certej updated definitive feasibility study summary technical...

Authors:- Patrick Forward February 26th 2009 Neil Liddell Tony Jackson

Certej Updated Definitive Feasibility Study

Summary Technical Report

Certej Updated Definitive Feasibility Study Summary Technical Report

Contents

3. SUMMARY ................................................................................................................................................................... 1

4. INTRODUCTION AND TERMS OF REFERENCE ................................................................................................................ 6

5. RELIANCE ON OTHER EXPERTS ...................................................................................................................................... 7

6. PROPERTY DESCRIPTION AND LOCATION ..................................................................................................................... 8

7. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY ............................................. 10

7.1. ACCESS ............................................................................................................................................................. 10

7.2. CLIMATE ........................................................................................................................................................... 10

7.3. LOCAL RESOURCES AND INFRASTRUCTURE ....................................................................................................... 10

7.4. PHYSIOGRAPHY ................................................................................................................................................ 11

8. HISTORY ..................................................................................................................................................................... 14

8.1. PRIOR TO WORLD WAR 2 .................................................................................................................................. 14

8.2. POST WORLD WAR 2 ........................................................................................................................................ 15

8.3. RECENT HISTORY .............................................................................................................................................. 16

8.3.1. HISTORICAL INFORMATION ON EXISTING DUMPS ............................................................................................ 19

9. GEOLOGICAL SETTING ................................................................................................................................................ 20

9.1. REGIONAL GEOLOGY......................................................................................................................................... 20

9.2. LOCAL GEOLOGY ............................................................................................................................................... 21

10. DEPOSIT TYPES .......................................................................................................................................................... 27

11. MINERALISATION ....................................................................................................................................................... 28

11.1. ALTERATION ..................................................................................................................................................... 30

11.2. DEPOSIT ZONATION .......................................................................................................................................... 31

12. EXPLORATION ............................................................................................................................................................ 36

12.1. EXPLORATION BY PREVIOUS OWNERS .............................................................................................................. 36

12.2. EXPLORATION BY CURRENT OWNERS ............................................................................................................... 36

12.3. EXPLORATION MANAGED BY RSG GLOBAL – 2000 TO 2003 ............................................................................... 36

12.4. INVOLVEMENT OF OTHER CONSULTANTS – NOVEMBER 2003 TO OCTOBER 2004 ............................................. 36

12.5. MANAGEMENT BY EUROPEAN GOLDFIELDS TECHNICAL TEAM ......................................................................... 37

12.5.1. EXPLORATION OF EXISTING DUMPS .............................................................................................................. 37

13. DRILLING .................................................................................................................................................................... 39

13.1. DRILLING OF DEPOSIT ....................................................................................................................................... 39

13.2. DRILLING OF EXISTING DUMPS ......................................................................................................................... 39

14. SAMPLING METHOD AND APPROACH ........................................................................................................................ 42

14.1. DUMP SAMPLING APPROACH ........................................................................................................................... 43

14.1.1. SURFACE CHANNEL ........................................................................................................................................ 43

14.1.2. REVERSE CIRCULATION .................................................................................................................................. 44

15. SAMPLE PREPARATION, ANALYSES AND SECURITY ..................................................................................................... 47

15.1. DUMP SAMPLE PREPARATION, ANALYSIS AND SECURITY ................................................................................. 49

16. DATA VERIFICATION ................................................................................................................................................... 51

16.1. ACCURACY ........................................................................................................................................................ 51

16.2. ASSAY PRECISION ............................................................................................................................................. 52

16.3. DATA VERIFICATION FOR THE DUMP EXPLORATION DATABASE. ...................................................................... 53

16.3.1. ASSAY STANDARDS ....................................................................................................................................... 53

16.3.2. DUMP FIELD DUPLICATES (RC AND CHANNEL) ............................................................................................... 54

16.3.3. LABORATORY REPLICATES AND REPEATS ....................................................................................................... 57

16.3.4. LABORATORY PREPARATION REPEATS – SPLITS (FOR 2004-2005) .................................................................. 59

16.3.5. LABORATORY BLANKS ................................................................................................................................... 61

16.3.6. INTER-LABORATORY ROUND ROBIN .............................................................................................................. 61

16.3.7. CONCLUSIONS DUMP QAQC .......................................................................................................................... 61

17. ADJACENT PROPERTIES .............................................................................................................................................. 63

18. MINERAL PROCESSING AND METALLURGICAL TESTING .............................................................................................. 64


18.1. INTRODUCTION ................................................................................................................................................ 64

18.2. MINERALOGY ................................................................................................................................................... 64

18.2.1. INTRODUCTION ............................................................................................................................................. 64

18.2.2. MINERALOGY PRE 2004 ................................................................................................................................. 65

18.2.3. DEPORTMENT OF GOLD AND SILVER IN CERTEJ FLOTATION CONCENTRATES AND TAILS ............................... 66

18.2.4. COMMENTS AND OBSERVATIONS ................................................................................................................. 72

18.3. METALLURGICAL TESTWORK ............................................................................................................................ 73

18.3.1. HISTORIC METALLURGICAL TESTWORK (PRE 2005) ........................................................................................ 73

18.3.2. 2005 TESTWORK SAMPLES ............................................................................................................................ 73

18.3.3. 2005 FLOTATION TESTWORK ......................................................................................................................... 74

18.3.4. 2008 FLOTATION TESTWORK ......................................................................................................................... 77

18.3.5. PRELIMINARY ALBION TEST WORK ................................................................................................................ 77

18.3.6. PROCESS SELECTION ...................................................................................................................................... 78

18.3.7. CONTINUOUS SCALE ALBION TESTWORK PROGRAMME ................................................................................ 78

18.3.8. DEPORTMENT OF GOLD & SILVER IN ALBION AND CIL RESIDUES, AMTEL, DECEMBER 2007 ........................... 82

18.3.9. COMMINUTION TESTWORK PROGRAMME .................................................................................................... 83

18.3.10. CYANIDE DETOXIFICATION TESTWORK ....................................................................................................... 85

18.3.11. FLOTATION TAILINGS & CIL RESIDUE PERCOLATION TESTS.......................................................................... 85

18.3.12. CERTEJ PROCESS PLANT SUMMARY ............................................................................................................ 85

18.4. PLANT DESCRIPTION ......................................................................................................................................... 86

18.4.1. OVERALL DESCRIPTION .................................................................................................................................. 86

18.4.2. CRUSHING ..................................................................................................................................................... 86

18.4.3. COMMINUTION ............................................................................................................................................. 86

18.4.4. FLOTATION CIRCUITS ..................................................................................................................................... 87

18.4.5. GRINDING & FLOTATION CIRCUIT CONTROL .................................................................................................. 87

18.4.6. FLOTATION CONCENTRATE AND TAILINGS HANDLING ................................................................................... 88

18.4.7. DESCRIPTION OF THE ALBION PROCESS PLANT .............................................................................................. 88

18.4.8. ISAMILL GRINDING ........................................................................................................................................ 88

18.4.9. ALBION OXIDATION PLANT ............................................................................................................................ 89

18.4.10. ALBION TAILS THICKENER ........................................................................................................................... 90

18.4.11. GOLD RECOVERY PLANT ............................................................................................................................. 90

18.4.12. CIL TAILINGS THICKENER............................................................................................................................. 91

18.4.13. ELUTION AND CARBON REGENERATION ..................................................................................................... 91

18.4.14. ELECTRO-WINNING AND SMELTING ............................................................................................................ 91

18.4.15. CYANIDE DETOXIFICATION ......................................................................................................................... 92

18.4.16. LIMESTONE MILLING .................................................................................................................................. 92

18.4.17. OXYGEN SUPPLY ......................................................................................................................................... 93

18.4.18. PLANT SERVICES ......................................................................................................................................... 93

18.4.19. TAILINGS MANAGEMENT FACILITIES AND WATER RETICULATION............................................................... 94

18.4.20. WATER TREATMENT PLANTS ...................................................................................................................... 95

19. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ........................................................................................ 109

19.1. BULK DENSITY ................................................................................................................................................ 109

19.2. DATABASE STRUCTURE AND CONTENT ........................................................................................................... 109

19.3. LITHOLOGICAL INTERPRETATION AND 3D MODEL CREATION .......................................................................... 111

19.3.1. MINERALISATION MODELLING .................................................................................................................... 112

19.4. BLOCK MODEL DEVELOPMENT ....................................................................................................................... 112

19.5. STATISTICAL ANALYSIS .................................................................................................................................... 115

19.5.1. STATISTICAL SUMMARY BY ESTIMATION DOMAIN ...................................................................................... 116

19.5.2. INVESTIGATION OF HIGH GRADE OUTLIERS ................................................................................................. 117

19.5.3. ANALYSIS OF DATA CLUSTERING ................................................................................................................. 117

19.5.4. CORRELATION ANALYSIS ............................................................................................................................. 118

19.6. VARIOGRAPHY ................................................................................................................................................ 118

19.7. BLOCK MODEL ESTIMATION AND VALIDATION ............................................................................................... 119


19.7.1. ORDINARY KRIGING..................................................................................................................................... 120

19.7.2. CHANGE OF SUPPORT ESTIMATES FOR GOLD .............................................................................................. 121

19.8. MINERAL RESOURCE/RESERVE CLASSIFICATION AND REPORTING .................................................................. 122

19.9. VALIDATION AND COMPARISON WITH PREVIOUS MODEL .............................................................................. 124

19.10. DEVAGOLD POLYGONAL RESOURCE CALCULATION COMPARISON .................................................................. 125

19.11. DUMP RESOURCE ESTIMATION ...................................................................................................................... 126

19.11.1. DESCRIPTION OF DUMP TOPOGRAPHIC SURVEY AND SURFACE GENERATION .......................................... 126

19.11.2. DUMP RESOURCE ESTIMATE DESCRIPTION ............................................................................................... 127

19.11.3. DUMP RESOURCE RESULTS ....................................................................................................................... 129

19.12. RESERVES ....................................................................................................................................................... 129

19.13. CONCLUSIONS ON GEOLOGY AND RESOURCES ............................................................................................... 130

20. OTHER RELEVANT DATA AND INFORMATION ........................................................................................................... 138

20.1. MINING CONSIDERATIONS ............................................................................................................................. 138

20.1.1. GEOTECHNICAL ANALYSIS............................................................................................................................ 138

20.1.2. PIT OPTIMISATION ...................................................................................................................................... 157

20.1.3. MINERAL RESOURCE MODEL ....................................................................................................................... 158

20.1.4. ECONOMIC INPUT AND PROCESS RECOVERIES PARAMETERS ...................................................................... 160

20.1.5. SENSITIVITY ANALYSIS ................................................................................................................................. 162

20.1.6. ENGINEERED PITS ........................................................................................................................................ 162

20.1.7. UPDATED RSG ENGINEERED PITS ................................................................................................................. 164

20.1.8. MINE PRODUCTION SCHEDULE .................................................................................................................... 164

20.1.9. HAUL DISTANCE PROFILE AND TRUCK FLEET DIMENSIONING....................................................................... 167

20.1.10. MINING EQUIPMENT ................................................................................................................................ 169

20.1.11. WASTE ROCK MANAGEMENT AND STOCKPILE STRATEGY ......................................................................... 173

20.1.12. MANPOWER REQUIREMENTS ................................................................................................................... 174

20.1.13. MINING PROGRAMME ............................................................................................................................. 175

20.1.14. MINING RESERVES .................................................................................................................................... 176

20.1.15. CONCLUSIONS ON MINING ....................................................................................................................... 177

20.2. TAILINGS MANAGEMENT FACILITY ................................................................................................................. 207

20.2.1. INTRODUCTION ........................................................................................................................................... 207

20.2.2. INITIAL DESIGN ............................................................................................................................................ 207

20.2.3. DATA REVIEW .............................................................................................................................................. 208

20.2.4. SITE LAYOUT ................................................................................................................................................ 208

20.2.5. DESIGN CRITERION ...................................................................................................................................... 208

20.2.6. DAM EMBANKMENT WALLS ........................................................................................................................ 209

20.2.7. WATER MANAGEMENT ............................................................................................................................... 211

20.2.8. TMF CONCLUSIONS ..................................................................................................................................... 213

20.3. ENVIRONMENTAL CONSIDERATIONS .............................................................................................................. 219

20.4. FINANCIAL ANALYSIS ...................................................................................................................................... 220

20.4.1. INTRODUCTION ........................................................................................................................................... 220

20.4.2. FINANCIAL ASSUMPTIONS ........................................................................................................................... 220

20.4.3. PRODUCTION SCHEDULE ............................................................................................................................. 220

20.4.4. OPERATING COSTS ...................................................................................................................................... 222

20.5. 13.5 OTHER COSTS .......................................................................................................................................... 226

20.5.1. CAPITAL COSTS ............................................................................................................................................ 227

20.5.2. TAXATION AND DEPRECIATION ................................................................................................................... 229

20.5.3. PROJECT FINANCING ................................................................................................................................... 229

20.5.4. OPPORTUNITIES FOR COST REDUCTIONS ..................................................................................................... 230

20.5.5. PROJECT RETURNS AND SENSITIVITY ........................................................................................................... 231

20.5.6. BENEFITS TO THE REGION AND TO ROMANIA .............................................................................................. 233

20.5.7. CONCLUSIONS ............................................................................................................................................. 233

21. INTERPRETATION AND CONCLUSIONS ...................................................................................................................... 235

21.1. GEOLOGY AND RESOURCES ............................................................................................................................ 235


21.2. MINING .......................................................................................................................................................... 236

21.3. PROJECT CONCLUSIONS .................................................................................................................................. 237

22. RECOMMENDATIONS ............................................................................................................................................... 238

23. REFERENCES ............................................................................................................................................................. 239

List of Tables

Table 8-1: Certej Historic Resource Estimation Summary ...................................................................................................... 18

Table 8-2: Certej Project Historical Work Summary .............................................................................................................. 18

Table 11-1: Mineralising Events .......................................................................................................................................... 29

Table 11-2: Deposit Zonation .............................................................................................................................................. 32

Table 13-1: Drilling and Sampling Statistics ......................................................................................................................... 39

Table 13-2 RC and Channel Summary ................................................................................................................................. 40

Table 14-1: Theoretical Sample Recoveries RC .................................................................................................................... 42

Table 14-2: Summary Calculated Diamond Drilling Recoveries ............................................................................................ 43

Table 16-1 Rocklabs standards submitted to SGS and ALS .................................................................................................. 53

Table 16-2 Comparison Statistics, Au field duplicates >= 0.1 Au ppm. ................................................................................ 54

Table 16-3 Comparison Statistics, Ag field duplicates >=0.5Ag ........................................................................................... 54

Table 16-4 Comparison statistics, Au laboratory replicates and repeats ............................................................................. 57

Table 16-5 Comparison statistics, Ag laboratory replicates and repeats ............................................................................. 57

Table 16-6 Comparison statistics, Au laboratory preparation repeats (pulp splits) ............................................................. 59

Table 16-7 Summary of check assay data from SGS & ALS 2004-2005 and 2008 ................................................................ 62

Table 18-1: Flotation Concentrate Composition from the 4 ore zones at Certej .................................................................. 67

Table 18-2: Refractoriness and Direct Leachability by ore zone ........................................................................................... 68

Table 18-3: Solid solution gold carried in sulphide minerals by ore zone ............................................................................. 69

Table 18-4: Distribution of Gold by ore zone in Certej ......................................................................................................... 70

Table 18-5: The Deportment of Silver in the flotation concentrate ..................................................................................... 71

Table 18-6: Distribution of Silver by ore zone in Certej ....................................................................................................... 72

Table 18-7: Summary of Total Samples produced for Metallurgical Testwork ..................................................................... 74

Table 18-8: Geological Composition of the Samples used for the Metallurgical Testwork ................................................... 74

Table 18-9: Locked Cycle Test Data ..................................................................................................................................... 76

Table 18-10: Input Figures for Desktop Evaluation of the Optimum Process Treatment Route of Certej Flotation

Concentrates.............................................................................................................................................................. 78

Table 18-11: Hellas Gold Head Sample Assays .................................................................................................................... 79

Table 18-12: Grades and Calculated Recoveries with Recirculation of Cleaner Tailings ....................................................... 79

Table 18-13: West Concentrate, Sulphide Oxidation / Gold Recovery Profile ...................................................................... 81

Table 18-14: Gold Deportment in Albion and CIP tailings .................................................................................................... 82

Table 18-15: Deportment in Albion and CIP tailings ............................................................................................................ 83

Table 18-16: Overall Recoveries from Hellas and HRL testwork .......................................................................................... 83

Table 18-17: Comminution Circuit Parameters ................................................................................................................... 84

Table 19-1: Bulk Densities by Lithology ............................................................................................................................. 109

Table 19-2: Block Model Dimensions ................................................................................................................................ 112

Table 19-3: Block Model Variables .................................................................................................................................... 112

Table 19-4: Block Model Bulk Density Assignment ............................................................................................................ 114

Table 19-5: Summary Statistics for Composite Data Composite Data Defined within the Mineralisation Envelopes .. 116

Table 19-6: Certej - Gold 3m Composites, Summary Statistics ........................................................................................... 116

Table 19-7: Outlier Analysis - Gold (g/t) ............................................................................................................................ 117

Table 19-8: Comparison of Raw and Declustered Mean Grades by Estimation Domain ............................................. 118

Table 19-9: Variogram Model Parameters ........................................................................................................................ 119

Table 19-10: Sample Search Parameters - Gold and Silver Ordinary Kriging ...................................................................... 120

Table 19-11: Uniform Conditioning Parameters ................................................................................................................ 122

Table 19-12: Resource Classification Criteria..................................................................................................................... 123

Table 19-13: Total Resources All Domains UC Estimate at 0.8 g/t cut Off ............................................................................ 124


Table 19-14: Comparison of Raw and De-clustered Mean Grades by Estimation Domain (based on cut gold and silver

data) 124

Table 19-15: Comparison between 2005 and 2007 SMU Estimates Uniform Conditioning - Gold g/t ......................... 125

Table 19-16: Deva Gold Resource Estimate Domains .................................................................................................... 125

Table 19-17: Certej Polygonal Resource Estimation Comparison with SMU model ............................................................ 126

Table 19-18 Polygonal Indicated Estimate for North and South Coranda Dumps. ............................................................. 129

Table 19-19: Certej Mineral Reserves ............................................................................................................................... 130

Table 20-1: Summary structural data from existing open pit............................................................................................. 139

Table 20-2: Summary structural data from geotechnical and re-logged exploration boreholes .................................. 143

Table 20-3: Open pit slope angles and heights in existing pit ............................................................................................ 144

Table 20-4: West pit typical final pit slope heights ............................................................................................................ 145

Table 20-5: West pit typical geology at final pit profile ..................................................................................................... 145

