13.5 Dalradian New Resource 10-CT-64 42.13 42.33 12.7 Dalradian New Resource 10-CT-64 55.75 56.55 2.3 Dalradian New Resource 10-CT-64 125.77 126.07 73.7 Dalradian New Resource 10-CT-64 222.70 222.95 75.2 Dalradian New Resource 10-CT-65 6.55 7.55 4.1 Dalradian New Resource 10-CT-65 170.15 170.90 6.8 Dalradian New Resource 10-CT-66 25.75 27.00 59.9 Dalradian New Resource 10-CT-66 44.76 44.96 6.0 Dalradian New Resource 10-CT-66 113.42 113.87 14.4 Dalradian New Resource 10-CT-66 118.65 118.80 33.3 Dalradian New Resource 10-CT-66 175.72 175.92 29.6 Dalradian New Resource 10-CT-66 177.22 177.52 16.8 Dalradian New Resource 10-CT-67 17.05 17.25 3.1 Dalradian New Resource 10-CT-67 36.80 36.98 14.0 Dalradian New Resource 10-CT-67 57.58 58.23 14.2 Dalradian New Resource 10-CT-67 76.06 76.16 20.1 Dalradian New Resource 10-CT-67 118.43 119.23 29.8 Dalradian New Resource 10-CT-67 197.40 199.19 27.1 Dalradian New Resource 10-CT-68 12.35 12.45 14.2 Dalradian New Resource 10-CT-68 35.40 35.52 2.0 Dalradian New Resource 10-CT-68 40.90 41.50 18.7 Dalradian New Resource 10-CT-68 42.75 42.90 11.2 Dalradian New Resource 10-CT-68 71.02 71.22 8.2 Dalradian New Resource 10-CT-69 61.55 62.30 9.7 Dalradian New Resource 10-CT-70 36.50 36.60 20.9 Dalradian New Resource 10-CT-70 65.93 66.53 17.8 Dalradian New Resource 10-CT-70 124.75 125.57 13.8 Dalradian New Resource 10-CT-70 164.00 164.30 19.9 Dalradian New Resource 10-CT-70 220.60 221.00 27.8 Dalradian New Resource 10-CT-71 12.47 12.58 5.1 Dalradian New Resource 10-CT-71 51.34 51.47 18.6 Dalradian New Resource 10-CT-71 83.16 83.31 48.9 Dalradian New Resource 10-CT-71 96.77 97.77 4.4 Dalradian New Resource 10-CT-71 157.35 157.75 4.2 Dalradian New Resource

278

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 10-CT-71 247.10 247.50 17.6 Dalradian New Resource 10-CT-73 124.20 126.35 37.4 Dalradian New Resource 10-CT-73 187.20 188.58 44.5 Dalradian New Resource 10-CT-73 251.57 252.06 12.1 Dalradian New Resource 10-CT-74 16.05 16.35 72.3 Dalradian New Resource 10-CT-74 189.12 189.37 110.7 Dalradian New Resource 10-CT-74 251.47 251.97 31.9 Dalradian New Resource 10-CT-74 344.00 344.75 21.4 Dalradian New Resource 10-CT-74 365.07 365.63 53.1 Dalradian New Resource 10-CT-75 104.50 105.49 12.9 Dalradian New Resource 10-CT-75 117.52 117.64 9.4 Dalradian New Resource 10-CT-75 155.23 156.52 17.3 Dalradian New Resource 10-CT-75 167.71 167.78 27.0 Dalradian New Resource 10-CT-75 220.36 220.81 41.0 Dalradian New Resource 10-CT-75 239.00 239.52 28.4 Dalradian New Resource 10-CT-76 189.29 189.64 1.0 Dalradian New Resource 10-CT-76 204.20 204.65 9.3 Dalradian New Resource 10-CT-76 260.48 261.20 19.9 Dalradian New Resource 10-CT-76 264.44 264.74 6.8 Dalradian New Resource 10-CT-76 327.50 327.67 1.1 Dalradian New Resource 10-CT-76 330.22 330.32 9.9 Dalradian New Resource 10-CT-76 369.95 370.05 8.7 Dalradian New Resource 10-CT-76 415.10 416.62 5.2 Dalradian New Resource 10-CT-76 454.50 455.55 4.6 Dalradian New Resource 10-CT-76 480.07 480.67 8.1 Dalradian New Resource 10-CT-76 486.43 486.56 182.6 Dalradian New Resource 10-CT-76 497.33 497.48 12.0 Dalradian New Resource 10-CT-76 506.97 507.07 2.3 Dalradian New Resource 10-CT-76 539.40 540.46 4.1 Dalradian New Resource 10-CT-76 562.18 562.24 34.6 Dalradian New Resource 10-CT-76 562.96 563.45 6.6 Dalradian New Resource 10-CT-76 583.80 583.92 10.9 Dalradian New Resource 10-CT-76 590.11 590.19 4.4 Dalradian New Resource 10-CT-76 599.35 599.77 1.9 Dalradian New Resource 10-CT-76 644.99 645.54 12.4 Dalradian New Resource 10-CT-76 674.30 677.03 15.3 Dalradian New Resource 10-CT-77 156.27 156.37 12.0 Dalradian New Resource 10-CT-77 186.68 186.82 4.4 Dalradian New Resource 10-CT-77 242.72 242.88 2.9 Dalradian New Resource 10-CT-77 247.00 247.13 3.6 Dalradian New Resource 10-CT-77 257.94 258.07 7.9 Dalradian New Resource 10-CT-77 265.41 266.16 10.3 Dalradian New Resource 10-CT-77 269.41 269.60 32.0 Dalradian New Resource 10-CT-77 291.70 291.80 14.4 Dalradian New Resource 10-CT-77 309.10 310.35 6.1 Dalradian New Resource 11-CT-77a 235.73 235.81 16.3 Dalradian New Resource 11-CT-77a 243.91 244.33 4.2 Dalradian New Resource 11-CT-77a 262.59 263.35 4.9 Dalradian New Resource 11-CT-77a 266.53 267.27 1.3 Dalradian New Resource 11-CT-77a 288.58 288.83 4.1 Dalradian New Resource 11-CT-77a 306.04 308.00 3.7 Dalradian New Resource 11-CT-77a 366.28 366.35 13.1 Dalradian New Resource 11-CT-77a 375.18 375.24 8.0 Dalradian New Resource