Table 20-6: Central pit typical final pit slope heights ......................................................................................................... 145

Table 20-7: Central pit typical geology at final pit profile .................................................................................................. 146

Table 20-8: East pit typical final pit slope heights ............................................................................................................. 146

Table 20-9: East pit typical geology at final pit profile ....................................................................................................... 146

Table 20-10: Summary of rock mass quality and geology for geotechnical boreholes .................................................. 149

Table 20-11: Summary of rock mass quality and geology for re-logged exploration boreholes ................................... 150

Table 20-12: Summary of rock mass for main rock types .................................................................................................. 151

Table 20-13: Rock mass shear strength for major depth zones within open pit ................................................................. 152

Table 20-14: Rock mass shear strength for main rock types within open pit ..................................................................... 152

Table 20-15: Slope stability assessments for each sector .................................................................................................. 153

Table 20-16: Block failure assessment for each sector ...................................................................................................... 155

Table 20-17: Slope profile summary ................................................................................................................................. 159

Table 20-18: Whittle input parameters source .................................................................................................................. 160

Table 20-19: Whittle economic input parameters and metallurgical recoveries ................................................................ 160

Table 20-20: Pit slope parameters .................................................................................................................................... 162

Table 20-21: Engineered pit material inventories ............................................................................................................. 163

Table 20-22: East Pit Material Inventories ........................................................................................................................ 164

Table 20-23: Updated Engineered pit material inventories ............................................................................................... 164

Table 20-24: Mine Production Schedule ........................................................................................................................... 165

Table 20-25: Truck operator’s efficiency profile ................................................................................................................ 167

Table 20-26: Loading unit parameters .............................................................................................................................. 168

Table 20-27: Truck fleet analysis summary ....................................................................................................................... 169

Table 20-28: Mining equipment ........................................................................................................................................ 170

Table 20-29: Major equipment shift operating time ......................................................................................................... 171

Table 20-30: Mine waste schedule .................................................................................................................................... 174

Table 20-31: Certej mineral reserves ................................................................................................................................ 176

Table 20-32: Catchment Area Summary ............................................................................................................................ 212

Table 20-33 Drainage Gallery Characteristics ................................................................................................................... 213

Table 20-34: Mine Schedule for Financial Model (Years 1-12) ........................................................................................... 221

Table 20-35: Mine Operating Cost Summary ..................................................................................................................... 224

Table 20-36: Concentrator Cost Summary ......................................................................................................................... 225

Table 20-37: Albion Cost Summary ................................................................................................................................... 225

Table 20-38: Composition for Doré Market Survey ........................................................................................................... 227

Table 20-39: Capex Inputs to Cash flow Model ................................................................................................................. 229

Table 20-40: Cash flow Model Indicators .......................................................................................................................... 231

Table 20-41: NPV Sensitivity to Commodity Prices ............................................................................................................ 232

Table 20-42: IRR Sensitivity to Commodity Prices ............................................................................................................. 232

Table 20-43: Payback Sensitivity to Commodity Prices .................................................................................................... 232

List of Figures

Figure 6-1: Certej Project Location ........................................................................................................................................ 9


Figure 7-1: Certej Local Infrastructure ................................................................................................................................. 13

Figure 9-1: The Golden Quadrilateral .................................................................................................................................. 23

Figure 9-2: Regional Tectonics ............................................................................................................................................ 24

Figure 9-3: Geology of the Certej Brad Belt ......................................................................................................................... 25

Figure 9-4: Geology of the Certej Concession ...................................................................................................................... 26

Figure 11-1: Certej West-East Geological Section ................................................................................................................ 33

Figure 11-2: Certej West-East Section - Alteration .............................................................................................................. 34

Figure 11-3: Certej Deposit Zonation .................................................................................................................................. 35

Figure 14-1: Distribution of Channel, Diamond Drillhole and RC Sampling – Plan View ....................................................... 45

Figure 14-2 RC Drillhole Collars and Channel Line Locations Over the North and South Coranda Dumps. ........................... 46

Figure 15-1 Conventional assay procedures and quality control at Deva Gold for gold and base metal analyses ................ 50

Figure 16-1 Scatter and Q-Q plots for Au in Dump Field Duplicates Database. ................................................................... 55

Figure 16-2 Scatter and Q-Q plots for Ag in Dump Field Duplicates Database. .................................................................... 56

Figure 16-3 Scatter Plots for Au and Ag in Dump Laboratory Duplicates Database. ............................................................ 58

Figure 16-4 Comparison plot for Au laboratory preparation repeats (pulp splits) ............................................................... 60

Figure 18-1: Plots of Arsenic and Tellurium ......................................................................................................................... 97

Figure 18-2: Sulphide Oxidation vs. Au recovery for Initial Albion Tests .............................................................................. 98

Figure 18-3: Comparison of Treatment Options; IRR ........................................................................................................... 99

Figure 18-4: Schematic Diagram of Continuous Albion Testwork Equipment .................................................................... 100

Figure 18-5: Gold and Silver Recovery vs. Sulphide Oxidation for West Concentrate ......................................................... 101

Figure 18-6: Schematic Diagram of Certej Process Flowsheet ............................................................................................ 102

Figure 18-7: General Schematic Diagram of Certej Processing Route ................................................................................ 103

Figure 18-8: Processing Route for the Treatment of the Flotation Concentrate ................................................................. 104

Figure 18-9: General Site Layout ....................................................................................................................................... 105

Figure 19-1: Certej Lithology Model – West East Cross Section ......................................................................................... 132

Figure 19-2: Grade Tonnage Curves for the UC and OK Certej Estimates ........................................................................... 133

Figure 19-3: Plot showing East Zone Intercepts versus Grade Model................................................................................. 134

Figure 19-4: Comparison of Drillhole Composite and SMU Model Grades, East Zone. ....................................................... 135

Figure 19-5: Comparison of 2005 and 2007 Certej SMU Resource Estimates ..................................................................... 136

Figure 19-6 North and South Dump Wireframes ............................................................................................................. 137

Figure 20-1: Stereonet of rock mass structure, existing open pit ....................................................................................... 178

Figure 20-2: Rosette of rock mass structure, existing open pit .......................................................................................... 179

Figure 20-3: Stereonet of rock mass structure, geotechnical and re-logged boreholes ...................................................... 180

Figure 20-4: Rosette of rock mass structure, existing open pit – joints >3m ...................................................................... 181

Figure 20-5: East pit section views .................................................................................................................................... 182

Figure 20-6: West pit section views .................................................................................................................................. 183

Figure 20-7: East and west pit long section views ............................................................................................................. 184

Figure 20-8: Plan of Open Pit Showing Sectors, Geotechnical and Re-Logged Boreholes ................................................... 185

Figure 20-9: Pit size vs. commodity price .......................................................................................................................... 186

Figure 20-10: Pit optimization results ............................................................................................................................... 187

Figure 20-11: Ore to waste relationship ............................................................................................................................ 188

Figure 20-12: Scheduling cutbacks .................................................................................................................................... 189

Figure 20-13: Topography ................................................................................................................................................. 190

Figure 20-14: Pre-Production ............................................................................................................................................ 191

Figure 20-15: Pit year 1 ..................................................................................................................................................... 192

Figure 20-16: Pit year 2 ..................................................................................................................................................... 193

Figure 20-17: Pit year 3 ..................................................................................................................................................... 194

Figure 20-18: Pit year 4 ..................................................................................................................................................... 195

Figure 20-19: Pit year 5 ..................................................................................................................................................... 196

Figure 20-20: Pit year 6 ..................................................................................................................................................... 197

Figure 20-21: Pit year 7 ..................................................................................................................................................... 198

Figure 20-22: Pit year 8 ..................................................................................................................................................... 199

Figure 20-23: Pit year 9 ..................................................................................................................................................... 200


Figure 20-24: Pit year 10 ................................................................................................................................................... 201

Figure 20-25: Pit year 11 ................................................................................................................................................... 202

Figure 20-26: Pit year 12 ................................................................................................................................................... 203

Figure 20-27: General mine layout .................................................................................................................................... 204

Figure 20-28: Rimpull and retarding charts for CAT 775E (C282) ....................................................................................... 205

Figure 20-29: Mine Organisation and Workforce .............................................................................................................. 206

Figure 20-30: Location of the Flotation TMF and CIL TMF .................................................................................................. 214

Figure 20-31: Plan of TMF Embankments .......................................................................................................................... 215

Figure 20-32: Flotation TMF Main Embankment Cross Section .......................................................................................... 216

Figure 20-33: Flotation TMF Saddle Embankment Cross Section ....................................................................................... 217

Figure 20-34: CIL TMF Embankment Cross Section ............................................................................................................ 218

Figure 20-35: NPV Sensitivity Plot ..................................................................................................................................... 234

Figure 20-36: IRR Sensitivity Plot ...................................................................................................................................... 234

List of Plates Plate 18-1: Typical Photomicrograph showing Arsenical Zoning of a Pyrite Grain ............................................................. 106

Plate 18-2: Photomicrographs showing the 5 Morphological forms of Pyrite described by Amtel ..................................... 107

Plate 18-3:Gypsum Crystal in Oxidised Albion Residue ..................................................................................................... 108

Certej Updated Definitive Feasibility Study Summary Technical Report Page 1 of 241

3. SUMMARY

European Goldfields Limited holds four mineral properties located within the “Golden Quadrilateral” area of western Romania. This is a mining district in the Apuseni Mountains of Transylvania covering an area of approximately 500 km2 immediately to the north of the city of Deva. The Certej property are Certej is held through the Company’s 80% interest in Deva Gold S.A. (Deva Gold). The remaining 20% of Deva Gold is held by Minvest S.A. (19.25%), a Romanian state-owned mining company, and three minority Romanian shareholders (collectively, 0.75%). Deva Gold has a mining licence and has applied for the environmental permit required to develop the new Certej gold/silver project. It is the strategy of European Goldfields to manage the development and ultimately the operations of the Romanian projects through Deva Gold. The Romanian mining industry has reduced significantly in terms of both personnel and capacity since the end of the communist era and more recently leading up to Romania joining the European Union in 2007. This was due to the removal of subsidies to non-profitable and polluting operations which was one of the conditions for joining the EU and in January 2007 Minvest S.A. closed its mining and processing operations at Certej which had been steadily run down over the preceding few years. European Goldfields/Deva Gold has carried out an extensive drilling and exploration programme which commenced in 2000 and as a result has developed for the Certej deposit a Measured and Indicated Resource comprising:

Total Resources All Domains Estimate at 0.8 g/t cut Off

Resource Category Million Tonnes Au g/t Ag g/t Measured 3.9 2.3 5 Indicated 37.6 1.9 11 Measured and Indicated Total 41.5 2.0 11 Inferred (East-West) 3.4 1.6 4 Inferred (Far West/Central Domains) 3.8 1.4 8 Inferred Total 7.1 1.5 6

In addition a polygonal estimate has been carried out on existing dumps at Certej:

Category Volume SG

In-situ

Tonnage

Au

g/t

Au

Ounces

Ag

g/t

Ag

Ounces

Certej North

Dump Indicated 2,141,307 1.44 3,083,482 0.53 52,500 8.29 821,800

Certej South

Dump Indicated 2,735,383 1.44 3,938,951 0.53 67,100 9.32 1,180,300

Total Indicated 4,876,690 1.44 7,022,434 0.53 119,600 8.87 2,002,100


The deposit resource was estimated using a block model with grade interpolation carried out using geostatistical methods with a change of support using Universal conditioning. The main mineralised zone is some 1,500 meters long by 500 meters wide and occurs as sub-horizontal, to moderately dipping zones hosted by Cretaceous and Neogene sediments with andesitic rock intrusions. The ICP, mineralogical and metallurgical test work clearly shows that the deposit can be divided into four main ore types, Central, Intermediate, East and West. It was apparent from the structure of the resource that open pit mining will be the most cost effective method of mineral extraction which was also used by the previous operator. The structure of the deposit and the behaviour of the four ore types results in the Central, Intermediate and East materials being mined and processed together while the West will be treated separately at the end of the mine life. The resource estimation study included a review of the exploration data quality, review and construction of appropriate geological constraints, statistical and geostatistical investigation, grade estimation, and classification of the estimate in accordance to the criteria laid out in the Canadian National Instrument 43-101. The project was then further advanced by developing mining and processing strategies with associated operating and capital costs which enabled the mineable reserve for the Certej deposit to be calculated which again conformed to N.I. 43-10 as follows:- Certej mineral reserves ‘000t Au Au Ag Ag

g/t Moz g/t Moz

Certej In-pit full grade

Probable 32,811 2.0 2.12 11.4 12.0

Certej In-Pit Lower Grade

Probable

7,829 0.7

0.18 14.0 3.5 Certej Existing Dumps

Probable

6,320 0.5

0.11 8.9 1.8

Total

46,960

2.41 17.3 The deposit extends from surface and will be mined by open pit methods with a strip ratio of 3.1:1. The project will involve the mining and processing of 3.0 million tonnes of ore per annum. In 2006, Deva Gold completed all the necessary Environmental Impact Assessments (Levels I and II) and a Social Impact Assessment Study in support of the permit application to develop the Certej project.


In September 2006 the Hunedoara County Council issued a General Urbanisation Certificate for the Certej project. This confirms the designation of Certej as an industrial mining area and verifies local community support for the project. Deva Gold submitted a Technical Feasibility Study to the Romanian government authorities, which include the National Authority for Mineral Resources, NAMR, in March 2007 and the Environmental Impact Study (EIS) in November 2007 to support the application for environmental and construction permits. The EIS comprised a detailed multi-discipline study assessing the environmental, social and health impacts of the project on the affected area. It was carried out over a period of a year covering the four seasons to accumulate all the base line data required. In parallel with the EIS the detailed urbanisation plan will be prepared and submitted at the same time. The necessary construction, mining and operating permits are expected to be awarded by mid-2009 following the public consultation process with the local community which will be attended by all interested parties. The process plant will comprise a conventional +/-9,500 tpd concentrator with crushing, grinding and sulphide flotation unit operations. The refractory gold and silver bearing sulphide flotation concentrate will then be directed to the Albion Process plant where the (pyrite sulphide) concentrate will be oxidised and the solid residue will be directed to the CIL and precious metals plant to recover the liberated gold and silver as a doré. The Albion Process is a combination of ultra-fine grinding of concentrates followed by oxidative leaching at atmospheric pressure. The process does not employ autoclaves or rely on bacterial cultures. The residues from the flotation and gold plants will be disposed of in two separate but adjoining TMFs, Tailings Management Facilities. The project is scheduled to produce an average of 308,000 tonnes of flotation concentrate per annum with high grades between 17 – 19 g/t gold and 80 – 130 g/t silver, depending on the source of the ore in the deposit, with a flotation gold and silver recoveries of approximately 90%. The Albion gold and silver recoveries will be approximately 90%and 85% respectively, resulting in total process gold and silver recoveries in the region of 81% and 77%. The silver recoveries for the west ore type will be slightly lower than the other three zones. This will yield approximately 160,000 oz of gold and 820,000 oz of silver per year in doré. European Goldfields has completed an extensive metallurgical testwork programme using the Albion Process at the facilities of Hydrometallurgical Research Laboratories Testing (HRL) in Australia. Initial small scale batch tests achieved gold recoveries of 90-93% and established the optimum conditions for the continuous pilot plant testwork required to prove the suitability of the Albion Process for the Certej concentrate. A Phase 1 continuous pilot scale test campaign was carried out on a composite of the Central, Intermediate and East ore types in the ratio that will be mined and this confirmed that gold extractions in excess of 90% could be achieved on a continuous basis at sulphur oxidation rates in the Albion Process of around 70%. The testwork also confirmed the consumable levels were in the expected range.


A programme of flotation concentrate grade optimisation work was then undertaken and on completion a second Phase 2 continuous pilot plant run was completed in June 2007. This confirmed the high gold extractions of 90% achieved in the Phase 1 run and that the flotation concentrate is amenable to treatment by the Albion Process. A continuous CIL pilot plant was then run for 4½ days on the product from the Phase 2 Albion Process run. Further investigatory work and equipment tests by potential equipment vendors was also undertaken and used as the process design basis for this engineering bankable feasibility study. During the testwork campaign at HRL a large batch scale Albion test was completed on the concentrate from the West ore zone of the Certej deposit which had proved to be less amenable to the Albion Process. This work demonstrated that by applying a slightly higher oxidation rate, a gold recovery of 90% could also be achieved from the West ore zone. A definitive mineralogical study describing the four ore zones of the Certej deposit was completed by Amtel of Canada which indicated that there is potential for increasing flotation gold and silver recoveries above those achieved in the laboratory scale tests. A review of the process testwork and an engineering study of the process plant and associated infrastructure has been undertaken by Aker Kvaerner Engineering Services of Stockton, UK. The conceptual design of the tailings management facilities was developed by world renowned Romanian professors and the detail study and cost estimated prepared by Golder Associates UK. The area has experienced a substantial economic revival in the past four years with major investments from international and local corporations. It is served by good infrastructure with 110kV power supply and water pipelines arriving within two kilometres of the mine. The project has paved roads directly to site and the region has a large road-building programme to improve these further. The Certej project also benefits from two rail loading facilities 18 kilometres distant at the major rail-head at Deva. Deva is connected to the main Black Sea port of Constantia by the Romanian highway and rail network and is serviced by three international airports, all within two hours drive of the project. The project will employ over 300 people from the Certej area whose recent mining history ensures a good skills base is available in the local labour force. There are no settlements in the vicinity of the proposed mine and TMF sites. Detailed field work has established that there are no archaeological remains on the site. Both the mine site and the TMFs are shielded by topography and there is no visual impact on settlements. All the necessary studies to comply with Romanian legislation and international best practice have been completed. The financial returns achieved by the project show that it is robust at metal prices of $650 per ounce for gold and $7.5 per ounce for silver and the IRR exceeds the company threshold of 20%. The Certej Definitive Feasibility Study, summarised in this report,


indicates a viable 16 year project and it is recommended that it should now progress to basic and final engineering stage prior to financing and construction on approval of the relevant environmental and construction permits.