279

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-77a 384.68 384.74 24.5 Dalradian New Resource 11-CT-77a 416.80 416.87 25.1 Dalradian New Resource 11-CT-77a 438.55 438.63 10.0 Dalradian New Resource 11-CT-77a 448.87 448.93 16.6 Dalradian New Resource 11-CT-78 26.90 27.95 10.4 Dalradian New Resource 11-CT-78 74.36 75.51 18.0 Dalradian New Resource 11-CT-78 181.19 181.29 195.2 Dalradian New Resource 11-CT-78 186.25 186.51 6.6 Dalradian New Resource 11-CT-78 234.90 235.09 27.2 Dalradian New Resource 11-CT-78 238.35 238.50 94.7 Dalradian New Resource 11-CT-78 242.61 243.46 63.4 Dalradian New Resource 11-CT-78 303.22 303.90 81.1 Dalradian New Resource 11-CT-78 306.88 307.96 6.5 Dalradian New Resource 11-CT-78 367.92 368.37 3.8 Dalradian New Resource 11-CT-78 382.53 382.86 244.4 Dalradian New Resource 11-CT-78 471.94 472.37 4.5 Dalradian New Resource 11-CT-78 482.30 483.48 18.8 Dalradian New Resource 11-CT-78 509.45 509.90 3.4 Dalradian New Resource 11-CT-78 528.00 528.12 15.7 Dalradian New Resource 11-CT-79 27.72 28.00 67.8 Dalradian New Resource 11-CT-79 39.19 40.28 21.8 Dalradian New Resource 11-CT-79 77.73 78.16 2.7 Dalradian New Resource 11-CT-79 84.09 84.47 49.0 Dalradian New Resource 11-CT-79 98.55 98.68 4.8 Dalradian New Resource 11-CT-79 103.25 103.64 7.3 Dalradian New Resource 11-CT-79 192.82 194.30 15.9 Dalradian New Resource 11-CT-80 17.71 17.92 5.1 Dalradian New Resource 11-CT-80 175.38 175.50 8.4 Dalradian New Resource 11-CT-80 254.80 255.10 52.6 Dalradian New Resource 11-CT-80 259.09 259.20 15.9 Dalradian New Resource 11-CT-80 314.82 314.92 3.1 Dalradian New Resource 11-CT-80 503.18 503.68 10.2 Dalradian New Resource 11-CT-80 551.91 552.89 10.8 Dalradian New Resource 11-CT-80 590.25 595.29 7.6 Dalradian New Resource 11-CT-80 616.04 616.52 1.4 Dalradian New Resource 11-CT-80 650.05 650.78 2.7 Dalradian New Resource 11-CT-80 652.60 652.75 4.0 Dalradian New Resource 11-CT-80 661.63 662.06 7.1 Dalradian New Resource 11-CT-80 679.55 679.96 4.2 Dalradian New Resource 11-CT-80 719.39 719.93 2.5 Dalradian New Resource 11-CT-80 721.09 721.45 2.6 Dalradian New Resource 11-CT-80 727.05 727.37 12.6 Dalradian New Resource 11-CT-80 741.56 741.81 3.6 Dalradian New Resource 11-CT-80 755.61 755.76 4.6 Dalradian New Resource 11-CT-80 756.89 757.16 2.7 Dalradian New Resource 11-CT-80 776.60 777.18 2.7 Dalradian New Resource 11-CT-81 27.30 27.50 8.6 Dalradian New Resource 11-CT-81 31.10 31.54 10.9 Dalradian New Resource 11-CT-82 105.19 105.27 58.9 Dalradian New Resource 11-CT-82a 167.66 167.73 35.2 Dalradian New Resource 11-CT-82a 283.35 286.93 9.5 Dalradian New Resource 11-CT-82a 305.80 306.15 4.0 Dalradian New Resource 11-CT-82a 359.03 359.40 7.9 Dalradian New Resource