4. INTRODUCTION AND TERMS OF REFERENCE

The Certej epithermal gold silver deposit is located in the “Golden Quadrilateral” area of the Apuseni Mountains of Transylvania in Western Romania immediately to the north of the city of Deva. The deposit was first identified in the 18th Century and was subject to minor ad-hoc mining activity until the 1960's when the Romanian state started to exploit the deposit on a small scale mainly for base metals from narrow sulphide veins from underground development. During the 1980's operation was concentrated on an open pit called the Coranda Pit situated in the Central part of the deposit which continued operating until 2006. European Goldfields acquired 80% of the project through its subsidiary Deva Gold SA in 2000. Deva Gold's activities were confined to exploration and the Coranda Pit operation remained in the hands of the Romanian State mining company Minvest. European Goldfields/Deva Gold has carried out an extensive drilling and exploration programme which commenced in 2000 and to date defines a Measured and Indicated Resource comprising 41.5 Mt of ore with grades of 2.0 g/t Au & 11 g/t Ag at a 0.8 g/t Au cut-off. The main mineralised zone is some 1,500 meters long by 500 meters wide and occurs as sub-horizontal, to moderately dipping zones hosted by Cretaceous and Neogene sediments and intrusive andesites. The deposit can be divided into four main ore types, Central, Intermediate, East and West. The mineable reserve comprises 32.8 million tonnes of ore grading 2.0 g/t gold and 11.4 g/t silver, representing 2.1 million ounces of gold and 12.0 million ounces of silver with extraction by conventional open pit with a strip ratio of 3.1:1. All resources and reserves are Canadian National Instrument 43-101 compliant The project will involve the mining and processing of 3.0 million tonnes of ore per annum over at least eleven years, with the Central, Intermediate and East zones being mined early in the project life with the West zone, which forms a separate open pit being mined towards the end of mine life. It is proposed that the run of mine (ROM) ore be processed first using flotation to produce a pyrite gold concentrate and then by ultra fine grinding and ambient pressure oxidation using the Albion method developed by Xtrata to produce gold and silver as doré on site. Annual metal production will, on average, be some 160,000oz’s of Au and 800,000oz’s of silver. The area is served by good infrastructure with an 110kV line and water supply arriving within two kilometres of the site. The historic mining conducted in the area also ensures a good skills base is available in the local labour force. It is expected to receive the necessary permits and secure finance to start building the project in 2009 for a plant start up in 2010.


5. RELIANCE ON OTHER EXPERTS

The mineral resources and reserves quoted in this report were estimated by RSG Global. Their report was filed in full on SEDAR in November 2007. The same resources and reserves are quoted here and all the data behind this is described in full in order that this technical report is Canadian National Instrument 43-101 (NI 43-101 compliant). The resource and reserve estimates follows the CIM guidelines currently utilised by NI 43-101 and use current cost and financial factors in reserve generation. In addition, work has been completed by various experts on behalf of European Goldfields to complete this definitive feasibility study as follows:-

• Geology - Deva Gold with review by RSG Global, now part of Coffey Mining • Resources and Reserves - RSG Global, now part of Coffey Mining • Mining Study – European Goldfields and In Silico Mining • Process Route - Core Resources (subsidiary of Xtrata) • Plant Design and Cost Estimation – Aker Kvaerner Engineering Solutions (AKES) • Dam Design - Golder Associates and Cepromin

This report reflects the findings of these studies but European Goldfields is responsible for the contents of this report.


6. PROPERTY DESCRIPTION AND LOCATION

The Certej Project is located in the southern part of the Apuseni Mountains in central Romania (Figure 6-1), some 12km NNE of the regional town of Deva in Hunedoara County. The area has a history of mining and the Coranda Open Pit situated in the central part of the deposit operated until early 2007. The nearby town of Deva is served by a National Highway and a railway line from Bucharest. A major river, the Mures, runs within 10 kilometres of the proposed site and 110 kVA power lines and water come to within 2 kilometres of the site. Tarmac run directly to the deposit. European Goldfields currently holds nine mineral properties located within the “Golden Quadrilateral” area of the southern Apuseni Mountains, Romania. The Certej license covers an area of approximately 26.7 square kilometres in the Certej area of Hunedoara county and the company holds a further eight licences covering some 507 square kilometres for around 534 square kilometres of licences over all. The license was granted for a period of 20 years, with an initial development-exploitation period of 5 years commencing on the day the license was gazetted on the 25th of January 2000, under the terms of the Romania Mining Law No. 61/1999. The concession is held in the name of European Goldfields subsidiary Deva Gold. The License was initially granted by the National Agency for Mineral Resources to Minvest as the “titleholder” and Deva Gold as the “affiliated company” and was subsequently transferred in accordance with its terms by Minvest to Deva Gold further to an additional act concluded on 10 December 2001, which entered into force on 18 December 2001, so that Deva Gold becomes the “titleholder” and Minvest the “affiliate”. Minvest currently own 20% of the project and Deva Gold hold 80% Deva Gold currently employs some 55 staff in Romania to pursue the exploration and successful development of its concessions and projects. To date, European Goldfields has invested over US$ 27.3 million in Romania, including US$ 21.7 million on the Certej Project. European Goldfields is engaged in the acquisition, exploration and development of precious and base metals mineral resource properties primarily in Romania and Greece. The company's main administrative office is located in London England.


Figure 6-1: Certej Project Location


7. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

7.1. Access

The Certej Project is accessed by a sealed two lane road from the edge of the proposed open pit to Certej and then by the National highway E68 to Deva, as detailed in Section 10. The National highway E68 connects Bucharest to Arad and passes through Deva. The National Rail system also connects Deva to Bucharest and several other regionally important towns such as Timisoara, Cluj and Arad. Access is not impeded during periods of heavy snow as the local and national road authorities have vehicles which clean the accumulated snow off the roads and also apply grit and salt to prevent snow and ice build up.

The closest international airport is at Sibiu approx 109km towards Bucharest from Deva on the National highway E68. International airports are also located at Cluj (180km) and Timisoara (168km). A regional airport is also planned for construction near Deva at the existing small airstrip as part of EU accession.

7.2. Climate

The climate is mild temperate-continental, with mean temperatures of around 23°C in summer and minus 3°C in winter. Snow cover can last until March and may be up to 80cm thick in the forested areas. During the winter, in snow-free areas, frost penetration is about 180cm. Precipitation as rain and snow annual average is 562 mm with a minimum monthly average 27.0 mm and a maximum monthly average 72.4 mm. Using monthly averages taken over a 20 year period from 1987 to 2006, the minimum monthly air temperatures range from -4.4 to 14.4oC in January and February and the maximum monthly average air temperature ranging from 2.4 to 27.9oC in August. Over the same period the maximum absolute air temperature was 39oC and the minimum absolute air temperature was -25oC. Exploration and mining activities can continue throughout the year with no problems.

7.3. Local Resources and Infrastructure

Deva is the largest town near the mining operation with a total population of 69,000 and housing the vast majority of the previous mine’s national workforce. The village of Certej de Sus, population around 3,500, is located approximately 15km northeast of Deva is located next to the operation and is also a source of labour (Figure 7-1). A sealed two lane road provides access between Deva and Certej and the main highway from Bucharest to Arad passes through Deva. The national railway system also connects Deva with Bucharest and the rest of Romania. Minvest is the current owner/operator of mining operations in the area. The Certej operation was shut down in 2006 because it was uneconomic and all subsidies were suspended as part of Romanians EU accession in 2007. A ready source of skilled mining labour is available in Certej and the surrounding villages due to the closure of the Minvest operations, leaving most of these people


unemployed due to the lack of suitable employment opportunities in the region. The unemployment statistics for the town are around 70% versus a national average of around 6%. The immediate area around the mining operation includes the three villages of Certej de Sus, Hondol and Bocsa. These villages and the town of Certej would benefit with better employment opportunities and the flow on effect of investment in the project. The Certej Gold Project is presently serviced by the Mintia power station which is run by the Romanian state and located 14km away with a 110Kv existing power line to the town of Certej, some 2 kilometres from the proposed plant site, and a 20Kv line from this substation to the current Certej open pit. Processing water is currently drawn from the Mures River about 11km away using an existing pipeline installed for the Minvest flotation plant at Certej, only 2 kilometres from the plant site proposed in this report, which can supply excess capacity than needed for the current plant design. Current operations have never been limited by the supply of water for processing. Potable water is currently sourced from an underground aquifer near the current Certej open pit. The national oil company Petrom has a depot in Deva for diesel fuel and oil supplies which supplies the entire Hunedoara County. It is envisaged diesel would be trucked to site from Deva. In addition the Heidelberg Cement, a multinational company, operates a major cement plant located 15km from Certej in Chiscadaga for the supply of cement to a local concrete plant for construction of the Certej Plant and infrastructure. Limestone for the Albion process will be sourced from a limestone deposit located within the company’s Voia concession 10km to the north of the Certej Project. Access to the limestone will be via an upgrade of an existing forestry road. No villages are located in this area only state owned forestry concessions. Expatriate staff and itinerant consultants and contractors presently reside in the town of Deva. All national employees live in the surrounding towns and are transported to and from work in a company supplied minibus from Deva.

7.4. Physiography

Elevations within the project areas range from 400 metres to over 1,100 metres above sea level, but with the hills around the deposit itself reaching a maximum of around 600 metres. Much of the area is hilly, with incised valleys. The hills are forested with beech and oak with occasional conifers, particularly at higher elevations. Scattered small rural settlements, associated pasture, and other agricultural land occur at lower elevations, on gentler slopes and in the river valleys. The proposed site for the Certej project itself comprises a large area previously used for open pit mining with associated dumps, secondary and tertiary forest and rough pasture.


The hills continue to the north, east and west of Certej. The Certej valley drains the project area and the terrain grades into the Mures River alluvial floodplain to the south of the project area.


Figure 7-1: Certej Local Infrastructure


8. HISTORY

The development of exploration and mining at Certej and surrounds can be broken down into the following periods:

• Austro - Hungarian Empire era (end of 17th Century to 1918). • Inter-war period (1918 to 1939). • Modem era (post World War II to present).

8.1. Prior to World War 2

Gold mining in the southern Apuseni Mountains dates back more than two millennia to pre-Roman times and it has been estimated that between 20 and 40 million ounces have been extracted to date. Mining in the Certej area itself dates back several hundred years. This has resulted in the presence of numerous old audits. Early information on the Certej deposit dates back to the 17th century and are included in the history of Alba Iulia from Prince Gabriel Bethlen’s times, considered to be a founding metropolis for SE European development. There is visible evidence of surface and underground mining operations performed in the 18th century with the boundaries of an artificial pond identified in the area. In the first half of the 18th century the most important mine that existed on the Certej Deposit was named “Csertes Berghandel” and another mine named “Dreifaltig Keit” (Sfanta Treime) was opened in 1744. Andrei Stutz, director of the Physical and Natural Sciences Society in Vienna, the author of the first monograph about mines in the Sacaramb area mentions that at around 1790, on the Baiaga hill slope in a place called Coranda the entire ground surface had been re-worked by human activity. In addition, he noted the existence of 40 stamp-mills along the valley, all discharging water from the underground workings. In September 1763, due to the intense mining activity in the area, a gold processing facility was commissioned and this was operating until 1882. Since 1774, the water required for the facility was sourced from the Certej stream when an artificial lake was built along the Faerag valley with a capacity of over 220,000 m3 which exists to this day. At the end of the 19th century a British company was operating at Certej which then sold the mining concession to Teodor Bornemisa, a local baron, who mined approximately 25 -30 kg of gold annually employing some 100 workers at a mine called “Regina”. In 1925, the “Fortuna and Coranda” company held concessions within the Hondol village area after which the “Mica Brad" Company purchased these and started underground mining operations. Mica Brad and another small company called “Concordia” then mined the area. After the Union of Transylvania with Romania all the local entrepreneurs in the region operated together with the State owned companies and developed mining operations at Sacaramb.


Mining operations were recorded in Bocsa Mare village in the 18th century when a lead ore deposit was partly mined and operations resumed when the Certej mine started in the 1970’s. In the Sacaramb area several native gold bearing veins were discovered and telluride minerals known locally as “sacarambit” and “silvanite” were recorded there for the first time. In 1835 the requirement to provide qualified, skilled personnel to work in the local underground mines resulted in the founding of one of the first technical mining schools in Europe at Sacaramb. The school ran 3 year courses and functioned for about 72 years producing a total of 1,213 graduates who all worked at Sacaramb or at other deposits in Transylvania. Between 1882 and 1917 a Sacaramb Mining Entrepreneur exploited the Hondol vein system at the Certej Deposit. The mining works resulted in large dumps being formed near the portals of the underground drives. In 1935 the Humbold Company built a cableway from Sacaramb to Certej and a flotation process plant at Certej in order to process these dumps and produce gold and base metal concentrates.

8.2. Post World War 2

The exploration work conducted before 1965 on the Coranda ore deposit was restricted to underground exploration drives designed to outline the Baiaga igneous strata and to identify zones of gold and silver vein mineralisation. In 1970 Minvest, the government mining company, commenced exploitation of the Bocsa base metal deposit located about 1km east of the Certej resource using the old ore processing plant. A review of exploration data in 1972 concluded that the hydrothermal alteration zone in Coranda hill was hosting a disseminated base metal and gold deposit. In 1973 IPEG, Deva, a Romanian Government organisation, conducted exploration work consisting of additional underground investigatory drives and 50m by 50m grid spaced drilling on the eastern side of Baiaga to further define the Certej orebody. Between 1975 and 1977 the flotation plant in Certej was expanded by installing a new processing line giving a capacity of around 90,000tpa. In 1982 the operation focussed on base metal, Pb and Zn, mining and a new Concentrator was constructed with a throughput of approximately 1,000,000tpa at Certej. In 1983 the Minvest owned Certej Mine took over the Baiaga-Hondol deposit, (the Central and West part of Certej), from IPEG-Deva and the exploration and pre-stripping work of the deposit continued. From 1985 to 1988 the Certej mine produced between 800,000 t and 1,000,000tpa of ore. The open pit was called Coranda. After 1991, Minvest undertook underground development and diamond drilling of the gold mineralisation at Dealul Grozii, (the East part of Certej deposit) and from 1993 underground exploratory drives and diamond drilling in the Coranda Mica zone (West). Detailed metal production records from the state run mining operations at Certej are not available. However it is reported the major product was a lead/silver/zinc concentrate


and, on occasion, a gold/silver bearing pyrite concentrate, depending on where the deposit was being mined. In January 2006 MINVEST S.A. Deva closed its mining and processing operations at the Coranda open pit and Certej town.

8.3. Recent History

The Certej licence was granted in 1999 and European Goldfields gained an 80% interest in the project in 2000 through its subsidiary Devagold SA. After some two years of exploration comprising surface and underground channel sampling and reverse circulation (RC) and diamond drilling on a nominal 80 metre spacing, an independent estimate of resources was made by consultants RSG Global of Perth of 34.7Mt at 2.1g/t Au and 10 g/t Ag. During this period some metallurgical testwork was undertaken, mostly on samples from existing underground development which were localised in their distribution and subsequent investigation has shown that they were partially oxidised. The resource was assessed by Minproc who proposed a bulk mining approach of 8Mtpa using gold recoveries by direct cyanidation of 56% and 63% for silver based on averages of the afore mentioned metallurgical sampling results. This did not prove to be viable $350/oz gold and $4.5/oz silver. Kvaerner were then commissioned to carry a study on the deposit. They re-examined the deposit and re-modelled the resource in an attempt to define high grade zones based on a series of north-south sections. Whilst the sections cut the overall east-west trend of the orebody, this did not recognise the fact that most of the gold mineralisation is hosted by north-south structures and so the resulting calculation understated the resource with some 25Mt at 2.1 g/t Au. Kvaerner assessed the project on the basis of gold recoveries by direct cyanidation of 69% and 63% for silver and 1.2Mt. They also looked at flotation to a pyrite gold concentrate followed by leaching using Geobiotics with overall 77% gold 56% silver with an option of also sweetening the feed with feed from European Goldfields project in Greece. Only the latter option gave viable returns using $375/oz gold price and $7/oz silver price though it was assumed that existing Certej flotation plant could be used which was never a realistic option given the state of the plant. Given the flawed nature of the resource estimation, little can be drawn from the Kvaerner work that followed.

The management of European Goldfields changed in the second half of 2004 and a competent internal technical team was employed. On re-examination of the deposit data and reports it was concluded by the European Goldfields management that, whilst much of the exploration data gathered by RSG was of a high quality, neither the deposit controls, particularly higher grade cores to the mineralisation, nor its mineralogy and therefore metallurgy were fully understood.


In November 2004 a drilling campaign was commenced that continued in to 2005 with an emphasis on tighter drill spacing in higher grade areas in order to demonstrate that a selective mining unit (SMU) model as applied by RSG in 2002, was appropriate and valid approach. In places the drill spacing was reduced to 20 metres and this indicated that the SMU model had correctly predicted where larger parent blocks contained high grade sub-blocks and so the approach was applied again in the 2005 estimate. The programme of work from October 2004 also emphasised the need for representative metallurgical sampling. In order to complete this, a full understanding of the mineralogy was required. Elements of the deposit zonation were known at this stage such as the presence of Tellurides in the Central area and the elevated arsenic due to arsenic rich pyrite rims in the East area but these had not been mapped accurately. Drill pulps were therefore analysed using ICP multi-element techniques in order to map the characteristics and from this the deposit zones described in the sections above were delineated. Representative samples of core from each zone were then identified to produce composite samples from each zone. This established that the gold was intimately associated with pyrite and was more refractory than previous testwork had indicated and it was decided that, whatever the downstream process, a flotation concentrate was the best way to initially recover the gold. Extensive flotation testwork was completed in 2005 at SGS, Wardell Armstrong and HRL including 11 locked cycle tests. Process options were considered for the flotation concentrate including, but not restricted to, roasting, pressure oxidation, biological oxidation, the Activox process, Geobiotics and the Albion oxidative leach process. Following an analysis of the capital and operating costs for each it was established that the Albion process, which involves an ultra fine grind followed by leaching at atmospheric pressure by an autothermic reaction in agitated tanks requiring the input of oxygen, offered the optimum process route and the best return on investment. At the same time in 2006 reserves were established based on the sale of pyrite-gold concentrates. Following further drilling in 2007 to convert inferred resources falling within the pit to indicated resources, a new resource estimation was carried out and a new pit optimisation and reserve estimation based on the use of the Albion process were completed resulting in the current reserve used in this study of 32.8Mt at 2.0g/t Au and 11 g/t Ag. A technical report was then submitted in 2007 to the Romanian authorities as part of the permitting and project approval process. An EIA was also produced at this time in order to establish the current environmental status of the project area and within this the previous operators of the Certej open pit, Minvest, accepted all responsibility for environmental liabilities outside the planned area of operation of the European Goldfields’ project. A 250 kilogramme bulk concentrate sample was produced from some 2.66 tonnes of drill core from across the deposit by European Goldfields flotation laboratory in Greece and this was used for two stage continuous testwork at HRL’s laboratory in Australia in 2007. The results of this testwork form the basis of the plant design proposed in this study.


The pilot plant scale continuous tests were completed during 2007 and proved the Albion process could provide +90% gold and silver recoveries from the Certej flotation concentrate. In 2007 an infill drilling programme was designed to convert inferred material to the indicated category inside the then current pit design. Drilling was completed and a new resource was calculated by RSG (Coffey Mining) and a new pit optimisation completed. Table 8-1: Certej Historic Resource Estimation Summary

Consultant Year Resource Estimate at 1.0g/t Au Cut-off

Contained Gold Moz’s

RSG 2000 8.1Mt @ 1.9g/t Au & 14g/t Ag 0.50

RSG 2002 34.7Mt @ 2.1g/t Au & 10g/t Ag 2.36

Kvaerner 2004 25.4Mt @ 2.1g/t Au & 11g/t Ag* 1.69

RSG 2005 31.35Mt @ 2.1g/t Au & 11g/t Ag 2.31

All resources are based on measured and Indicated resources classified using the JORC system with the exception of the Kvaerner resource which was unclassified. Table 8-2: Certej Project Historical Work Summary Year Supervision Exploration Completed 1999 Deva Gold Surface Channel Sampling

2000-2003

RSG

Surface and UG Channel Sampling and mapping Surface drilling, Aeromagnetic data capture Completion of Surface and UG sampling 80m spaced surface drilling programme Surface Geochemical sampling programme DGPS Control Point Survey Satellite deposit exploration

2003 Deva Gold

Infill surface drilling programme 40m spacing Satellite Exploration

2004-2006

Deva Gold

Surface channel sampling to define extensions to Certej mineralisation Nominal 20m spaced surface drilling programme in high grade areas Further Infill drilling for metallurgy and upgrading of Inferred to Indicated Geotechnical drilling testwork and report by Golders Pilot plant scale metallurgical testwork EIA Level I and II completed Certej Feasibility Study – Cepromin completed

2007 DevaGold

Pilot Plant scale metallurgical testwork at HRL Infill drilling of Inferred resources Gold deportment mineralogical work by AMTEL EIS started with a consortium of 15 local Romanian Companies


8.3.1. Historical Information On Existing Dumps

Preparation for the dumps commenced in the early 1980’s, prior to Minvest Deva pre-stripping oxide from the top of Certej deposit. Although it has not been possible to obtain complete trucked tonnes for both north and south dumps from Minvest Deva, it is estimated that total ore excavated between 1986 to 2006, from Coranda open pit was 9.2Mt, with no records being available for 1984 and 1985. Based on this information the total production of ore from the Coranda open pit over its operational life is estimated at 10Mt. The strip ratio of ore to waste is unknown making it difficult to estimate the dump volumes. It has also been recognised that parts of the dumps were used for the stockpiling of oxide and primary ore that was never reclaimed. As the original focus of the operation was for base metals, analysis for gold was not a priority. This resulted in higher grade areas in the dumps that contained elevated gold and silver credits. In 2004 S.A. Belevion Ltd was engaged by S.A. Deva Gold Ltd to provide an estimate of the dumps volume and a global estimate of 4.96Mt was calculated. This was based on a volume of 1.11Mm3 and 1.81Mm3 in the north and south dumps respectively, considering a bulk density of 2.42 and void factor of 0.7. It is understood that Belevions lower surface was based on the Romanian Cadastral 1980, 1:5,000 topographical sheets. With drilling information now available over most of the dumps, it was established that the 1:5,000 topography can be unreliable at times. This explains Belevions conservative volume. The bulk of the Coranda open pit dumps have been systematically engineered, with steeped benches ranging in height from 5-20m metres. The dumps are still stable with slips and minor slumping evident only on the dumps margins. Vehicle and drill rig access in still easily achieved over most of the dumps benches. The dumps are mainly composed of primary material from the historic open pit operation that has undergone conventional blasting. The dumps do also contain a low percentage of oxide material inkeeping with the deposit itself, with the north dump having more than the south. Field observations confirm that the dumps contain a higher percentage of fines to coarse material. Large boulders are still present over limited parts of the dumps but do not pose any concern. Altered Cretaceous sediments and Neogene sediments and andesite are the three main lithological units within the dumps. Based on field observations of RC chips, surface channel material and upper dump surface, the north dump lithology contains approximately 65% to 35%, Cretaceous to Neogene rocks respectively, with the south dump containing approximately 70% to 30%, Cretaceous to Neogene rocks respectively.


9. GEOLOGICAL SETTING

The Certej exploitation license lies within the ‘Golden Quadrilateral’ an area located in the Apuseni Mountains of Transylvania in Romania, which is part of the Carpathian-Balkan province of the Tethyan Arc and has produced some 40 million ounces of gold according to some estimates (Figure 9-1). The district hosts numerous known epithermal and mesothermal Ag-Au, Cu-Au and Cu, gold, silver and base metal deposits associated with the Neogene andesitic-dacitic volcanic and sub-volcanic bodies which are intruded into a variety of lithologies along three distinct west-northwest trending belts; the Rosia Montana Belt in the north, the Zlatna belt in the centre and the Certej Brad belt, which hosts the Certej deposit, in the south. The South Apuseni Mountains are a range of mountains developed just to the north of the main Carpathian Chain.

9.1. Regional Geology

The geological formations of the South Apuseni Mountains were formed as an accretive geological terrain relating to the collision of the African and European tectonic plates and the subduction of the African plate which resulted in the development of the Carpathian Mountain chain in Central Europe. Lithologies comprises Mesozoic, shallow marine and non-marine sedimentary rocks overlying Palaeozoic and Precambrian sedimentary and metamorphic basement. North-directed thrust faulting during the late Cretaceous resulted in a series of nappes. After this, collision and subsequent rotation of the Pannonian microplate occurred with the European continental plate (Figure 9-2). This was accompanied by northeastward subduction, including roll-back and slab detachment of the subducted material resulting in partial melting, in the Late Eocene to Lower Miocene. This resulted in volcanic activity and, combined with graben style extension in the overlying plate due to dextral rotation leading to tearing in the accretionary wedge, allowed the emplacement of numerous calc-alkaline volcano-intrusive complexes. The three west northwest trending volcanic belts of the South Apuseni are the manifestation of this graben controlled volcanic activity. The structural framework developed during the microplate rotation also provided the fluid pathways for the high-level gold-silver mineralisation and porphyry copper deposits of the Golden Quadrilateral. Tertiary volcanism in the area has been subdivided into three cycles. The earliest cycle is interpreted as middle Tortonian (11.6 to 7.2 Ma) age and comprises andesitic volcanics and rhyolitic ignimbrite overlain by andesitic and rhyodacitic volcanics. Volcanogenic sediments occur throughout this cycle and widespread hydrothermal alteration overprints all rock types. Rocks of the second cycle outcrop the most extensively and are characterised by andesite and dacite overlain by a very thick sequence of quartz andesite that is, in turn, overlain by pyroxene andesite. The sequence is interpreted to be late Tortonian. The middle and upper sequence of this cycle represents the principal host to gold-silver mineralisation currently being mined in Romania, as well as significant occurrences of copper, lead, zinc and mercury.


The third and final cycle of volcanism continued into the Quaternary era and is characterised by pyroxene andesite, basaltic andesite and potassic basalt. The Certej Complex represents part of the southernmost belt of the three major west-northwest-trending belts of volcanism and associated mineralisation identified within the Golden Quadrilateral (Figure 9-1). This graben hosted volcanic belt is traversed by north-south trending transfer faults that and may also act as channel ways for mineralising fluids (Figure 9-3). While movement along these transfer faults is difficult to determine due to the lack of good marker beds, it seems to have been predominantly vertical to sub vertical in nature.

9.2. Local Geology

The Brad-Sacaramb Neogene volcanic belt occurs within a major northwest-southeast trending graben which is approximately 20 kilometres long and 10 kilometres wide. The basement in the Certej area on either side of the graben has been mapped as Basalt of Jurassic age. The basalts are commonly spilitically altered on extrusion, resulting in replacement minerals such as carbonate, albite and chlorite. Cretaceous sediments within the graben are quartz rich and intercalated with black shales and interpreted as Barremian-Aptian (130 to 120Ma) age. Neogene conglomerates, grits, marls and grey shales overlie the Cretaceous sediments. The Certej licence is dominated by extrusive and intrusive andesite. Magnetics show a distinct textural difference between the andesites to the east and those in the western half of the tenement. Andesites stocks throughout the area are commonly propylitically altered and surrounded by lenses of extrusive andesite, which blankets the surrounding Neogene sediments. The south-western quarter of the tenement comprises a thicker sequence of Neogene sediments. Andesites and associated extrusives are the main host for gold-silver mineralisation in the concession with the exception of the Certej deposit (Figure 9-4). At the Certej deposit the sedimentary packages are cut by Neogene porphyritic andesites of both intrusive and extrusive origins have been divided into the ‘Hondol’ and ‘Sacaramb’ types. The main characteristics of which are both are of similar age (K-Ar dating gives around 11Ma) and composition except the Sacaramb Andesite contains biotite and is coarser grained. The Baiaga Andesite is interpreted to postdate the Hondol and Sacaramb Andesites and is of similar composition but is typically fine grained. All lithologies are commonly brecciated from crackle type breccias in the andesite to milled polymictic breccias which contain all rock types. Thrusts and conjugate faulting are the dominant structural elements in the license area. The main structural control of the pre-Miocene basement was thrust faulting. While very little of the basement is exposed within the license area, possible east-west trending thrusts have been interpreted from airborne magnetic survey and are interpreted to have exerted some control on the distribution of younger rocks. The main structural control on the distribution of the Miocene andesite intrusives and volcanics is the intersection of the older EW thrusts and later conjugate faulting, which was coeval with


the andesite volcanism. The main fault trends are approximately NNW-SSE, and WNW-ESE. While movement along these faults is difficult to determine, due to the lack of good marker beds, it seems to have been predominantly vertical. The Hondol and Dealul Grozii Andesites are interpreted to have been emplaced at the intersection of an older EW thrust and younger NW faults. The area between the andesites consisted of Cretaceous shales overlain by Neogene sandstones which was subsequently buckled upwards with the intrusion of the younger Baiaga Andesite.


Figure 9-1: The Golden Quadrilateral


Figure 9-2: Regional Tectonics


Figure 9-3: Geology of the Certej Brad Belt


Figure 9-4: Geology of the Certej Concession


10. DEPOSIT TYPES

The deposits along the Certej Brad belt are classified as low to medium sulphidation epithermal deposits. The Certej deposit itself is mid- to shallow-level and has been driven by hydrothermal fluids probably of a mainly magmatic origin, possibly from a porphyry style system at depth, though given the regional scale of the structural setting, deep crustal fluids cannot be ruled out. The deposit itself is a structural trap with distinct chemical and physical characteristics that have controlled the deposition of metals, as outlined in the next section. Epithermal gold silver deposits such as Certej are found in association with accretive geological terrains associated with subduction zones along plate boundaries such as that found in the Apuseni Mountains, as outlined in the previous section. Other examples of these accretive terrains hosting epithermal deposits can be found along the Tethyan Arc as well as in the Andes and Rocky Mountain chains of the Americas and in the Indonesian Archipelago and all around the Pacific Rim. Resources have also been estimated in existing dumps at the Certej site. To accommodate the proposed Certej pit shell designed footprint, it is a requirement that significant volume of the existing north and south Coranda dumps be relocated. Reverse circulation (RC) exploration drilling in 2004-2005 targeted selective areas of the dumps that were believed to be historically higher in gold and silver grade. This drilling did return encouraging results and a second phase of exploration took place in third quarter of 2008. This phase targeted the remaining dump areas that had not been tested with RC drilling and surface channel sampling. The objective of the RC and channel sampling was to generate a global polygonal estimate from the north and south dump. Ultimately, the exploration programmes aim was to identify additional ROM feed that could potentially be economically placed through the Certej Processing Plant, after all in-pit reserves have been exhausted. It was decided to estimate the resource quantities of the entire dump as total extraction represents a more realistic option than attempting to high-grade the deposits which would involve complex grade control.


11. MINERALISATION

The Certej gold deposit is hosted by Cretaceous and Miocene sediments and Neogene andesites. The deposit is situated within a within a 2.5km by 1.5km magnetic low which is bounded by unaltered biotite andesite to the west, north and east and at least partly bounded by the Notandreas fault trending WNW-ESE to the south. The magnetic low is due to magnetite being replaced by pyrite during hydrothermal alteration. Within the magnetic low significant gold mineralisation has been outlined in a 1500m long zone extending approximately east-west from Dealul Grozii to the NNE-trending Lidia Fault. Within the zone north-northwest trending structures, typified by breccias, host disseminated pyrite mineralisation. An East-West section showing the interpreted geology is detailed in Figure 11-1. A generalised account of the mineralising events that led to the Certej deposit is given in Table 11-1.


Table 11-1: Mineralising Events Event Comment Structural Preparation 1. Hondol (West) and Dealul Grozii (East) Andesites emplaced along major east-west structure representing a reactivation of early thrusts within the Certej-Brad graben. Development of breccia zones in sediments and along andesite margins.

Occurs as regional Certej-Brad graben opens due to rotation of the Pannonian microplate and partial melting of subducted material.

Structural Preparation 2. Intrusion of Baiaga (Central) Andesite buckles sediment package. Sediments and early andesite contact zones are compressed and brecciated to produce north-south trending structures. Pervasive silica alteration of the sediment packages and continuing upward movement of the Baiaga Andesite leads to brittle fracturing of the Cretaceous sediments.

Occurs as regional Certej-Brad graben opens due to rotation of the Pannonian microplate and partial melting of subducted material.

Hydrothermal system driven by deep crustal source and / or intrusion at depth leads to propylitic, argillic and silica alteration focussed at Certej by channel ways from earlier intrusion and deformation. Silica and gold/silver associated with fine grained, disseminated pyrite deposition occurs in pressure shadow above the Biaga ridge. Results in elongate east-west zone of gold mineralisation but with mineralisation hosted within the north south structures. Hydrothermal brecciation of all rock types occurs. Final part is arsenic / gold rich resulting in arsenical pyrite rims in the eastern zone.

Has resulted in general fine grained gold (+/- silver) mineralisation, particularly disseminated in breccias and sediments and gold in pyrite. Late arsenic-gold stage has resulted in arsenical rims to pyrite grains in the east of the deposit.

Later stage of hydrothermal activity produces Pb-Zn rich fluids, which fill brittle north trending fractures, and some reactivation of the gold, which overprints earlier mineralisation. Last part of event included tellurium and silver rich fluids and is concentrated in the central part of the orebody due to shifting pressure shadow as area. rotates and the other structures now being blocked by silicification.

Reflected in late stage Pb-Zn veining and Te, Ag rich veining in the Central zone

Late movement along the NW-SE faults has remobilised mineralisation along with emplacement of the late stage quartz veining with galena, sphalerite, pyrite and barites which can also host higher grade gold mineralisation.

The orebody has been separated into four zones; East, Intermediate, Central and West. With the exception of the Intermediate zone, these were originally designated by the dominant andesite present but later work has shown that each has its own mineralogical characteristics. Details of each zone are given below:-


West (Hondol) The Hondol area is broadly characterised by the intrusion of the Hondol Andesite and includes the Hondol Carol and Coranda Mica zones. The surface geology of the Hondol Carol and Coranda Mica prospects is dominated by an intensely altered intrusive andesite. The altered andesite overlies a sub-vertical west dipping wedge of mineralised sediments has been compressed between the Hondol and Baiaga Andesites to the east of Hondol. The wedge of sediments is typically brecciated and is mineralised at depth and represents the fluid pathway for the western mineralisation. The west mineralisation shows a mixed signature for the multi-element ICP work completed to date. Variable arsenic and tellurium concentrations show overprinting of mineralising fluids in the west. Numerous higher grade NW-SE veins and mineralised andesite stocks also occur in the west zone at Certej. Central (Baiaga) The Central (Baiaga) zone includes the Coranda Open Pit and the area to the northwest of the pit. Drilling has shown that the Baiaga Andesite forms an approximately east-west trending ridge with its highest point on the west side of the pit. The andesite has intruded and deformed Neogene and Cretaceous sediments and pushed aside overlying Neogene andesite. The Cretaceous and Neogene sediments are draped over the east-west trending ridge of Baiaga Andesite and plunge steeply to the north and south and to the west and east over the keel of Baiaga Andesite. The ICP work demonstrated the central zone to have low As and elevated Te concentrations associated with Nagyagite a Pb-Te-Sb-Au telluride which is associated with higher gold grades in the central area. Higher grade Pb-Zn mineralisation is also concentrated along the margin of the western and northern flanks of the Baiaga Andesite in this zone at depth. East (Dealul Grozii) The Dealul Grozii area is located to the east of the Coranda Open Pit. A strongly potassic and silica altered andesite of similar composition to the Hondol andesite (the Dealul Grozii Andesite) intrudes into or is faulted adjacent to Cretaceous sediments. The andesite has also intruded along the Cretaceous/ Neogene contact and has been deformed by the intrusion of the Baiaga Andesite. The andesite hosts the bulk of the eastern mineralisation associated with fine grained pyrite with arsenic-rich rims which contain higher gold than the cores of the pyrites. The andesite is commonly brecciated along its margin and the sediments to the west of the andesite are commonly brecciated by both mechanical and hydrothermal breccias. An Intermediate zone exists between the Central and East zones and is essentially a transition zone between the two. Mineralisation styles identified at Certej within the zones are as follows:

• Disseminated and stockwork/breccia hosted gold mineralisation. • Vein-hosted base metal (with gold and silver) mineralisation.

11.1. Alteration

An extensive zone of strong hydrothermal alteration hosts the Certej deposit. Different hydrothermal alteration packages have been mapped, dependent on the host lithology, as follows:-


Cretaceous sediments:-

• Argillic alteration, quartz, adularia with associated barite.

Badenian (Neogene) conglomerates:- • Quartz-adularia-pyrite and intense silicification.

Hondol-style Neogene andesites:- • Argillic and propylitic alteration, with adularia, silica and clay minerals.

Sacaramb/Baiaga-style andesites:- • Strong argillic alteration.

The distribution of alteration assemblages is complex, but falls into two main groupings:- • Clay-sericite-pyrite (‘argillic’) assemblages, which generally occur peripheral to

the core zones of gold-silver mineralisation. • Silica-adularia-pyrite-sericite (‘silicic/potassic’) assemblages, which usually

represent the core zone of the Certej deposit.

In addition, chlorite-carbonate (propylitic) alteration assemblages are regionally developed within the andesites, but have been overprinted by the argillic and silica alteration assemblages in the Certej Deposit. The destruction of primary magnetite and replacement with pyrite in the alteration halo surrounding the Certej deposit, results in a marked magnetic ‘low’ being associated with the mineralisation. An East-West section showing an east-west section of the interpreted alteration is detailed in Figure 11-2.