280

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-82a 370.10 370.67 2.3 Dalradian New Resource 11-CT-82a 384.46 385.83 4.7 Dalradian New Resource 11-CT-82a 461.79 461.87 33.1 Dalradian New Resource 11-CT-82a 528.62 528.73 27.7 Dalradian New Resource 11-CT-82a 552.42 552.80 3.2 Dalradian New Resource 11-CT-83 60.12 60.41 17.9 Dalradian New Resource 11-CT-83 105.05 105.86 34.8 Dalradian New Resource 11-CT-83 107.75 107.85 4.9 Dalradian New Resource 11-CT-83 141.86 142.03 5.1 Dalradian New Resource 11-CT-83 171.00 171.35 5.3 Dalradian New Resource 11-CT-83 238.12 238.67 34.6 Dalradian New Resource 11-CT-84 50.78 51.12 13.5 Dalradian New Resource 11-CT-84 85.60 85.71 5.3 Dalradian New Resource 11-CT-84 114.40 115.00 55.4 Dalradian New Resource 11-CT-84 221.66 222.17 2.8 Dalradian New Resource 11-CT-84 257.67 257.73 15.0 Dalradian New Resource 11-CT-85 126.04 126.49 28.3 Dalradian New Resource 11-CT-85 226.88 229.42 20.6 Dalradian New Resource 11-CT-85 333.60 334.13 57.2 Dalradian New Resource 11-CT-85 358.48 358.86 76.8 Dalradian New Resource 11-CT-85 454.78 455.00 5.5 Dalradian New Resource 11-CT-85 504.79 505.90 26.8 Dalradian New Resource 11-CT-85 637.61 637.96 21.4 Dalradian New Resource 11-CT-85 639.25 639.73 1.6 Dalradian New Resource 11-CT-85 645.88 646.54 5.0 Dalradian New Resource 11-CT-85 690.01 690.17 114.6 Dalradian New Resource 11-CT-86a 23.30 23.68 4.8 Dalradian New Resource 11-CT-86a 247.29 248.25 2.3 Dalradian New Resource 11-CT-86a 330.14 330.68 1.7 Dalradian New Resource 11-CT-86a 351.97 352.12 33.6 Dalradian New Resource 11-CT-86a 438.58 439.41 6.9 Dalradian New Resource 11-CT-86a 483.98 484.18 16.5 Dalradian New Resource 11-CT-86a 488.52 488.85 19.5 Dalradian New Resource 11-CT-86a 501.89 502.40 8.9 Dalradian New Resource 11-CT-86a 507.04 507.54 24.2 Dalradian New Resource 11-CT-86a 542.64 543.22 14.3 Dalradian New Resource 11-CT-86a 547.01 547.24 19.8 Dalradian New Resource 11-CT-86a 556.32 556.53 51.8 Dalradian New Resource 11-CT-86a 562.78 562.97 70.1 Dalradian New Resource 11-CT-86a 612.05 612.21 82.6 Dalradian New Resource 11-CT-86a 666.00 666.12 11.4 Dalradian New Resource 11-CT-86a 667.31 667.38 36.2 Dalradian New Resource 11-CT-86a 735.09 735.33 33.1 Dalradian New Resource 11-CT-86a 757.21 757.92 6.4 Dalradian New Resource 11-CT-87 18.53 18.81 2.8 Dalradian New Resource 11-CT-87 72.63 73.73 3.9 Dalradian New Resource 11-CT-87 130.45 131.00 24.9 Dalradian New Resource 11-CT-87 235.44 235.59 20.1 Dalradian New Resource 11-CT-87 324.42 324.62 4.8 Dalradian New Resource 11-CT-87 339.97 340.24 4.9 Dalradian New Resource 11-CT-88 30.18 30.43 26.2 Dalradian New Resource 11-CT-88 134.86 135.40 9.9 Dalradian New Resource 11-CT-88 151.47 152.02 1.1 Dalradian New Resource

281

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-88 225.49 225.64 11.3 Dalradian New Resource 11-CT-88 265.21 265.31 3.7 Dalradian New Resource 11-CT-88 328.98 329.09 3.6 Dalradian New Resource 11-CT-88 363.95 364.05 4.5 Dalradian New Resource 11-CT-88 440.81 441.46 1.4 Dalradian New Resource 11-CT-89 30.59 32.65 116.1 Dalradian New Resource 11-CT-90 210.89 211.08 6.0 Dalradian New Resource 11-CT-90 225.10 225.29 17.6 Dalradian New Resource 11-CT-90 286.70 287.05 20.0 Dalradian New Resource 11-CT-90 410.72 411.12 204.8 Dalradian New Resource 11-CT-90 531.28 532.23 33.9 Dalradian New Resource 11-CT-90 538.86 539.36 16.1 Dalradian New Resource 11-CT-90 545.34 545.64 4.7 Dalradian New Resource 11-CT-90 548.14 548.24 1.8 Dalradian New Resource 11-CT-90 553.97 557.68 6.6 Dalradian New Resource 11-CT-90 578.83 579.05 7.6 Dalradian New Resource 11-CT-90 585.40 585.74 13.8 Dalradian New Resource 11-CT-90 599.85 600.55 4.9 Dalradian New Resource 11-CT-90 627.00 627.57 21.0 Dalradian New Resource 11-CT-90 649.35 650.12 2.3 Dalradian New Resource 11-CT-90 663.48 663.58 14.4 Dalradian New Resource 11-CT-90 668.88 669.33 2.5 Dalradian New Resource 11-CT-90 671.29 671.67 3.9 Dalradian New Resource 11-CT-90 795.33 795.63 2.0 Dalradian New Resource 11-CT-90 804.81 804.96 3.6 Dalradian New Resource 11-CT-92 19.20 19.56 13.8 Dalradian New Resource 11-CT-92 83.66 83.84 3.7 Dalradian New Resource 11-CT-92 108.42 108.52 9.4 Dalradian New Resource 11-CT-92 198.04 198.84 3.1 Dalradian New Resource 11-CT-92 295.31 295.50 4.2 Dalradian New Resource 11-CT-94 150.47 150.68 11.1 Dalradian New Resource 11-CT-94 196.26 196.35 100.5 Dalradian New Resource 11-CT-94 265.25 265.47 11.8 Dalradian New Resource 11-CT-94 340.42 348.76 5.8 Dalradian New Resource 11-CT-94 375.94 377.31 1.4 Dalradian New Resource 11-CT-94 405.73 405.86 3.4 Dalradian New Resource 11-CT-94 407.81 407.89 52.5 Dalradian New Resource 11-CT-94 440.56 441.32 3.6 Dalradian New Resource 11-CT-94 460.33 460.89 5.6 Dalradian New Resource 11-CT-94 549.49 549.70 3.1 Dalradian New Resource 11-CT-94 592.20 592.32 64.8 Dalradian New Resource 11-CT-95 143.05 143.40 21.5 Dalradian New Resource 11-CT-95 188.70 188.80 5.0 Dalradian New Resource 11-CT-95 328.05 328.28 2.5 Dalradian New Resource 11-CT-95 378.13 378.33 1.7 Dalradian New Resource 11-CT-95 492.42 492.47 25.0 Dalradian New Resource 11-CT-96 24.08 24.18 8.4 Dalradian New Resource 11-CT-96 97.30 97.41 7.1 Dalradian New Resource 11-CT-96 137.13 137.25 2.8 Dalradian New Resource 11-CT-96 164.50 165.28 5.6 Dalradian New Resource 11-CT-96 324.19 324.30 16.2 Dalradian New Resource 11-CT-96 414.24 414.89 2.5 Dalradian New Resource 11-CT-96 434.50 436.00 2.8 Dalradian New Resource