11.2. Deposit Zonation

As stated above, the deposit has been divided into four zones, East, Intermediate, Central and West, based on the dominant rock types, mineralisation styles and geochemistry. In order to define the zones a total of 873 five metre composite samples were assayed for Te, Ag, As, Cd, Fe, S and Sb using ICP/MS and OES at the Ultratrace laboratory in Perth during 2006. A probability plot was used to define the population distribution of the element values in the ICP samples. This allowed an interpretation of anomalous values to be applied to the plots of the elements. Plots showed the distribution of both Te, As and Ag demonstrated zonation in the deposit and these plots are shown in Figure 11-3. This work was used to define the metallurgical sampling and the following table indicates the main characteristics:-


Table 11-2: Deposit Zonation Ore-Zone Dominant

Lithology Au g/t

Ag g/t

S:Au ratio

Te ppm

As ppm

Reserve Au Metal

content % East Andesite

Neogene 50% 22%

2.26 5.4 1.67 3.4 1052 30

Intermediate Cretaceous Breccia

48% 28%

2.2 10.5 2.16 14.5 576 20

Central Andesite Cretaceous Breccia

22% 39% 22%

1.78 18.2 1.96 25.4 232 23

West Andesite 85% 1.82 12.4 2.31 3.2 434 27

The arsenic distribution defined from the ICP work shows higher arsenic values in the East due the presence of the arsenic rich pyrite rims on the pyrite. The arsenic level is a mixed signature in the Intermediate Zone reflecting a zone of mixing between the East and Central Zones due to the intense brecciation. The Central Zone is low in arsenic and the West Zone is mixed.

The tellurium distribution shows an inverse relationship to the arsenic distribution with low tellurium in the East with the highest values in the Central Zone. This also reflects the mineralogical work which detected tellurides in the Central Zone flotation concentrate. The Cretaceous sediments are interpreted to be a chemical trap for the tellurides. The sediments occur mainly in the Central Zone and to a lesser extent in the Intermediate zone in the form fine grained matrix in breccias. This explains the moderate levels of tellurium found in samples from the Intermediate zone. Tellurium levels are low in both the East and West Zones which are dominated by andesitic rocks. The tellurides are interpreted to be a later event in the Certej deposit as the silica alteration in the andesites will have impeded the fluid flow in these rocks during the deposition of the tellurium enriched fluids, forcing them in to the sediments and breccias. The silver distribution also reflects the telluride distribution with abundant stutzite (Ag telluride) noted in the central zone. Thus the silver distribution tends to mimic the telluride distribution with low silver in the East and highest in the Central Zone.


Figure 11-1: Certej West-East Geological Section


Figure 11-2: Certej West-East Section - Alteration


Figure 11-3: Certej Deposit Zonation


12. EXPLORATION

Exploration data used in the resource calculation relate entirely to the exploration carried out by the current owner. Previous exploration is described below for information only.

12.1. Exploration by Previous Owners

Since the 1970s, exploration at Certej has continued under the control of the Romanian state companies S.C. Minexfor S.A. (Minexfor) and Regia Autonoma a Cuprului Deva or Minvest (Regia Deva), and has consisted of the following key methods:-

• Underground development (adits, drives and cross-cuts). • Underground diamond and short hole rotary drilling. • Channel sampling of existing and new underground development. • Surface rock chip sampling. • Limited surface diamond drilling.

12.2. Exploration by Current Owners

The exploration under the Deva Gold ownership can be split in to three phases as follows.

12.3. Exploration managed by RSG Global – 2000 to 2003

Exploration was initially conducted by Deva Gold and then from June 2000 to the end of 2003 was managed on behalf of Deva Gold by consultants RSG Global. The programme during this period comprised the following work

• Underground channel sampling of all accessible underground drives and cross cuts.

• Surface channel sampling. • Limited blast hole sampling and rock chip sampling was carried out in the existing

open pit on the Central zone. • Surface drilling programme conducted between January to October 2001. • Exploration data gathered from the underground and surface channel sampling

and the drilling programme provides the basis for an independent resource estimate by RSG Global in November 2000 and this was updated by RSG in January 2002 (see section 8).

12.4. Involvement of Other Consultants – November 2003 to October 2004

Exploration from November 2003 to October 2004 and was managed by Deva Gold and consisted of surface RC and diamond drilling. This period also saw the involvement of consultants Kvaerner and Minproc. Kvaerner re-examined the deposit and remodelled the resource in an attempt to define high grade zones based on a series of north-south sections. Whilst the sections cut the overall east-west trend of the orebody, this did not recognise the fact that most of the gold mineralisation is hosted by north-south structures and so the resulting calculation understated the resource with an unclassified estimation of some 25Mt at 2.1 g/t Au. Given the flawed nature of the resource estimation, little can be drawn from the Kvaerner work that followed. Minproc used the RSG resource estimate proposed a bulk mining approach of 8Mtpa using gold recoveries


by direct cyanidation of 56% and 63% for silver based on averages of the afore mentioned metallurgical sampling results. This did not prove to be viable at the $350/oz gold and $4.5/oz silver prices used by Minproc.

12.5. Management by European Goldfields Technical Team

From October 2004 European Goldfields’ management changed to the current team which included for the first time experienced technical staff. On re-examination of the deposit data and reports it was concluded that, whilst much of the exploration data gathered by RSG was of a high quality, neither the deposit controls, particularly higher grade cores to the mineralisation, nor its mineralogy and therefore metallurgy were fully understood. In November 2004 a drilling campaign was commenced that continued in to 2005 with an emphasis on tighter drill spacing in higher grade areas in order to demonstrate that a selective mining unit (SMU) model as applied by RSG in 2002, was appropriate and that the change of support by uniform conditioning was a valid approach. In places the drill spacing was reduced to 20 metres and this indicated that the SMU model had correctly predicted where larger parent blocks contained high grade sub-blocks and so the approach was applied again in the 2005 estimate. The programme of work from October 2004 also emphasised the need for representative metallurgical sampling. In order to complete this, a full understanding of the mineralogy was required. Elements of the deposit zonation were known at this stage such as the presence of Tellurides in the Central area and the elevated arsenic due to arsenic rich pyrite rims in the East area but these had not been mapped accurately. Drill pulps were therefore analysed using ICP multi-element techniques in order to map the characteristics and from this the deposit zones described in the sections above were delineated. Representative samples of core from each zone were then identified to produce composite samples from each zone.

12.5.1. Exploration of Existing Dumps

Exploration over the Coranda dumps took place in two phases, with the first phase between 2004-2005, targeting higher-grade oxide material and mineralised Cretaceous sediment with vertical RC drill holes. A second exploration phase in 2008, covered the remaining accessible dump areas with vertical RC drill holes and channel lines.

The programme of exploration over the dumps can be summarised as:

• 2003 to 2004 - systematic grab sampling, trenching and hand auger programmes over selective parts of the dumps. Results from this programme are not included in this polygonal estimate

• 2004 to 2005 – RC drill holes in dumps, totalled 56 holes for 1,088m (DRSD030 – 042 and DRSD082 - 089). Depth of holes ranged between 3-39m and average sample recovery of 9.56kg per 1 metre of RC

• 2004-2005 – Internal polygonal estimate of 0.68Mt @ 1.28 g/t Au & 18.85 g/t Ag (Indicated) on a 40m x 40m grid


• 2008 – RC drill holes in dumps, totalled 103 holes for 1,799m (DRSD102 – 146 and DRSD175 – 232). Depth of holes ranged between 3-37m and average sample recovery average of 9.40kg per 1 metre of RC

• 2008 – Channel sampling lines along berms and toes, totalled 112 channel lines for 7,438m (WDSC001 – 112). Channel interval of 2m and average weight of 3.68kg per 2m sample.

The exploration activities over the Coranda dumps between 2004-2008 includes:

• 232 RC drill holes (4 1/2“ diameter hole) for 2,587m, average of 9.40kg per 1 metre of RC

• 114 channel lines with 2m sample intervals for 7,438m, average weight of 3.68kg per 2m sample

• 1 channel line with 10m composite intervals for comparative work for 220m (not included in this polygonal estimate)

• In-situ volume for north and south wireframe is 2,141,307m3 and 2,735,383m3 respectively, with a cumulative in-situ total of 4,876,690m3.


13. DRILLING

13.1. Drilling of Deposit

Drilling and sampling statistics from the exploration programmes described above and used in the current resource estimate are presented in Table 13-1 below. No drilling information prior to European Goldfields involvement in the project has been used in the estimation of resources due to the lack of quality control when the data was collected. Table 13-1: Drilling and Sampling Statistics

Exploration Programme

Method Number Average Length

Total Metres

Number of Assays

1 (July 1998 - May 2000)

DDH 0 - - - RC/DDH 0 - - - RC 0 - - - Surface Channels 9 45 405 376 Underground Channels 74 96 7103 5873

2 (June 2000 to May 2003)

DDH 1 245 245 245 RC/DDH 42 350 14680 14540 RC 120 160 19170 18822 Surface Channels 161 65 10487 9527 Underground Channels 49 67 3276 2836

3 (June 2003 to December 2004)

DDH 53 219 11595 10326 RC/DDH 10 287 2871 2520 RC 16 236 3772 2956 Surface Channels 33 30 982 460 Underground Channels 0

4 (January 2005 to August 2007)

DDH 52 104 5395 4471 RC/DDH 10 181 1805 1775 RC 13 70 905 896 Surface Channels 4 31 123 114 Underground Channels 0

Sub - Total

DDH 106 163 17234 15042 RC/DDH 62 312 19355 18835 RC 149 160 23847 22674 Surface Channels 207 58 11997 10477 Underground Channels 123 84 10379 8709

Combined All drill holes All channels Combined

317 330 647

191 68

128

60436 22376 82812

56551 19186 75737

13.2. Drilling of Existing Dumps

A gridded drill spacing of approximately 40m x 40m was set up over the dumps, however access was not always possible to every collar and adjustments were made at


times to accommodate berms and large boulders on the dumps. As dumps are not homogeneous, variations in drill hole collars placements will not affect the integrity of the polygonal estimate. In addition to 2008 drilling, channel sampling lines were completed along all accessible berms and dump toes. Sampling was typically done about half way up the berm, and along selected toes, as well as a selection of transverse channel lines from an upper to a lower bench. The channel lines provide a comprehensive assessment of the upper dump surface.

Table 13-2 RC and Channel Summary

2004 DRSD030 - 042 RC 13 321 281* 1mRC2005 DRSD082 - 089 RC 8 82 69* 1mRC

DRSD102 - 118DRSD175 - 202WDSC001 - 029 SCS 29 2,764 1,382# 2mSCS

Totals: 95 4,056 2,537

DRSD043 - 046DRSD050 - 068

2005 DRSD090 - 101 RC 12 168 142* 1mRCDRSD119 - 146DRSD203 - 232WDSC030 - 112 SCS 86 4,674 2337# 2mSCS

Totals: 179 6,269 3,773

Total Metres

Total Sample

s

Comments

CERTEJ CONCESSIONRC Drillings/ & Surface Channels 2004-2008

Year Site ID Type No. of Hole Collars/Chan

nel Lines

1mRC805*

NORTH DUMP

2008RC 45 889

SOUTH DUMP

2004 RC 23 517 443* 1mRC

RC2008

58 910 851* 1mRC

*Variance between total metres drilled and total number of samples submitted for analysis is due to poor recovery and some RC intervals are missing. #Variance between total channel metres and sample number submitted for analysis is due to 2m intervals collected and some RC intervals are missing. RC = Reverse circulation sample & SCS = Surface channel samples.


14. SAMPLING METHOD AND APPROACH

Both reverse circulation (RC) and diamond drilling (DDH) methods have been used to define the Certej deposit (Figure 14-1). Sample weights have been routinely measured for RC drilling on a metre by metre basis as part of the standard RC drilling and sampling procedures. The resultant data has been statistically analysed. Theoretical percentage recoveries have been based on calculating the volume of a hole produced using a 5.25 inch hammer and using a density of 2.4t/m3 (deposit average) to determine the theoretical sample weight. On this basis, the theoretical weight of a metre sample is 33kg. RC samples were then sent to the laboratory for drying and pulverising to 85% passing -75µm. A 50gm charge is then taken for fire assay analysis by standard industry techniques. Full details on the RC sampling and assaying procedures are in the Certej Technical Report by RSG dated October 2007. Table 14-1 summarises the result of the statistical analysis and the calculated theoretical percentage sample recovery for the RC drilling programme based on assumed densities and therefore weights for each of the rock types. High sample recoveries by industry standards for RC have been determined, due to the routine use of ‘blow-backs’ during the course of drilling. Table 14-1: Theoretical Sample Recoveries RC

No. Minimum Weight (kg)

Maximum Weight (kg)

Mean Weight (kg)

Median Weight (kg)

CV % Recovery

28,861 0.05 56 27.7 30 0.3 83

A combination of surface diamond drilling has been completed. The drilling comprised HQ core diameter (66% of metres drilled) and NQ core diameter (34% of metres drilled).

To ensure a high sample quality, stringent data collection quality control procedures were applied. The standard quality control procedures for the diamond drilling programs included:-

• Use of triple tube sample collection in areas of non-optimal ground conditions. • Use of sub-three metre core runs in areas of non-optimal ground conditions. • Use of specialised drilling mud in areas of non-optimal ground conditions. • Recording of core recovery (see below for analysis). • Photography of all core prior to sampling.

The diamond core was marked off at 1m sample intervals and cut lengthways using a diamond saw to produce half-core samples for assay. The first 27 (RC and diamond) drillholes of Exploration programme 3 were sampled on a 2m basis. (Note: The 2m samples were subsequently re-fire assayed on a 1m basis for mineralised intervals over 0.5g/t Au for the 2005 resource estimate). Each metre sample of half HQ drillcore was bagged into a separate calico bag. The first half of the 2m composite was labelled with the sample number and an “A” to denote the first half of


the composite and the second half of the composite was annotated with a “B”. The individual metre samples were then crushed and pulverised at the Gura Rosiei laboratory and exactly 150g of pulp was measured from each mill and combined to form the 2m composite sample. The individual metre pulps were stored and the residues discarded. Field duplicates were produced from the same half-core following jaw crushing. Drill core recoveries were calculated by comparing the measured length of recovered core with the distance recorded on the core blocks between each drill run. Diamond core recoveries are routinely recorded as part of the standard geological logging practise. The vast majority of diamond core is highly competent and recoveries of 100% are the norm, and rarely fall below 95%. Table 14-2 below demonstrates that very high core recoveries have been achieved. Table 14-2: Summary Calculated Diamond Drilling Recoveries No. Minimum (%) Maximum (%) Mean (%) Median (%) CV 25,773 1.0 100.0 95.91 100.0 0.13

14.1. Dump Sampling Approach

The surface channel and down hole RC sampling method and approach used in the two phases of exploration over to the dumps was a continuation of Coffey Mining (formally RSG Global) procedures documented earlier in this section. The Deva Gold sampling team in both 2004-2005 and 2008 sampling programmes was made up of an experienced team that have been provided with the appropriate training to complete all aspects necessary to undertake the sampling in a diligent manner and minimise contamination. All sampling is supervised by experienced Romanian geologists. The distribution of sampling is shown in Figure 14-2.

14.1.1. Surface Channel

Channel samples were collected from the existing dumps surface along a pre-determined line. The horizontal two-metre sample intervals were measured out with a tape and the start and end of each interval was marked with paint. The distance along the channel line was also recorded. A canvas sheet was then placed below the interval line and a trench was systematically dug between the two paint marks. The excavated material was then transferred from the canvas sheet into a sample bag and routinely weighed. The transverse channel samples were collected from the side of a typically one metre high face cut into the dump. The cut in the dumps was achieved by a dozer blade cutting a track down the dump faces. The same sampling principle was applied and paint marks were placed on the dump face. As the samples were collected the geologist on site recorded the lithology of the sample and other reportable information. On completion of the sampling a surveyor recorded the start and end of each sample interval using a total station and calculate a dip and azimuth for the interval. In this way, the channel line could be treated as a drill hole in the database and could be


accurately plotted in 3D space and enable the intervals to be captured into the polygonal estimate. Results of the comparison and validation exercise are described in detail in sections below.

14.1.2. Reverse Circulation

The drilling in the dumps is entirely RC, however a limited number of diamond resource pre-collar RC intercepts are included in the wireframes. These holes were drilled as part of the exploration of the main Certej Deposit. RC samples were collected at one metre intervals and split with a Jones riffle splitter. The bags of cuttings were routinely weighed prior to taking the sub-sample via the splitter. Field duplicates were also taken via the splitter every 20 samples and assigned the next consecutive sample number. All RC samples have been weighed on a metre by metre basis as part of the standard RC sampling practice. All the RC samples have been collected in a robust and appropriate manner following the stringent procedures set out for resource estimation studies and detailed earlier in this section. Theoretical percentage recoveries of samples have been based on calculating the volume of a RC hole using a 4 ½ inch hammer with <5mm of space surrounding the hammer, using a density of 2.4t/m3 (deposit average) to determine the theoretical sample weight. On this basis, the theoretical weight of a metre sample is 27kg. As we are drilling dumps a void factor of 0.7, has been applied to compensate for this. This provides an approximate average weight of 19kg. Later Sections discusses bulk density and void factor in more detail. Samples larger than 19kg have either been drilled through an area that has a lower void factor of 0.7, or contain additional material that has been stripped from the hole.


Figure 14-1: Distribution of Channel, Diamond Drillhole and RC Sampling – Plan View


Figure 14-2 RC Drillhole Collars and Channel Line Locations Over the North and South Coranda Dumps.