282

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-97b 424.94 425.06 14.6 Dalradian New Resource 11-CT-97b 498.58 499.11 2.6 Dalradian New Resource 11-CT-97b 537.01 538.35 9.4 Dalradian New Resource 11-CT-97b 554.36 556.86 13.6 Dalradian New Resource 11-CT-97b 601.40 602.14 6.7 Dalradian New Resource 11-CT-97b 604.65 605.39 5.5 Dalradian New Resource 11-CT-97b 615.43 616.05 5.9 Dalradian New Resource 11-CT-97b 624.25 624.56 1.2 Dalradian New Resource 11-CT-97b 636.38 637.38 2.2 Dalradian New Resource 11-CT-97b 652.67 652.77 2.0 Dalradian New Resource 11-CT-97b 658.50 659.41 1.1 Dalradian New Resource 11-CT-97b 663.88 665.48 1.3 Dalradian New Resource 11-CT-97b 666.61 666.74 3.2 Dalradian New Resource 11-CT-97b 714.84 715.41 7.6 Dalradian New Resource 11-CT-97b 785.94 786.60 1.7 Dalradian New Resource 11-CT-97b 839.07 839.49 1.1 Dalradian New Resource 11-CT-98 238.44 239.14 2.4 Dalradian New Resource 11-CT-98 280.18 281.20 14.1 Dalradian New Resource 11-CT-98 329.83 330.00 6.2 Dalradian New Resource 11-CT-98 399.76 400.04 1.5 Dalradian New Resource 11-CT-98 408.23 408.26 23.3 Dalradian New Resource 11-CT-98 432.00 432.85 1.1 Dalradian New Resource 11-CT-98 463.30 463.40 5.6 Dalradian New Resource 11-CT-98 523.07 523.34 19.7 Dalradian New Resource 11-CT-98 531.39 532.57 6.7 Dalradian New Resource 11-CT-98 580.90 582.00 1.0 Dalradian New Resource 11-CT-98 647.80 648.53 2.8 Dalradian New Resource 11-CT-98 656.42 656.53 2.0 Dalradian New Resource 11-CT-98 751.14 752.48 10.7 Dalradian New Resource 11-CT-98 842.09 842.37 6.9 Dalradian New Resource 11-CT-98 844.67 846.31 6.1 Dalradian New Resource 11-CT-98 969.00 970.42 4.7 Dalradian New Resource 11-CT-99 467.11 468.49 5.3 Dalradian New Resource 11-CT-99 471.14 474.34 5.3 Dalradian New Resource 11-CT-99 616.92 617.02 11.5 Dalradian New Resource 11-CT-99 686.41 686.61 41.0 Dalradian New Resource 11-CT-99 702.29 702.39 8.3 Dalradian New Resource 11-CT-99 788.00 788.58 16.8 Dalradian New Resource 11-CT-99 800.26 801.36 5.0 Dalradian New Resource 11-CT-100 37.18 37.95 13.1 Dalradian New Resource 11-CT-100 142.37 142.49 35.4 Dalradian New Resource 11-CT-100 150.25 151.01 11.2 Dalradian New Resource 11-CT-100 222.62 222.65 15.9 Dalradian New Resource 11-CT-100 247.88 248.22 5.8 Dalradian New Resource 11-CT-100 287.36 287.72 2.5 Dalradian New Resource 11-CT-100 290.54 291.69 1.3 Dalradian New Resource 11-CT-100 355.68 356.14 13.0 Dalradian New Resource 11-CT-100 376.00 376.13 4.0 Dalradian New Resource 11-CT-102 145.36 145.90 1.3 Dalradian New Resource 11-CT-102 192.41 193.00 2.4 Dalradian New Resource 11-CT-102 228.30 229.10 5.1 Dalradian New Resource 11-CT-103 44.47 44.57 2.2 Dalradian New Resource 11-CT-103 107.46 107.50 1.0 Dalradian New Resource

283

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-103 121.34 121.39 8.1 Dalradian New Resource 11-CT-103 182.55 182.74 1.0 Dalradian New Resource 11-CT-103 268.35 268.67 1.0 Dalradian New Resource 11-CT-103 275.45 275.60 2.4 Dalradian New Resource 11-CT-103 329.70 329.74 7.8 Dalradian New Resource 11-CT-103 431.75 432.30 1.8 Dalradian New Resource 11-CT-103 473.37 475.03 3.5 Dalradian New Resource 11-CT-103 591.15 592.45 4.4 Dalradian New Resource 11-CT-103 603.81 604.65 3.6 Dalradian New Resource 11-CT-103 617.00 617.65 6.7 Dalradian New Resource

284

31.0 APPE NDI X 3 D A L R A DI A N R ESO UR C ES DRI L L H O L E IN T E RSE C T I O NS, POST 2011 M IN E R A L R ESO UR C E EST I M A T E

285

Hole_ID F rom To Au G rade Source/End Use (m) (m) (g/t) 11-CT-101a 124.78 124.88 2.18 Dalradian Updated Info 11-CT-101a 141.21 142.16 1.89 Dalradian Updated Info 11-CT-101a 346.40 346.84 5.14 Dalradian Updated Info 11-CT-101a 520.25 520.85 3.61 Dalradian Updated Info 11-CT-104 506.44 507.09 1.00 Dalradian Updated Info 11-CT-104 320.38 320.50 1.05 Dalradian Updated Info 11-CT-104 951.00 951.50 1.14 Dalradian Updated Info 11-CT-104 749.90 750.10 1.42 Dalradian Updated Info 11-CT-104 935.29 935.92 1.45 Dalradian Updated Info 11-CT-104 622.68 622.82 1.57 Dalradian Updated Info 11-CT-104 598.78 599.00 1.59 Dalradian Updated Info 11-CT-104 705.77 706.39 2.04 Dalradian Updated Info 11-CT-104 844.74 845.14 2.28 Dalradian Updated Info 11-CT-104 513.77 513.90 2.73 Dalradian Updated Info 11-CT-104 954.95 955.11 3.12 Dalradian Updated Info 11-CT-104 594.23 594.68 4.04 Dalradian Updated Info 11-CT-104 620.00 620.50 5.08 Dalradian Updated Info 11-CT-104 562.38 562.88 5.98 Dalradian Updated Info 11-CT-104 540.77 540.93 6.98 Dalradian Updated Info 11-CT-104 629.37 629.48 7.88 Dalradian Updated Info 11-CT-104 624.07 625.87 12.29 Dalradian Updated Info 11-CT-105 1005.16 1005.66 1.68 Dalradian Updated Info 11-CT-105 385.96 386.39 6.27 Dalradian Updated Info 11-CT-105 899.31 899.68 7.51 Dalradian Updated Info 11-CT-106 132.65 132.75 1.03 Dalradian Updated Info 11-CT-106 150.00 150.60 2.35 Dalradian Updated Info 11-CT-106 227.98 228.11 3.66 Dalradian Updated Info 11-CT-106 287.45 288.25 8.21 Dalradian Updated Info 11-CT-107 111.05 111.30 1.70 Dalradian Updated Info 11-CT-107 201.70 201.85 2.64 Dalradian Updated Info 11-CT-107 170.85 171.83 2.79 Dalradian Updated Info 11-CT-108 123.94 124.54 1.31 Dalradian Updated Info 11-CT-108 88.30 88.45 1.82 Dalradian Updated Info 11-CT-108 110.75 110.85 1.92 Dalradian Updated Info 11-CT-108 98.96 99.21 2.80 Dalradian Updated Info 11-CT-108 112.00 112.29 3.29 Dalradian Updated Info 11-CT-108 161.55 162.55 4.10 Dalradian Updated Info 11-CT-108 102.03 102.69 4.36 Dalradian Updated Info 11-CT-108 67.00 67.11 6.55 Dalradian Updated Info 11-CT-108 78.85 79.93 6.58 Dalradian Updated Info 11-CT-108 156.47 156.97 7.68 Dalradian Updated Info 11-CT-108 118.10 120.26 22.12 Dalradian Updated Info 11-CT-108 126.35 127.69 37.71 Dalradian Updated Info 11-CT-109 66.33 66.45 1.04 Dalradian Updated Info 11-CT-109 43.77 44.02 1.76 Dalradian Updated Info 11-CT-109 97.40 97.50 2.72 Dalradian Updated Info 11-CT-109 90.80 91.50 3.11 Dalradian Updated Info 11-CT-109 85.49 86.99 5.11 Dalradian Updated Info 11-CT-109 78.58 78.85 9.51 Dalradian Updated Info 11-CT-109 132.70 133.12 9.84 Dalradian Updated Info 11-CT-109 120.49 121.86 10.86 Dalradian Updated Info