15. SAMPLE PREPARATION, ANALYSES AND SECURITY

Prior to March 2005, all drilling and channel samples were sent to the Gura Rosiei SGS laboratory in Romania. Following March 2005, drillhole and channel samples were submitted to both the Gura Rosiei SGS laboratory in Romania and the Krumovgrad Bulgaria, SGS exploration sample preparation facility, and assayed at the SGS Chelopech laboratory in Bulgaria. Gura Rosiei (Romania) Samples were prepared and assayed at the Gura Rosiei custom-built assay facility managed by SGS until 2006 and then owned and managed by ALS Chemex, both internationally accredited assay laboratory groups. The Gura Rosiei assaying from 2000 to 2004 was monitored by umpire assays completed by Bondar Clegg, ALS Chemex and SGS. RC and channel samples were dried and the whole sample pulverised to 95% passing 75µm in a LM5 Mixermill. Core samples were crushed in a jaw crusher prior to the pulverising stage. A 50g split was submitted for assay for gold by fire assay followed by an atomic absorption spectrophotometry (AAS) finish. Silver, copper, lead and zinc were routinely analysed by multi-acid digest followed by an AAS finish. The RC samples were prepared and collected from the drill rig and delivered to the onsite office on a daily basis. Similarly, the diamond core is transferred to the core yard adjacent to the onsite office on a daily basis for processing and sampling prior to submission. The entire procedure is undertaken by Romanian national geologists and field technicians, and is closely supervised by expatriate geological personnel. Samples are delivered to the Gura Rosiei laboratory by Deva Gold staff. The rapid submission of samples from drilling for analysis, and the close scrutiny of procedures by Romanian and expatriate technical staff, provides little opportunity for sample tampering. Equally, given the rigorous umpire assaying via external international laboratories and the regular ‘blind’ submission of international standards to the primary and umpire assay facilities, any misleading analytical data would be readily recognised and investigated. Reference material for all samples is appropriately retained and stored, including chip trays derived from RC drilling, half-core and photographs generated by diamond drilling, and assay pulps of all submitted samples. Assessment of the data indicates that the assay results are generally consistent with the logged alteration and mineralisation, and are entirely consistent with the historical and anticipated tenor of mineralisation. Chelopech (Bulgaria) Samples were prepared at the custom built Krumovgrad Bulgaria, sample preparation facility, managed by SGS and assayed at the custom-built assay facility in Chelopech Bulgaria, managed by SGS, an internationally accredited assay laboratory group. Senior staff from Deva Gold and European Goldfields have visited both facilities during May


2006 and found both to be operating above the requirements of a commercial laboratory. The Chelopech assaying from 2006 onwards was monitored by umpire assays completed by both ALX Chemex and SGS. RC and channel samples were dried and the whole sample pulverised to 90% passing 75µm in a LM5 Mixermill. Barren quartz sand is routinely used to clean the LM5 mixermills every 20 samples routinely and additionally on request if higher gold grades were anticipated. Samples were routinely checked for pulverisation at a rate of 1 in 20. A 100gram scoop was sieved and the coarse reject weighed to determine if the pulverisation passed 90% at 75µm. Additional random checks of the coarse rejects are performed by the laboratory manager. For each batch of samples an initial four samples are pulverised at varying times and sieved to check the residence time required is the same. Core samples were crushed in a jaw crusher prior to the pulverising stage and a barren material was routinely used to clean the jaw crusher every 20 samples. A 50g split was submitted for assay for gold by fire assay followed by an atomic absorption spectrophotometry (AAS) finish. Silver, copper, lead and zinc were routinely analysed by multi-acid digest followed by an AAS finish. Additionally a full 300gram pulp is taken and returned to Deva Gold for future reference. The RC samples were prepared and collected from the drill rig and delivered to the onsite office on a daily basis. Similarly, the diamond core is transferred to the core yard adjacent to the onsite office on a daily basis for processing and sampling prior to submission. The entire procedure is undertaken by Romanian national geologists and field technicians, and is closely supervised by expatriate geological personnel. Samples are delivered to the Krumovgrad facility laboratory by a transport contractor employed by Deva Gold. The samples were cleared customs in Deva and delivered directly to the Krumovgrad facility where the customs seal is broken ensuring the chain of custody is complete. The rapid submission of samples from drilling for analysis, and the close scrutiny of procedures by Romanian and expatriate technical staff, provides little or no opportunity for sample tampering. Equally, given the rigorous umpire assaying via external international laboratories and the regular ‘blind’ submission of international standards to the primary and umpire assay facilities, any misleading analytical data would be readily recognised and investigated. Additionally both laboratories have security guards and the pulps are stored in a locked cage inside a secure building at the Krumovgrad facility, before transport by SGS staff to Chelopech. Reference material for all samples is appropriately retained and stored, including chip trays derived from RC drilling, half-core and photographs generated by diamond drilling, and assay pulps of all submitted samples. Assessment of the data indicates that the assay results are generally consistent with the logged alteration and mineralisation, and are entirely consistent with the historical and anticipated tenor of mineralisation.


15.1. Dump Sample Preparation, Analysis and Security

Samples were prepared and assayed at SGS Gura Rosiei and ALS Chemex Gura Rosiei, assay facility managed by SGS in 2004-2005 and ALS Chemex in 2008. Both are internationally accredited assay laboratory groups. RC & channel samples were dried to between 110-120oC and then the whole sample was pulverized to 95% passing -75 μm in a LM5 mixer mill. A 50g charge was taken from the bowl and submitted for gold fire assays followed by an AAS finish and silver, lead, zinc, copper were analyzed by multi-acid digest followed by an AAS finish. The implemented sampling - assaying procedure flow sheet is presented in Figure 15-1. The precision and accuracy of the assay data was continually assessed by the use of standards, field and laboratory duplicates, and internal laboratory standards. Deva Gold included internationally accredited Rocklabs gold-silver standards that were submitted at a rate of between 5-10% of all samples (1:20) laboratory repeats on preparation (1:25 – 4-5%) laboratory repeats on assays (as required) laboratory blanks (1:50 – 2%) and internal round robin samples on a monthly basis. The laboratory QAQC is considered to be extremely robust and today, so obvious issues exists. The independently submitted quality control samples, the internal laboratory quality control data and the inter-laboratory verifications are assessed in detail below.


Conventional Assay Procedures and Quality Control at DevaGold

Sampling - Assaying QA/QC

Surface Channel Samples2 meter

(m = 5 kg)

Reverse Circulation Drilling Samples1 meter or 2 meter composites, riffle split on site (to m = 3-4 kg)

1/20th re-sampled parallel to original sample<QA/QC - Field Duplicates>

SGS Analabs / ALS Chemex* Laboratory- Gura Rosiei, RomaniaSamples are racked on trolleys in order by following the submission sheet.

Sample numbers entered in computer. CCLAS / LIMS* program generates paperwork, creates replicates and splits.

Oven –S001, S002 / DRY21*All samples received are dried at 110-120 C°

LM5 Disc Pulverize – S022 / PUL – 21b*All samples are pulverized for minimum of 5 minutes. 95% passing through a 75 um screen. A 300 gram scoop is taken from the bowl and placed in a

labeled paper bag (pulp). Sample pulps are kept for storage and rechecks. <QA/QC –External Duplicates>

The rest of sample is placed back in its original bag (residue),after final report are discarded or used for screen fire assay.

Randomly (5%)* are created replicates. A replicate is a sample taken from the LM5 bowl and assayed twice <QA/QC-Lab Replicates>

1/10th are generated second splits, same sample taken again from the LM5 bowl, placed in a separate bag and assayed. <QA/QC-Lab Splits>

Base Metal Analysis – A108 / AA47*0.4 g from pulp bag weighted, placed in test tube, aqua regia digested, made to volume,

and analyzed by AAS.

RCField Duplicates :

1/20th sent to Analabs

Screen - F642, F642T500 g sample is wet screened through

a 75 um sieve. The plus fraction (sample in sieve), and the minus fraction (sample

that passed through), are separatelycollected.

Screen FireCalculation - F644

The plus fraction results and the average result of the three minusfractions are combined. CCLAScomputer program automatically

calculates the final grade.

ReportAssays are reported by Lab as sif files. Assay data (including QA/QC Data) is

stored in Certej GBis Relational Database.

* LabDuplicates: 1/20th reanalyzed by Analabs

* LabSplits: 1/10th reanalyzed by Analabs

* LabReplicates: 1/10th reanalyzed by Analabs

* External Duplicates: 1/5th sent to Chemex, Vancouver

Independent – internationally accredited,randomly selected Au - Ag Standards

and Blanks included (1/20th ) in both Analabs and Chemex batches

Laboratories internal QA/QC data (InternalStandards, Blanks, Rechecks)are stored.

Samples are submitted to laboratory.Sample numbers and dispatch numbers are entered in Certej Relational Database

* ALS Chemex from September 2006

If Screen FireAssay

is required

Fire Assay – F650, A10850 gram charge from the pulp bag. A normal fire has a batch of 50 assays. It consist of 40 samples, 4 replicates, 3 splits,

2 standards, and a blank. Samples mixed with flux (lead oxide).Fusion at 1100°C. Cupellation at 1000°. Prill (Ag-Au) placed in test

tube, aqua regia digested, made to volume, and analyzed by AAS.

Fire Assay – AA26*50 gram charge from the pulp bag. A normal fire has a batch

of 50 assays. It consist of 45 samples, 2 replicates, 2 Lab standards, and a blank. Samples mixed with flux (lead oxide).Fusion at 1100°C. Cupellation at 1000°. Prill (Ag-Au) placed in test

tube, aqua regia digested, made to volume, and analyzed by AAS.

Geostats Sample and Assay Monitoring Service and the PTP-MAL Proficiency Testing Program for

Mineral Analysis Laboratories (Canada)

Figure 15-1 Conventional assay procedures and quality control at Deva Gold for gold and base metal analyses


16. DATA VERIFICATION

The current quality control procedures include the submission of internationally recognised standards, umpire assaying at two internationally recognised laboratories in Canada (ALS Chemex) and Australia (SGS), duplicate and replicate sample analyses and the submission of RC field duplicate samples at a rate of 1:20, with the latter providing a comparison of the total sampling and analytical error. Additional quality control employed by Deva Gold for external monitoring of the precision and accuracy of the assay analyses include:-

• Submission of internationally accredited gold and silver standards, produced by Rocklabs of New Zealand, routinely inserted into the sample stream at a frequency of 1 in 20 routine exploration samples.

• Routine collection of duplicate RC drill sample splits at a frequency of 1 in 20 routine samples.

• Routine collection of duplicate channel samples at a frequency of 1 in 20, approximately 20cm above the primary channel sample location.

• Collection of a duplicate sample split after jaw crushing of trench and core samples at a frequency of 1 in 20.

16.1. Accuracy

The accuracy of the gold and silver assay data and the potential for cross contamination of samples during sample preparation has been assessed based on the assay results for the laboratory internal standards and blanks and the Deva Gold submitted Rocklabs standards. RSG Global considered the gold and silver standards analysed by the SGS Gura Rosiei and Chelopech laboratories are accurate and appropriate for resource estimation studies. The following standards and blanks datasets have been used to assess the accuracy of the SGS Gura Rosiei gold and silver assay data:-

• Gold assay data for blanks and 6 internal laboratory standards routinely analysed during Exploration Programme 2 and 4 internal laboratory standards routinely analysed during Exploration Programme 3.

• Gold assay data for 5 Rocklabs standards submitted by Deva Gold throughout Exploration Programme 2 and 8 Rocklabs standards submitted during Exploration Programme 3.

• Silver assay data for blanks and 2 internal laboratory standards routinely analysed during Exploration Programme 2;

• Silver assay data for 5 Rocklabs standards submitted by Deva Gold throughout Exploration Programme 2 and 4 standards submitted by Deva Gold throughout Exploration Programme 3.

The results of the statistical analysis of the blanks and standards analysed by the SGS Gura Rosiei facility can be summarised as follows:-

• Blanks report gold and silver assays at or near the detection limits and hence, there is no evidence of cross contamination of samples.


• Gold and silver assays of the laboratory and Deva submitted standards are within ±10% of the expected standard values.

• There is no evidence of significant bias in gold assays for any of the analysed standards.

• There is evidence of a negative bias in the silver assays of up to -9.6%, for standard S3, however this is not seen as material and for the majority of the standards the bias is well within industry acceptable levels.

16.2. Assay Precision

The precision of the gold, silver, copper, lead and zinc assay data for the Certej deposit has been statistically assessed based on the following comparative sample/data types:-

• Duplicate RC and trench samples collected in the field - allows assessment of overall precision, reflecting total combined sampling and analytical errors (field and laboratory).

• Duplicate splits of core and trench samples after jaw crushing - allows assessment of overall laboratory precision, reflecting sample preparation and analytical errors at the laboratory.

• Duplicate pulp splits of samples after pulverisation - allows assessment of laboratory precision inclusive of sample splitting and analytical errors after sample pulverisation.

• Repeat analyses of laboratory pulp samples - allows assessment of laboratory analytical precision (exclusive of dominant sampling errors).

• Primary versus umpire laboratory analyses of duplicate pulp splits - allows assessment of inter-laboratory precision and relative accuracy inclusive of sample splitting and analytical errors after sample pulverisation.

The order of the comparative data types listed above reflects the successive removal of sampling and sample preparation error thus allowing the precision associated with each stage in the sampling, sample preparation processes and the sample analyses to be assessed. Details of the available datasets and results of the statistical analyses are summarised below, while a full compilation of statistical plots of the comparative datasets is in the Certej Gold Silver Romania Technical Report by RSG. Statistical analysis of the gold datasets has considered only the assay data greater than or equal to 10 times the SGS analytical detection limit (i.e. data at or above 0.1g/t Au). Similarly, a lower selection threshold of 5 times the SGS analytical detection for silver (i.e. data at or above 5g/t Ag) has been used for statistical analysis of the silver assay datasets. Data has been grouped based on collection method, including surface and underground channel samples, RC samples collected over 1m and 2m intervals, and diamond samples collected over 1m and 2m intervals. The results of the statistical analysis of the routine comparative QAQC gold and silver assay data can be summarised as follows:-

• Increasing levels of precision are reported in relation to each successive sampling stage approaching laboratory analysis for both gold and silver.

• No change in relative error was noted across the grade range (T and H plot) for all sample types for both gold and silver.


• Similar precision levels are reported for the assay datasets for the different input sample types after entry into the sample preparation stream (duplicate crusher splits onwards).

• The lowest precision (±18.9%) is reported for the duplicate (twin) trench samples, reflecting the collection of samples from distinctively separate locations (20cm apart).

• In conclusion, industry acceptable levels of precision are reported for all of the sampling stages for the purpose of resource estimation.

The inter-laboratory precision and relative accuracy between the primary laboratory (SGS, Gura Rosiei) and umpire laboratories (SGS Perth and ALS Chemex, Vancouver) has been assessed based on the primary and umpire laboratory assays of the duplicate pulp splits collected during both exploration programme at Certej. All primary and umpire assay datasets for both gold and silver have been compared and no bias was found.

16.3. Data Verification for the Dump Exploration Database.

16.3.1. Assay Standards

The standards selected for this programme have been in use at Certej Project for many years and are produced by Rocklabs.

In total 327 standards have been included into the two phases of RC and channel sampling and are in-line with the 1:20 ratio outlined in the Deva Gold quality control procedure. The Rocklab standards selected and numbers included in the 2004-2005 and 2008 programmes are presented in Table 16-1 together with their expected Au and Ag values (EV).

Deviation from the expected Au value are within tolerance and no significant issues exist with the reported values of all six standards. The reported Au results are considered to have excellent accuracy and precision. The reported Ag values have a higher variance however, it is not considered to be unacceptable. Overall, the standards results reported by SGS and ALS over the 2004-2008 period, when compared against expected Rocklab values, demonstrate an acceptable analytical accuracy in relation to industry requirements.

Table 16-1 Rocklabs standards submitted to SGS and ALS

Standard Type

Expected Value -

EV (ppm)

No of Analyses

Minimum Maximum Mean Stand Dev.

Coeff. Of variation

Outside of +/-5% of EV as %

Outside of +/-10% of EV as %

Au - SF12 0.819 69 0.77 0.84 0.803 0.018 0.022 4 0Au - SH13 1.315 66 1.21 1.36 1.292 0.031 0.024 6 0Au - S8 1.887 36 1.75 2.05 1.870 0.061 0.033 14 0Au - SK11 4.823 61 4.66 5.07 4.869 0.091 0.019 0 0Au - S6 13.89 22 13.00 14.35 13.816 0.317 0.023 5 0Ag - S8 14.77 36 13 16 14.1 0.826 0.059 56 25Ag -S6 17.08 22 14 18 15.9 0.971 0.061 36 36

Statistical Summary of Au and Ag, Standard data Assays, Reported by ALS


16.3.2. Dump Field Duplicates (RC and Channel)

Duplicate samples are collected in parallel to the original sample split on site at a ratio of 1:20. In total 120 field duplicates have been included in the two phases of RC sampling as outlined in the Deva Gold quality control procedure.

Correlation plots for Au and Ag are presented in Figure 16-1 and Figure 16-2, along with descriptive statistics in Table 16-2 and Table 16-3.

Assay results returned from SGS and ALS laboratories are considered to be above average and correlation coefficients for both Au and Ag are well above minimum industry requirements. There is very limited scatter in Au and Ag and this would be considered acceptable for a mineralised Au and Ag systems. On the whole the repeatability of the field duplicates for Au and Ag are considered to show a high level of accuracy and precision and no significant bias has been observed.

Table 16-2 Comparison Statistics, Au field duplicates >= 0.1 Au ppm.

Au Data Original Dupl.Mean 0.41 0.40

Minimum 0.10 0.10

Maximum 1.18 1.17

Pop Number 105 105

Pop Std Dev. 0.23 0.22

Coeff of variation 0.56 0.55

Corr Coeff. 0.997

Table 16-3 Comparison Statistics, Ag field duplicates >=0.5Ag

Ag Data Original Dupl.Mean 7.3 7.3

Minimum 1 1

Maximum 40 43

Pop Number 109 109



Corr Coeff. 0.992


Figure 16-1 Scatter and Q-Q plots for Au in Dump Field Duplicates Database.

Comparison plot for Au filed duplicates, Au >=0.1 ppm Au original (x-axis [alpha]) vs Au duplicate (y-axis [dup]) Q-Q plot. Datasets have been binned into 10 deciles, Au >=0.1 ppm


Comparison plot for Ag field duplicates, Ag > 0.5 ppm Ag original (x-axis [alpha]) vs Ag duplicate (y-axis [dup]) QQ Plot. Datasets have been binned into 10 deciles, Ag > 0.5 ppm

Figure 16-2 Scatter and Q-Q plots for Ag in Dump Field Duplicates Database.


16.3.3. Laboratory Replicates and Repeats

Replicates are selected at random by SGS-CCLAS (2004-2005) and ALS-LIMS (2008) laboratory computing systems at a frequency between 2-4 samples (2-in 2008) per standard fire assay batch of 50 samples. The replicates are prepared the same way as any other sample and are pulverized for a minimum of 5 minutes, so that 95% of the sample passes through a 75 μm screen. The replicate is a repeat sample taken from the pulp bag and assayed twice. For the purpose of this section, all Au and Ag repeats have been included in this section, as they also represent a second sample taken from the pulp bag and assayed twice.

Correlation coefficient plots for Au and Ag are presented in Figure 16-3, along with descriptive statistics in Table 16-4 and Table 16-5.

Statistics from the replicates do not highlight any concerns, in either the 2004-2005 and 2008 programmes. A calculate >0.99 correlation coefficient for both Au and Ag, confirms the repeatability. No significant bias has been observed.