286

11-CT-109 81.79 82.99 18.38 Dalradian Updated Info 11-CT-110 155.63 156.12 1.33 Dalradian Updated Info 11-CT-110 135.41 135.69 1.68 Dalradian Updated Info 11-CT-110 23.32 23.81 1.69 Dalradian Updated Info 11-CT-110 97.91 98.44 2.10 Dalradian Updated Info 11-CT-110 80.32 82.18 2.35 Dalradian Updated Info 11-CT-110 51.85 52.00 2.64 Dalradian Updated Info 11-CT-110 20.22 21.82 2.71 Dalradian Updated Info 11-CT-110 130.00 131.25 4.71 Dalradian Updated Info 11-CT-110 133.18 133.39 5.61 Dalradian Updated Info 11-CT-110 95.26 95.41 13.63 Dalradian Updated Info 11-CT-110 72.16 73.90 15.35 Dalradian Updated Info 11-CT-110 83.58 84.28 17.62 Dalradian Updated Info 11-CT-110 138.54 139.57 33.57 Dalradian Updated Info 11-CT-111 55.62 55.85 1.46 Dalradian Updated Info 11-CT-111 270.80 270.92 2.94 Dalradian Updated Info 11-CT-111 160.85 162.10 12.51 Dalradian Updated Info 11-CT-111 143.15 143.69 13.76 Dalradian Updated Info 11-CT-111 75.59 76.25 18.26 Dalradian Updated Info 11-CT-111 191.00 192.30 23.81 Dalradian Updated Info 11-CT-112 101.28 101.58 1.10 Dalradian Updated Info 11-CT-112 16.15 16.45 1.44 Dalradian Updated Info 11-CT-112 125.52 125.82 1.92 Dalradian Updated Info 11-CT-112 135.87 138.15 2.36 Dalradian Updated Info 11-CT-112 167.94 168.24 3.94 Dalradian Updated Info 11-CT-112 70.27 71.46 21.16 Dalradian Updated Info 11-CT-113 78.46 79.96 2.02 Dalradian Updated Info 11-CT-113 157.05 157.71 2.35 Dalradian Updated Info 11-CT-113 160.21 160.50 10.67 Dalradian Updated Info 11-CT-114 77.70 79.23 4.22 Dalradian Updated Info 11-CT-115 137.65 137.83 1.21 Dalradian Updated Info 11-CT-115 173.20 173.50 1.48 Dalradian Updated Info 11-CT-115 61.18 61.43 2.29 Dalradian Updated Info 11-CT-115 10.58 12.50 3.52 Dalradian Updated Info 11-CT-115 130.30 132.00 5.16 Dalradian Updated Info 11-CT-115 19.50 23.50 11.03 Dalradian Updated Info 11-CT-116 112.95 113.12 1.22 Dalradian Updated Info 11-CT-116 61.68 61.83 1.99 Dalradian Updated Info 11-CT-116 68.04 68.26 3.27 Dalradian Updated Info 11-CT-116 132.07 132.37 4.61 Dalradian Updated Info 11-CT-117 56.22 57.22 1.96 Dalradian Updated Info 11-CT-117 99.14 99.93 10.21 Dalradian Updated Info 11-CT-117 50.99 54.41 11.44 Dalradian Updated Info 11-CT-118 72.64 72.77 1.51 Dalradian Updated Info 11-CT-118 57.50 58.50 1.58 Dalradian Updated Info 11-CT-118 140.93 141.60 2.28 Dalradian Updated Info 11-CT-118 110.24 111.60 17.25 Dalradian Updated Info 11-CT-119 145.86 146.06 1.01 Dalradian Updated Info 11-CT-119 131.39 132.65 3.03 Dalradian Updated Info 11-CT-119 75.65 76.79 3.14 Dalradian Updated Info 11-CT-119 117.63 117.75 4.68 Dalradian Updated Info 11-CT-119 160.55 162.06 6.73 Dalradian Updated Info 11-CT-120 138.18 138.35 1.43 Dalradian Updated Info 11-CT-120 147.12 147.23 1.55 Dalradian Updated Info 11-CT-120 132.38 132.50 1.60 Dalradian Updated Info

287

11-CT-120 70.34 71.02 1.74 Dalradian Updated Info 11-CT-120 62.67 62.84 3.62 Dalradian Updated Info 11-CT-120 103.85 104.11 6.14 Dalradian Updated Info 11-CT-120 143.10 143.65 8.62 Dalradian Updated Info 11-CT-121 129.32 129.50 6.75 Dalradian Updated Info 11-CT-121 87.85 88.71 16.70 Dalradian Updated Info 11-CT-122 33.43 33.63 1.21 Dalradian Updated Info 11-CT-122 111.65 112.00 2.04 Dalradian Updated Info 11-CT-122 61.00 61.17 9.04 Dalradian Updated Info 11-CT-122 46.06 46.29 13.27 Dalradian Updated Info 11-CT-123 39.00 39.23 1.59 Dalradian Updated Info 11-CT-123 111.85 112.15 1.95 Dalradian Updated Info 11-CT-123 91.45 92.43 2.87 Dalradian Updated Info 11-CT-124 72.10 72.53 3.45 Dalradian Updated Info 11-CT-124 141.00 141.60 3.48 Dalradian Updated Info 11-CT-125 79.55 79.73 2.08 Dalradian Updated Info 11-CT-126 74.55 75.05 1.09 Dalradian Updated Info 11-CT-126 66.25 66.35 1.90 Dalradian Updated Info 11-CT-126 72.95 73.05 5.82 Dalradian Updated Info 11-CT-127 188.05 188.55 1.18 Dalradian Updated Info 11-CT-127 41.47 41.57 2.67 Dalradian Updated Info 11-CT-127 185.72 187.05 3.22 Dalradian Updated Info 11-CT-127 52.40 55.00 3.99 Dalradian Updated Info 11-CT-127 164.60 166.79 6.33 Dalradian Updated Info 11-CT-127 128.05 129.79 9.63 Dalradian Updated Info 11-CT-127 110.74 111.47 20.34 Dalradian Updated Info 11-CT-127 31.55 31.73 25.34 Dalradian Updated Info 11-CT-128 153.50 154.88 2.55 Dalradian Updated Info 11-CT-128 151.90 152.30 2.83 Dalradian Updated Info 11-CT-129 79.57 79.71 1.18 Dalradian Updated Info 11-CT-129 80.74 82.09 1.44 Dalradian Updated Info 11-CT-129 152.83 153.00 2.24 Dalradian Updated Info 11-CT-129 88.08 88.20 5.86 Dalradian Updated Info 11-CT-129 47.57 47.80 6.04 Dalradian Updated Info 11-CT-130 134.80 134.92 2.71 Dalradian Updated Info 11-CT-130 127.74 127.86 5.38 Dalradian Updated Info 11-CT-131 18.67 18.92 1.12 Dalradian Updated Info 11-CT-131 93.69 94.79 1.55 Dalradian Updated Info 11-CT-131 128.88 129.00 1.55 Dalradian Updated Info 11-CT-131 141.95 142.14 2.50 Dalradian Updated Info 11-CT-131 86.69 88.19 3.87 Dalradian Updated Info 11-CT-131 136.58 138.18 4.41 Dalradian Updated Info 11-CT-132 142.22 143.47 1.30 Dalradian Updated Info 11-CT-132 94.89 95.39 1.34 Dalradian Updated Info 11-CT-132 89.00 89.25 1.73 Dalradian Updated Info 11-CT-132 156.00 156.50 2.22 Dalradian Updated Info 11-CT-132 92.02 93.39 5.05 Dalradian Updated Info 11-CT-133 81.32 81.45 1.17 Dalradian Updated Info 11-CT-133 133.90 134.02 1.28 Dalradian Updated Info 11-CT-133 207.85 208.60 2.95 Dalradian Updated Info 11-CT-133 70.20 70.30 3.00 Dalradian Updated Info 11-CT-134 275.45 275.69 1.39 Dalradian Updated Info 11-CT-134 16.97 17.14 1.97 Dalradian Updated Info 11-CT-134 125.80 125.92 2.32 Dalradian Updated Info 11-CT-134 237.12 237.70 3.04 Dalradian Updated Info