Table 16-4 Comparison statistics, Au laboratory replicates and repeats

Au Data Original RepeatMean 0.88 0.88

Minimum 0.10 0.10

Maximum 9.97 10.40

Pop Number 375 375



Corr Coeff. 0.996 For Au >= 0.1 ppm

Table 16-5 Comparison statistics, Ag laboratory replicates and repeats

Ag Data Original RepeatMean 14.1 14.1

Minimum 1 1

Maximum 383 357

Pop Number 553 553

Pop Std Dev. 25.93 25.16


Corr Coeff. 0.992 For Ag > 0.5 ppm


0.1

1

10

100

1000

0.1 1 10 100 1000

Dup

licat

e A

ssay

Primary Assay

Correlation Plot, Ag

Comparison plot for Au laboratory replicates and repeats, Au >= 0.1 ppm

Comparison plot for Ag laboratory replicates and repeats, Au >= 0.1 ppm

Figure 16-3 Scatter Plots for Au and Ag in Dump Laboratory Duplicates Database.


16.3.4. Laboratory Preparation Repeats – Splits (for 2004-2005)

Preparation repeats are selected at random by SGS- CCLAS (2004-2005) and ALS-LIMS (2008) laboratory computing system at a frequency on preparation (1:25 – 4%) and are included in a standard fire assay batch of 50 samples. The preparation repeats are prepared the same way as any other sample and are pulverized for a minimum of 5 minutes, so that 95% of the sample passes through a 75 μm screen. The preparation repeat is a seperate sample taken from the LM5 bowl and placed in a separate bag for assaying.

Correlation coefficient plots for Au only are presented in Figure 16-4, along with descriptive statistics in Table 16-6.

Statistics from the preparation repeats are excellent. A calculate correlation coefficient 1.00 for Au, confirms the excellent repeatability. No significant bias has been observed.

Table 16-6 Comparison statistics, Au laboratory preparation repeats (pulp splits)

Au Data Original RepeatMean 0.56 0.56

Minimum 0.09 0.10

Maximum 3.65 3.68

Number 15 15

Std Dev. 0.88 0.89


Corr Coeff. 1.00


0.01

0.1

1

10

0.01 0.1 1 10

Split

Ass

ay

Primary Assay

Correlation Plot, Au

Figure 16-4 Comparison plot for Au laboratory preparation repeats (pulp splits)


16.3.5. Laboratory Blanks

In total 514 blanks were included in the RC and channel sample batches submitted to SGS and ALS, between 2004-2005 and 2008. A total 234 were assayed for Au and 280 for Ag. The outcomes are:

• 180 retuned <0.01 g/t Au >77%

• 52 returned 0.01 g/t Au

• 2 retuned 0.02 g/t Au

• 278 returned <1 g/t Au <1%

• 1 returned 2 g/t Ag <1%

• 1 returned 1 g/t Ag <1%

Au-blanks returned a Std Dev. of 0.002 and no significant variance. The Ag-blanks returned a Std Dev. of 0.10 and no significant variance. Statistics from the blanks are excellent.

16.3.6. Inter-Laboratory Round Robin

ALS Chemex Gura Rosiei participates in the monthly Fire Assay round robin which consists of analyzing 10 samples for gold by fire assay with an AAS finish. During 2008 the average bias was -0.66% and a ranking of 1. Only one failure was recorded in fire Assay. ALS Romania results can be considered to be of a high quality and no areas of concern are observed.

16.3.7. Conclusions Dump QAQC

The findings from the check methods for standards, field and laboratory replicates, repeats, blanks and round robin reported by SGS and ALS, between the period 2004-2005 and 2008 are all well within industry standards for fire assay and AAS.Table 16-7, provides a summary of QAQC of assay data.

The assay dataset is therefore considered to be robust and reliable and of high standard for reporting this polygonal estimate.


Table 16-7 Summary of check assay data from SGS & ALS 2004-2005 and 2008

Summary of Au and Ag Check Assay Data

Au Ag Data

Comparison Total No. of

Analysis Sample Type Confidence

Level at 90% Corr. Coef.

Confidence Level at 90%

Corr. Coef.

Field duplicates 120 RC &

Channel 0.036 0.997 1.127 0.992

Lab replicates and repeats

423 (Au) & 604 (Ag)

RC & Channel

0.102 0.996 1.788 0.992

Laboratory preparation

repeats (Pulp Splits)

15 RC &

Channel 0.398 1.000 - -


17. ADJACENT PROPERTIES

Not Applicable.


18. MINERAL PROCESSING AND METALLURGICAL TESTING

18.1. Introduction

In the early stages, post 2000, of the modern development of Certej the early interpretation was based on the mistaken belief that Certej was similar to the Rosia Montana deposit. In 2004 a new team was put in place at EGL to move the Certej project forward and the first question to address was to determine whether Certej was free milling. It was known that the gold was mostly associated with pyrite. A metallurgical review of all the available information was compiled which highlighted the shortcomings of the earlier work particularly that the samples were not representative and a significant number were likely to have been oxidised resulting in artificially high direct cyanidation gold extractions. A new drilling programme was carried out to ensure the highest degree of representivity across the Certej deposit and over a 100 samples, carefully selected throughout the deposit, demonstrated that the Certej deposit was refractory in nature throughout giving an average gold extraction of only around 26% and with considerable variation.

It became clear that in order to achieve high gold and silver extractions an oxidation step would be needed and this focussed the testwork on floatation to produce a gold containing pyrite based concentrate and the subsequent extraction of the contained gold and silver. Extensive flotation testwork confirmed that high gold and silver recoveries of over 90% could be achieved into a pyrite concentrate grading up to 20g/t gold and 100g/t silver.

In parallel with the flotation work, mineralogical investigations were conducted and it became apparent that the Certej deposit can be considered to comprise four ore zones, the Central, the East, the Intermediate and the West, all of which have different chemical and mineralogical signatures and differing degrees of refractoriness

A desktop study concluded that the best option for liberating the refractory gold and silver was the Albion process on the grounds of achieving high recoveries and competitive capital and operating costs compared to other processes. The Albion process was then evaluated fully with continuous pilot scale work to confirm the selection. Sufficient quantities of concentrates were produced at the company’s laboratory facilities in Greece from more than 2 tonnes of drill core representing the 4 ore zones. Two continuous Albion runs were carried out on a composite of the Intermediate, Central and East ore zones in the proportion that will be mined in the first 8-9 years of operation. Both pilot plant runs achieved gold and silver extractions by subsequent cyanidation of the oxidised residue of over 90% and 85% respectively. The pilot scale work programme also included a continuous CIL campaign on the Albion residue to establish the reagent dose rates and other operating criteria for the CIL plant.

18.2. Mineralogy

18.2.1. Introduction

There have been extensive mineralogical and petrographic studies carried out on the Certej deposit, the recent studies date back to 1993. These have incorporated works by


Romanian institutions and independent world leading consultants including SGS Lakefield and Amtel; a reference list is included in this document.

The deposit at Certej contains rock masses comprising three main types, Andesite/Dacite intrusives, Cretaceous sandstone and Neogene conglomerate and sandstones. Hydrothermal and tectonic brecciation occurs extensively within the deposit. Subsidiary andesite lithology types, spatially distinct within the Certej resource, have been labelled as Baiga Andesite, Sacaramb Andesite and Hondol Andesite.

The Certej resource can be identified broadly by distinct chemical and mineralogical signatures into four main and discrete ore zones. These have been classified as the East, the Intermediate, the Central and the West zones. The gold and silver values are mostly distributed as solid solution and very fine particles within pyrite, in gold bearing minerals locked in pyrite grains, and, for the Central and Intermediate ore zones, associated with complex Tellurium minerals. There is a strong compositional association between gold and arsenic over the resource. The studies identified arsenean pyrite zones within the pyrite grains particularly as rims on the pyrites in the East ore and within the pyrite grains from the West. Arsenopyrite is only a minor gold carrier mainly limited to the East and Intermediate ores, in some studies it was not observed at all. As expected there is a very good correlation between the gold and silver.

18.2.2. Mineralogy Pre 2004

Early mineralogical analysis identified the major sulphide minerals as pyrite, significant sphalerite and marcasite. They concluded that no visible gold was detected and the gold was effectively in solid solution, mainly in arsenical pyrite rims associated with pyrite grains. It was observed that there is no or very little arsenopyrite mineralization.

Some of the associated cyanidation leach tests gave high gold extractions in excess of 80%, depending on the lithology and the sample source. This indicated that the gold is not entirely locked up and that the ore has some free milling characteristics.

The gangue minerals noted were quartz and K feldspar with minor plagioclase and muscovite.

Lakefield obtained a total of 197 electron microprobe analyses from pyrite grains in the Mozley concentrate prepared from the Cretaceous Ore. Approximately 39% of the pyrite analyses reported elevated arsenic contents ranging from 0.99 wt.% to 8.23 wt.%, averaging 2.62 wt.% As, and were termed “arsenical pyrite” in the Lakefield report. Backscatter electron imaging of the pyrite revealed distinctive rhythmic compositional banding relating to regions of relatively low and high arsenic content.

SIMS analysis was performed on a total of 30 pyrite grains from the Cretaceous Ore Mozley Concentrate. Solid solution gold in pyrite was identified in 27 of 32 individual spot analyses. Analyses ranged from less than the instrument detection limit of 0.17 ppm Au to 37.3 ppm Au, with an average of 2.8 ppm Au. Solid solution gold distribution in pyrite was associated with regions of elevated arsenic content, primarily within euhedral (also called zoned), concentric zones on the rims of pyrite grains. In addition, a sub-micrometer-sized inclusion of native gold in pyrite was noted during the SIMS study, indicating the presence of extremely fine-grained native gold in the sample.


The Back-scattered Electron Imaging, BEI, photomicrograph below is typical of the zoning observed in this study. The average gold content of this grain was reported to be 5.39ppm, the bright regions are arsenic rich Plate 18-1: Typical Photomicrograph showing Arsenical Zoning of a Pyrite Grain.

In order to determine how best to characterise the deposit by mineralogy and therefore chemistry ICP data from all the drilling was examined in an orebody mapping exercise. This displayed that the deposit could be divided in to the four zones, West, Central Intermediate and East, by a combination of arsenic, silver and tellurium concentrations. Boundaries between the Central and Intermediate zones were not as well defined though. Concentrate analysis had revealed the importance of tellurium sulphide minerals as gold carriers in the Central part of the deposit and so tellurium mapping was conducted by ICP on composite drill samples. This enabled hard boundaries to be defined between the Central the Intermediate zones with the boundary modelled using tellurium and arsenic concentrations with gold as a guide.

The resulting observations were made following the orebody mapping:-

• ICP plots of As vs. Ag define the boundary between the East and the Intermediate metallurgical domains very clearly, this boundary also agrees with the higher Ag in the model for the Intermediate zone.

• Ag grades in the Block Model define a high grade population in the Central zone and the boundary to the West area is defined where the Ag grades fall away. This is very clear in the Block Model.

• Te plots clearly define the Central zone with elevated Te values compared to the rest of the deposit.

• The Intermediate zone is a combination of moderate As and Te values and thus shows some features of the East and some of the Central zones.

Plots of As and Te are shown on the schematic sections in Figure 18-1: Plots of Arsenic and Tellurium,

18.2.3. Deportment of Gold and Silver in Certej Flotation Concentrates and Tails

This key defining study by Amtel of Canada describes the mineralogy of flotation concentrates from the 4 Certej ore zones and the tailings from three; (the tailings from the Central zone had such low levels in contained value as to make their examination unnecessary). The report is very comprehensive and contains a large amount of information. This study is the standard reference text on the mineralogy of the Certej ore as the samples analysed are the most representative taken, the flotation was carried out well and the concentrate produced made up the feed for the Albion pilot plant continuous testwork. The mineralogical study included optical microscopy, X-Ray diffraction, SIMS and BSE Imaging.

18.2.3.1. Gold Deportment

The material generated for the Amtel studies were produced from drill core which had been scientifically selected using all the considerable information then known about the deposit. This ensured that the material was representative in terms of ore zone


speciation, grade, lithology and of course spatial location. This was not the case for the previous mineralogical studies.

The concentrates arising from the four ore zones are differentiated by individual chemical and mineralogical signatures, the Central is characterised by high Ag, high Te, low As; the East by low silver, high arsenic, low Te, the Intermediate between these levels, (a reflection of the Intermediate’s zone’s physical location between the Central and the East) and the West which has elevated Ag with low Te and As. This is shown in Table 18-1: Flotation Concentrate Composition from the 4 ore zones at Certej.

Table 18-1: Flotation Concentrate Composition from the 4 ore zones at Certej

Ore Zone Au, g/t Ag, g/t Te, g/t S tot, % As, % Pb, %

East Concentrate

18.2 37.5 16.8 35.3 1.00 0.96

Intermediate Concentrate

11.7 63.9 64.8 35.9 0.73 0.82

Central

Concentrate

17.6 174 233 39.5 0.23 0.83

West Concentrate

14.6 102 15.3 40.0 0.49 0.47

The gold in the Certej deposit is found in 3 main forms:

• Gold in solid solution in sulphide minerals mainly pyrite but also, in the East zone ore very minor marcasite and arsenopyrite.

• Gold minerals such as native gold, electrum and Au-Ag tellurides

• Gold bearing minerals such as silver tellurides and nagyagite

By far the dominant form of gold is in pyrite both as solid solution in the 5 forms of pyrite identified in the Amtel work and associated with gold minerals. This explains the refractory nature of the deposit. However provided that the first two mineral types above are exposed to solution, i.e. not locked, then these gold species will be leachable by direct cyanidation and are not refractory. The West ore type contains more solid solution gold in pyrite than the other 3 zones. The Amtel study has estimated the amount of refractory gold by ore type at the test grind (target P80 of ~75microns but actually coarser than this at +/-93 microns) based on mineral species and liberation. This is summarised in the Table 18-2: Refractoriness and Direct Leachability by ore zone.


Table 18-2: Refractoriness and Direct Leachability by ore zone

Ore Zone Estimated Refractory Au, % Estimated Leachable Au, %

East

67 – 84 16 -33

Intermediate 64 -82 18 - 36

Central 40 – 60 40 - 60

West 80 – 93 7 - 20

The Central zone is less refractory due to the higher content of liberated tellurides and gold minerals.

Free and liberated gold minerals are only minor contributors to the overall gold balance for each ore zone.

This extent of direct leachability is also a function of the degree of ore oxidation and this contributed to some of the earlier work leading to the erroneous belief that Certej could be considered to be free milling.

The 5 forms of pyrite present, their gold grade, contribution to the total concentrate sample grade and their approximate make up in the ore zones is shown in the Table 18-3: Solid solution gold carried in sulphide minerals by ore zone. The table also includes the solid solution gold in marcasite and arsenopyrite.


The following observations can be made from the information in the Table 18-3:

Table 18-3: Solid solution gold carried in sulphide minerals by ore zone

Pyrite Morphology Total Marcasite Arseno-

pyrite Coarse Zoned Porous Fine Micro-

Crystalline

East

Concentrate

Abundance,

wt %

17.6 0.8 32.5 2.8 4.5

7.42

3.3 1.46

Au grade,

ppm

0.9 9.9 2.1 33.8 123.7 29.3 190

Au carried, g/t 0.16 0.08 0.69 0.95 5.55 0.96 2.78

RA of Au in Py

form

2.2% 1.1% 9.3% 12.8% 74.6% 100.

Intermediate

Concentrate

Abundance,

wt %

16.42 2.9 40.9 1.4 0.9

1.92

0.3 0.95

Au grade,

ppm

0.4 22.6 1.8 10.2 34.8 29.3 141

Au carried, g/t 0.07 0.65 0.75 0.14 0.3 0.08 1,34

RA of Au in Py

form

3.7% 34% 39.3% 7.3% 15.7% 100

Central

Concentrate

Abundance,

wt %

37.2 0.2 32.8 0.7 0.4

2.70

0.1 0.17

Au grade,

ppm

0.5 1.8 6.8 5.0 61.1 13.6 132

Au carried, g/t 0.19 0.00 2.24 0.04 0.23 0.02 0.22

RA of Au in Py

form,

7.0% 0.0 83.0% 1.5% 8.5% 100%

West

Concentrate

Abundance,

wt %

30.4 0.8 26.0 1.9 13.7

10.01

0.2 0.48

Au grade,

ppm

0.8 105.6 8.8 20.8 45.5 29.3 23.8

Au carried, g/t 0.24 0.82 2.29 0.4 6.26 0.06 0.11

RA of Au in Py 2.4% 8.2% 22.9% 4.0% 62.5% 100%


• Coarse and porous pyrites are the most common forms in all the 4 ore zones.

• The grains of arsenopyrite, microcrystalline pyrite, zoned pyrite and marcasite are the highest in terms of gold grade. The grades vary considerably according to the ore zone.

• Amtel reported that there are no significant differences in pyrite compositional, (Au/As), zoning between pyrites. The East zone is highest in As.

• The lowest gold grade pyrite is the coarse material containing < 1g/t gold. This form of pyrite is the largest component in the Central and the West. The overall grades of these concentrates are increased by the presence of Telluride and gold minerals.

• The West zone mineralogy is variable.

• The porous pyrite type is the most common form in the East and Intermediate and the second most common constituent of the Central and West.

• The microcrystalline is more abundant in the West and the fine grained, (disseminated in rock), in the East.

• Marcasite and Arsenopyrite only contribute significantly to the gold content of the East zone.

• Considering all the forms of pyrite the West at 10.01 g/t and East at 7.42g/t have the highest grading pyrites.

• Allowing for the relative amounts and gold grades the pyrite morphologies containing the most gold are the Microcrystalline for the East and West concentrates and the Porous type for the Central and Intermediate.

The table below summarises the contained ounces of gold in each zone.

Table 18-4: Distribution of Gold by ore zone in Certej

Ore Zone Au , ozs Gold Distribution,%

East Concentrate 636,558 30.0

Intermediate 485,036 22.9

Central 414,769 19.6

West 581,803 27.5

Total 2,118,166 100

The photomicrographs in Plate 7.2 show the five morphological forms of pyrite described by Amtel.

There is a linear relationship between the gold and silver content across the ore types.

The most common form of Te is in the metal/silver/gold/telluride minerals significant in the Central and to a lesser extent in the Intermediate ore zones. These minerals include petzite/stutzite, calaverite, nagyagite and sylvanite.


18.2.3.2. Silver Deportment

The assayed silver contents of the samples of concentrates from the four ore zones are: East – 37.5 g/t, Intermediate – 63.9 g/t, Central – 174 g/t, West – 102 g/t, indicating the silver variability of the deposit.

The silver in the Certej deposit is found in 3 main forms:

• Silver minerals, electrum and Silver Tellurides

• Minerals with silver as an “impurity” in the crystal lattice, tetrahedrite

• Colloidal size silver sulphide in pyrite and solid solution silver in sulphide minerals, galena, Pb sulphosalts and pyrite

By far the dominant form of silver is colloidal size, <0.5 microns silver sulphides in pyrite. Free pyrite, (and minor marcasite), is the principal carrier of silver accounting for 62% of the East and 85% of the West silver grade in the concentrate which is shown in Table 18-5 below.