288

11-CT-134 247.50 249.40 3.25 Dalradian Updated Info 11-CT-134 100.50 101.23 3.54 Dalradian Updated Info 11-CT-134 131.90 132.23 4.05 Dalradian Updated Info 11-CT-134 240.70 242.00 4.14 Dalradian Updated Info 11-CT-134 156.56 156.66 7.26 Dalradian Updated Info 11-CT-135 76.40 76.90 1.73 Dalradian Updated Info 11-CT-135 31.82 32.04 1.73 Dalradian Updated Info 11-CT-135 52.50 53.20 2.34 Dalradian Updated Info 11-CT-135 74.32 75.21 3.17 Dalradian Updated Info 11-CT-135 136.00 136.15 3.47 Dalradian Updated Info 11-CT-136 149.00 149.50 1.68 Dalradian Updated Info 11-CT-136 154.78 154.98 8.02 Dalradian Updated Info 11-CT-136 134.26 135.00 8.38 Dalradian Updated Info 11-CT-136 68.39 69.85 13.09 Dalradian Updated Info 11-CT-137 165.75 166.38 5.72 Dalradian Updated Info 11-CT-137 86.00 86.22 10.71 Dalradian Updated Info 11-CT-138 166.00 167.69 3.71 Dalradian Updated Info 11-CT-138 176.68 176.98 3.73 Dalradian Updated Info 11-CT-138 61.28 61.73 5.26 Dalradian Updated Info 11-CT-138 38.33 38.71 9.94 Dalradian Updated Info 11-CT-138 172.19 172.59 12.74 Dalradian Updated Info 11-CT-139 108.30 108.55 21.84 Dalradian Updated Info 11-CT-140 86.30 87.33 3.03 Dalradian Updated Info 11-CT-140 10.48 11.69 13.98 Dalradian Updated Info 11-CT-141 67.68 69.30 3.20 Dalradian Updated Info 11-CT-142 129.59 129.69 1.57 Dalradian Updated Info 11-CT-142 167.11 168.65 1.89 Dalradian Updated Info 11-CT-142 86.76 87.05 3.51 Dalradian Updated Info 11-CT-142 163.51 165.11 3.56 Dalradian Updated Info 11-CT-142 193.31 193.84 5.59 Dalradian Updated Info 11-CT-142 191.32 192.07 7.18 Dalradian Updated Info 11-CT-142 148.80 150.15 8.19 Dalradian Updated Info 11-CT-142 199.95 200.68 8.80 Dalradian Updated Info 11-CT-142 169.19 169.85 9.08 Dalradian Updated Info 11-CT-142 122.40 124.40 9.55 Dalradian Updated Info 11-CT-143 49.60 50.10 1.06 Dalradian Updated Info 11-CT-143 67.18 68.09 1.99 Dalradian Updated Info 11-CT-143 30.37 30.55 2.43 Dalradian Updated Info 11-CT-143 52.70 53.60 4.55 Dalradian Updated Info 11-CT-143 38.84 40.16 6.33 Dalradian Updated Info 11-CT-143 37.40 37.60 7.72 Dalradian Updated Info 11-CT-143 42.90 44.48 13.04 Dalradian Updated Info 11-CT-143 146.90 148.26 18.36 Dalradian Updated Info 11-CT-144 33.60 34.10 1.12 Dalradian Updated Info 11-CT-144 47.60 48.50 4.56 Dalradian Updated Info 11-CT-145 9.56 9.74 3.56 Dalradian Updated Info 11-CT-145 183.39 184.39 3.94 Dalradian Updated Info 11-CT-145 208.37 208.80 9.20 Dalradian Updated Info 11-CT-145 99.38 100.98 9.71 Dalradian Updated Info 11-CT-145 146.32 146.81 9.86 Dalradian Updated Info 11-CT-145 180.74 181.89 12.68 Dalradian Updated Info 11-CT-146 58.00 58.11 1.52 Dalradian Updated Info 11-CT-146 13.40 14.14 2.09 Dalradian Updated Info 11-CT-147 213.92 214.22 2.49 Dalradian Updated Info 11-CT-147 166.14 166.58 2.83 Dalradian Updated Info

289

11-CT-147 65.77 66.19 4.43 Dalradian Updated Info 11-CT-147 100.01 100.20 4.62 Dalradian Updated Info 11-CT-147 128.95 131.09 4.90 Dalradian Updated Info 11-CT-147 162.29 163.12 6.48 Dalradian Updated Info 11-CT-147 114.22 114.90 6.82 Dalradian Updated Info 11-CT-147 124.55 124.74 10.70 Dalradian Updated Info 11-CT-148 181.50 181.69 1.05 Dalradian Updated Info 11-CT-148 110.50 111.30 1.31 Dalradian Updated Info 11-CT-148 24.78 24.89 1.39 Dalradian Updated Info 11-CT-148 125.56 125.76 1.39 Dalradian Updated Info 11-CT-148 144.23 144.63 1.49 Dalradian Updated Info 11-CT-148 112.52 113.02 2.19 Dalradian Updated Info 11-CT-148 150.93 152.14 2.39 Dalradian Updated Info 11-CT-148 177.87 178.00 3.89 Dalradian Updated Info 11-CT-148 201.57 201.80 5.57 Dalradian Updated Info 11-CT-148 167.40 168.00 9.33 Dalradian Updated Info 11-CT-148 170.90 171.28 11.70 Dalradian Updated Info 11-CT-149 34.82 34.95 1.57 Dalradian Updated Info 11-CT-149 127.00 128.00 2.50 Dalradian Updated Info 11-CT-149 138.40 139.09 3.52 Dalradian Updated Info 11-CT-149 152.10 154.20 3.57 Dalradian Updated Info 11-CT-149 164.70 166.81 5.71 Dalradian Updated Info 11-CT-149 147.90 150.30 7.29 Dalradian Updated Info 11-CT-149 170.28 170.68 8.25 Dalradian Updated Info 11-CT-149 141.00 142.90 9.51 Dalradian Updated Info 11-CT-149 145.17 146.00 11.82 Dalradian Updated Info 12-CT-150 58.80 59.05 1.25 Dalradian Updated Info 12-CT-150 28.60 29.38 1.36 Dalradian Updated Info 12-CT-150 240.57 241.04 1.78 Dalradian Updated Info 12-CT-150 140.94 141.91 1.86 Dalradian Updated Info 12-CT-150 118.53 118.63 2.06 Dalradian Updated Info 12-CT-150 113.48 113.67 2.07 Dalradian Updated Info 12-CT-150 260.11 262.22 3.98 Dalradian Updated Info 12-CT-150 252.61 253.77 4.27 Dalradian Updated Info 12-CT-150 255.09 256.90 4.97 Dalradian Updated Info 12-CT-150 219.63 221.30 13.24 Dalradian Updated Info 12-CT-150 67.28 67.59 14.57 Dalradian Updated Info 12-CT-150 305.47 308.82 44.15 Dalradian Updated Info 12-CT-150 200.93 201.73 47.06 Dalradian Updated Info 12-CT-151 135.37 135.47 1.05 Dalradian Updated Info 12-CT-151 127.89 128.00 1.32 Dalradian Updated Info 12-CT-151 69.16 69.28 1.33 Dalradian Updated Info 12-CT-151 294.24 295.38 1.52 Dalradian Updated Info 12-CT-151 83.38 83.56 1.88 Dalradian Updated Info 12-CT-151 282.76 283.32 2.14 Dalradian Updated Info 12-CT-151 249.42 249.60 2.69 Dalradian Updated Info 12-CT-151 160.11 160.93 3.69 Dalradian Updated Info 12-CT-151 198.06 198.19 4.67 Dalradian Updated Info 12-CT-151 77.21 78.13 5.28 Dalradian Updated Info 12-CT-151 156.81 157.26 6.88 Dalradian Updated Info 12-CT-151 112.77 113.54 9.27 Dalradian Updated Info 12-CT-151 187.44 187.95 9.38 Dalradian Updated Info 12-CT-151 5.00 5.50 10.40 Dalradian Updated Info 12-CT-151 279.08 279.23 14.48 Dalradian Updated Info 12-CT-151 131.61 132.07 22.83 Dalradian Updated Info

290

12-CT-152 373.78 373.98 1.47 Dalradian Updated Info 12-CT-152 221.52 222.21 3.55 Dalradian Updated Info 12-CT-152 57.00 57.39 3.64 Dalradian Updated Info 12-CT-152 297.58 298.22 5.74 Dalradian Updated Info 12-CT-153 435.68 435.78 1.25 Dalradian Updated Info 12-CT-153 330.98 331.82 1.33 Dalradian Updated Info 12-CT-153 255.63 256.24 1.53 Dalradian Updated Info 12-CT-153 203.52 203.63 1.58 Dalradian Updated Info 12-CT-153 79.24 79.44 1.70 Dalradian Updated Info 12-CT-153 122.51 124.07 2.75 Dalradian Updated Info 12-CT-153 74.50 74.95 4.60 Dalradian Updated Info 12-CT-153 253.87 254.20 4.67 Dalradian Updated Info 12-CT-153 311.21 311.39 5.47 Dalradian Updated Info 12-CT-153 337.02 337.80 7.96 Dalradian Updated Info 12-CT-154 40.50 40.66 1.90 Dalradian Updated Info 12-CT-154 348.59 349.24 9.78 Dalradian Updated Info 12-CT-155 33.07 33.20 1.09 Dalradian Updated Info 12-CT-155 415.93 416.04 1.46 Dalradian Updated Info 12-CT-155 414.22 414.74 1.93 Dalradian Updated Info

291