Table 18-5: The Deportment of Silver in the flotation concentrate

Carrier East Concentrate

Intermediate Concentrate

Central Concentrate

West Concentrate

Free Ag Tellurides >10µm 3.4 8.0 6.6 0.6

Free Ag Tellurides <10µm 9.8 12.1 21.1 3.9

Free Galena+ Pb sulphosalts>10 µm

5.2 3.8 1.5 1.4

Free Galena+ Pb sulphosalts<10 µm

16.2 9.0 3.7 6

Associate Ag/solid solution Ag with free pyrite>10 µm

53.0 51.4 56.4 57.8

Associate Ag/solid solution Ag with free pyrite<10 µm

9.4 14 9.8 28.1

Associate Ag/solid solution Ag with sulphide rock

middlings

1.9 0.8 0.4 1.3

Associate Ag/solid solution Ag with rock particles

1.2 0.9 0.4 0.9

Ag grade from mineralogy, g/t

39.7 63.8 168.6 98.4

Mineralogically accounted for, %

105.7 99.9 96.9 97.2


The Table 18-6 below summarises the contained ounces of silver in each zone.

Table 18-6: Distribution of Silver by ore zone in Certej

Ore Zone Ag , ozs Silver Distribution,%

East 1,561,148 13.0

Intermediate 2,325,217 19.3

Central 4,203,625 34.9

West 3,955,826 32.8

Total 12,045,817 100.

18.2.4. Comments and Observations

• Four ore zones have been identified which make up the Certej deposit and they are characterised by individual chemical and mineralogical signatures, the Central is characterised by high Ag, high Te, low As; the East by low silver, high arsenic, low Te, the Intermediate between these levels, (a reflection of the Intermediate zone’s physical location between the Central and the East) and the West which has elevated Ag with low Te and As.

• The gold is mostly refractory in nature to varying degrees depending on the ore zone with the Central being least refractory and the West zone the most. The gold is mostly in the form of microscopic, sub microscopic and solid solution contained in iron sulphides, predominantly pyrite. The Central zone is less refractory because a significant proportion of the gold is contained in free Gold/Tellurium mineral grains, and this also applies, to a lesser degree, to the Intermediate ore.

• Five morphological forms of pyrite have been identified; Coarse pyrite, which is the principal form in the Central and West and the second most important form in the other two zones, it is low in gold grade with a reported average of 0.63ppm Au; Zoned (euhedral) pyrite which is more abundant in the Intermediate ore but the least abundant overall and relatively high in gold grade averaging 47ppm; Porous pyrite has an average reported grade of 5.3ppm Au and is the principal pyrite morphology in the East and Intermediate ores; Microcrystalline pyrite is significantly more abundant in the West ore and has an average grade of 58ppm and finally fine grained pyrite has an average grade of 19ppm Au and is more common in the East ore zone. The gold rich arsenean rims reported in the East and Intermediate materials indicate that these ore zones should be more susceptible to oxidative extraction methods.

• Organic carbon is present in only minor proportions and will not cause gold losses due to preg-robbing in the gold recovery plant.


18.3. Metallurgical Testwork

18.3.1. Historic Metallurgical Testwork (pre 2005)

Prior to 2005, five significant testwork programmes had been completed; one by AMMTEC and four by SGS-Lakefield Research Ltd (“SGS”), termed SGS Phases I to IV. These programmes included bench scale cyanidation, grinding and flotation, roasting, bioleaching and pressure oxidation investigations. In 2005 a major definitive metallurgical testwork programme was carried out at SGS-Lakefield which is discussed in Section 7.3.3.

The SGS Phase IV programme was a metallurgical mapping exercise on 106 samples taken throughout the deposit. This showed that the deposit was more refractory than previously thought with an average gold recovery by direct cyanidation of only 26%, which is a much lower figure than that used in the Minproc and Aker Kvaerner reports. There was a wide variation in the observed gold recoveries of between 2% and 91% and generally the near surface samples yielded higher average gold extractions attributed to in situ oxidation.

The conclusions of the metallurgical testwork from the pre-2005 metallurgical studies were:

• There is very little free (or gravity recoverable) gold; the majority of the gold is associated with sulphides and is refractory. Some gold is encapsulated with gangue silicates and can not be recovered even after oxidation.

• Flotation recovers around 90% of the gold but the silicate locked gold cannot be recovered.

• Oxidation of a sulphide concentrate, by roasting, pressure oxidation or bacterial oxidation, allows cyanidation to achieve higher recoveries. However, almost complete oxidation of the sulphides is required to obtain significant gold recoveries and the associated high capital and operating expenditure impacts negatively on the project viability.

18.3.2. 2005 Testwork Samples

Inductively Coupled Plasma, ICP, analysis work during this period defined four main zones in the Certej Deposit defined as:-

• The East Zone - characterised by low Ag, elevated As, and low Te

• The Central Zone - characterised by high Ag, high Te, low As

• The Intermediate Zone with Te and As levels between the East and Central zone (a reflection of the Intermediate zone’s physical location between the Central and the East).

• The West which has elevated Ag with low Te and As.

For the 2005 metallurgical testwork programmes carried out by SGS-Lakefield and Hellas Gold laboratories, representative samples were taken from the four distinct ore zones.

The number of samples taken for the 2005 metallurgical testwork is shown in Table 18-7.


Table 18-7: Summary of Total Samples produced for Metallurgical Testwork

Ore-type Laboratory No. of

Samples

East SGS-Lakefield 96

Hellas Gold, Greece 96

West SGS-Lakefield, 61

Intermediate SGS-Lakefield 36

Central

SGS-Lakefield 94

Hellas Gold, Greece 94

Wardell-Armstrong, UK 32

Each sample was taken from a one metre length of quarter drill-core, weighing approximately 1.5 kg/m. The Wardell-Armstrong sample was taken at a later date to check the metallurgical results of the Central sample achieved by SGS-Lakefield and Hellas Gold and locked-cycle flotation testwork verified the findings.

The approximate composition of the material types in the ore-zones are shown in Table 18-8. The gold and silver grades are derived from the geological model and the tellurium and arsenic assays from ICP measurements on selected samples.

Table 18-8: Geological Composition of the Samples used for the Metallurgical Testwork

Ore-type Rock type Alteration Au

(g/t) Ag

(g/t) Te

(ppm) As

(ppm)

East

Andesite Cretaceous Neogene Breccia

50% 13% 22% 15%

Silica Argillic Carbonate Potassic

34% 58% 2% 6%

2.26 5.5 3 1,052

West:


85% 3% 7% 5%

Silica Argillic Carbonate Potassic Propylitic

45% 42% 2% 4% 7%

1.70 11.6 3 434

Interm:


14% 48% 10% 28%

Silica Argillic Carbonate

69% 28% 3% 2.14 13.3 15 576

Central:


22% 39% 17% 22%

Silica Argillic

73% 27%

1.71 21.0 25 232

18.3.3. 2005 Flotation Testwork

Extensive flotation testwork was undertaken on the Central and East zones followed later in the programme by the West and Intermediate material. The aim of the tests was to


produce a bulk sulphide concentrate utilising the knowledge from the previous Certej test work programmes.

After grind size and reagent optimisation were completed, cleaner flotation tests were conducted and finally locked-cycle (“LC”) tests were carried out with either two or four stages of cleaning. In total forty four flotation tests were undertaken. Locked-cycle data is presented in Table 18-9.


Table 18-9: Locked Cycle Test Data

Sample/Facility Area

Feed Grades (%, ppm)

Test

Concentrate Grades (%, ppm) Recoveries

Au [g/t]

Ag [g/t]

As [%]

Sb [ppm]

Te [ppm]

Au [g/t]

Ag [g/t]

As [%]

Sb [ppm]

Te [ppm]

Au [%]

Ag [%]

SGS

Central Zone 2.58 49 0.02 170 48 LC21 24.9 450 0.18 - - 97.1 98.7

LC23 28.8 540 0.19 2,300 440 96.6 97.2

Avg 26.8 495 0.19 2,300 440 96.8 97.9

East Zone 2.73 6 0.19 140 <4 LC22 23.3 47.4 1.57 - - 76.2 82.1

LC24 22.2 52.2 1.48 - - 82.1 90.1

LC26 20.3 46.1 1.37 1,200 17 80.8 89.6

LC27 19.4 42.3 1.31 1,200 17 84.2 79.4

West Zone 2.08 12 0.06 47 <2 LC42 18.7 79.2 0.52 470 6 84.2 84.4

Intermediate Zone 2.38 12 0.08 130 16 LC43 21.3 49.3 0.45 1,500 110 88.4 88.1

(Central/East) LC44 19.8 76.1 0.38 - - 89.5 92.8

Other labs. East (1) 3.16 6 0.17 - - 25.9 49.5 1.30 - - 84.5 73.9

Central (2) 2.08 27 0.01 78 73 25.4 233 0.15 783 731 98.1 97.4

Notes: 1. Hellas Gold Laboratory 2. Wardell-Armstrong International UK


The Central zone ore exhibited the best flotation characteristics with faster flotation kinetics, higher gold and silver recoveries and higher concentrate grades. This is due to a significant proportion of the gold from the central zone being associated with a gold telluride mineral, which floats readily. To verify this, a repeat sample was generated using drill core from the central zone and independently tested at Wardell-Armstrong International, UK, which confirmed the SGS- results.

A weighted average of the locked-cycle test results for each ore-type gave an overall gold recovery of 87.5%, producing a concentrate gold grade of 21g/t.

18.3.4. 2008 Flotation Testwork

Early in 2008 Cytec carried out a series of flotation tests on a composite sample of East Central and Intermediate samples which had been stored at SGS-Lakefield since the 2005 testwork programme. This was instigated by Cytec to assess various collector combinations. The base case parameters were set by EGL and specified a flotation feed size of 120 microns. The results were positive and verified the selection of a coarser flotation feed also concluded in the Amtel mineralogical work.

18.3.5. Preliminary Albion Test Work

18.3.5.1. Description of the Albion Process and Selection Criteria

The Albion Process was developed by Xstrata Technology, formerly Mount Isa Mines Technology, to process refractory precious and base metal sulphide ores and concentrates. The process comprises ultra fine grinding, UFG, in an IsaMill followed by an ambient pressure oxidative leach where the fine sulphide particles are oxidised by oxygen to liberate the values. In the case of gold and silver pyrite concentrates, as is the case for Certej, the oxidised Albion leach residue is then processed by conventional cyanidation in a CIL plant and the extracted gold and silver electro-won and smelted to doré bars.

Many treatment options were considered for processing the Certej concentrate based on bench scale testwork and financial evaluation, refer to Section 7.3.6

18.3.5.2. Albion Batch Testwork

In 2005 and 2006 several batch tests were carried out on samples of the SGS-Lakefield locked cycle flotation concentrates to recover the gold using the Albion process to oxidise the concentrate before gold recovery with cyanidation. The concentrates from the four ore-types were tested separately at a variety of oxidation levels. The results were reported in a series of reports (HRL Technical Memorandum No’s 0806, 0810, 0814, 0827, and 0838). The results of these initial batch tests are represented graphically in Figure 18-2: Sulphide Oxidation vs. Au recovery for Initial Albion Tests.

The conclusions of the early Albion batch testwork were;

• Flotation concentrate from the Central ore domain required lower levels of oxidation

• For concentrates from the East and Intermediate ore domains the gold recovery increased with oxidation

• Concentrate from the West domain did not seem to be amenable to the Albion process


In 2005 and 2006 a total of 22 batch tests were carried out. The final tests were carried out on a blend of Central, Intermediate and East and 92% Au recovery and 83% Ag recovery were achieved at an oxidation level of 83% (HRL Technical Memo No. 0838).

18.3.6. Process Selection

There are a variety of processes to recover the gold and silver locked in the concentrate such as Pressure Oxidation, Bacterial leaching and Roasting.

In order to evaluate the viability of the various processes a desktop study was undertaken in June 2006. The main input factors were the capital and operating costs and the gold and silver recoveries obtained for each process. Table 18-10 summarises this input data. The data used was considered to be realistic at the time.

Table 18-10: Input Figures for Desktop Evaluation of the Optimum Process Treatment Route of Certej Flotation Concentrates

Technique Capex

US$ M

Opex

US$/t conc.

Reference Metal Recoveries Testwork Source

Au % Ag %

Pressure Oxidation $138.2M 42.5 AKES Olympias study 94% 7% AMMTEC report A8025

Bacterial Leaching $61.3M 9.2 Goldfields databse 83% 53% Cepromin testwork 2005,

SGS testwork 2003

Geocoat Process $14.6M 11.0 Geobiotics cost study 75% 75% Estimate

Albion Process $29.0M 36.8 Core Resources ref paper 84% 93% HRL testwork

Activox Process $55M >25 Western Mining 83% 53% BiOx results

Roasting On-site $71.6M 18.4 Lurgi Roaster/ Acid plant 84% 79% AMMTEC A8025 report

Roasting Off-site $0.2M 87.9 Roast at Eti Bor and return 84% 79% As above“

The economic returns were plotted, Figure 18-3: Comparison of Treatment Options; IRR, which shows the most economic treatment process was the Albion process.

18.3.7. Continuous Scale Albion Testwork Programme

18.3.7.1. Bulk Flotation Concentrates Preparation

In Q2 & Q3 2006, 1,856m of NQ core was drilled by the Deva Gold Geological team from the four main ore-zones at Certej to provide 2.66 tonnes of material for flotation and comminution testwork. The samples were carefully selected in terms of geology, lithology, mineralogy and metal grades. Subsequent sulphur mapping showed that the sulphur to gold ratios, (STGR), of the samples also verified the sample validity. These samples are therefore highly representative of the four zones within the Certej deposit and therefore the subsequent investigatory work reflects the behaviour of the entire Certej resource.

Samples were sent to the Hellas Gold test facility as quarter drill core and were prepared for the testwork using industry standard methods. Head samples reported the following grades:


Table 18-11: Hellas Gold Head Sample Assays

Sample Au Ag S Pb Zn As

Central 2.00 25.7 4.87% 0.13% 0.35% 0.02% Intermediate 2.28 11.8 5.77% 0.16% 0.69% 0.14%

East 2.33 4.1 4.49% 0.14% 0.63% 0.17% West 2.00 25.7 4.87% 0.13% 0.35% 0.02%

A total of 22 flotation tests were carried out in Denver 3-litre and 6-litre cells and Outokumpu 50-litre and 1.5-m3 cells. After verifying the flotation conditions on old samples from the 2005 testwork, eight tests were carried out to optimise conditions. The bulk flotation concentrates were produced in the industrial 1.5 m3 Outokumpu cell, refer to the report Preparation of Bulk Flotation Concentrates from Four Samples of Certej Drillcore, Hellas Gold Laboratory.

A total of 250 kgs of bulk concentrate was produced and shipped to HRL comprising;

i. Central : 55 kg

ii. Intermediate : 62 kg

iii. East : 112 kg

iv. West : 21 kg

The flotation results obtained were from batch tests without any recirculation of cleaner tailings which would be carried out in locked-cycle tests. In order to better assess the flotation recoveries expected in a process plant an iterative calculation was carried out with different ‘Cell factors’ applied to the recovery for the re-circulated tailings. A conservative factor was applied and the following recoveries obtained.

Table 18-12: Grades and Calculated Recoveries with Recirculation of Cleaner Tailings

Hellas Flotation Concentrate after Regrinding

Gold Silver Sulphur

g/t % Rec. g/t % Rec. % % Rec.

Central 18.1 95.3 158 94.3 36.0 92.9

Intermediate 14.5 88.4 58.5 87.4 35.6 90.7

East 18.2 88.8 32.3 82.5 34.9 94.2

Composite of

Central, East & Intermediate 17.0 90.3 78.8 90.0 35.5

West 14.5 87.6 105 88.7 41.4 89.2


These results are very close to the Locked Cycle flotation results obtained by SGS Lakefield et al but cannot be compared directly as the material tested here is more representative of the Certej deposit than that used in the earlier programme.

18.3.7.2. Phase I Albion Continuous Test

A composite sample of Central, Intermediate and East from Hellas was prepared according to the proportions of the three domains in the deposit and after some initial batch tests the first continuous Albion pilot test was carried out in December 2006 (HRL Technical Memorandum No, 0849). The continuous leach reactor consisted of four stirred tanks with a total volume of 120 litres. Oxygen flows were controlled manually and limestone addition was automatically added to maintain a pH of 5.5.

The continuous reactor commenced treating the composite sample on the 8th December and concluded on the 15th December. The target sulphide oxidation was 70 – 75% and, from the batch tests, a residence time of 40 hours was targeted for a feed rate of 11.6 kg/day.

Daily samples were taken from all four tanks and submitted for XRD analysis, Au and Ag determination, sulphur speciation and elemental analysis by XRF. The level of oxidation after 9 days continuous operation was 70% requiring 750 kg limestone/tonne and 450 kg of oxygen/tonne of IsaMill feed. Assuming a utilisation of 80% the oxygen consumption would increase to 563 kg/tonne IsaMill feed.

A series of optimisation tests were carried out on a sample of the oxidised residue from day 7 from the continuous reactor. These tests were aimed at optimising cyanidation conditions and to determine kinetic data for design purposes. The use of a lime boil prior to cyanidation had a detrimental effect on metal recovery by cyanidation. The average Au and Ag recoveries of the non-lime boil tests were 91.6% and 84.9% respectively.

18.3.7.3. Phase II Albion Continuous Test

After further batch tests had confirmed that gold recoveries of over 90% could be achieved on the feed-stock the continuous leach reactor was operated on this Certej composite feed from June 13th to June 23rd. The one difference from the Phase I set-up was that the Phase II continuous reactor consisted of 4 x 20 litre reactors with a lower total volume of 80 litres. After 5 days operation, steady state was reached in that a stable sulphide oxidation level of 70% was achieved. The limestone consumption was 751 kg/tonne and the oxygen consumption 550 kg/tonne, (the Phase I consumptions were 750 kg/tonne and 563 kg/tonne respectively).

Batch cyanidation tests were carried out on the discharge slurry from the final 5 days operation of the Albion pilot plant. A further 15 cyanidation tests were carried out to optimise cyanide leach conditions as for Phase I. Neglecting the tests with clear adverse conditions (low CN level and high slurry density) the average recoveries were 89.9% Au and 60.9% Ag. The gold recovery is comparable with the batch testwork and the results of Phase I but the silver recovery is significantly lower.

More test data was obtained describing the relationship between sulphide oxidation and gold recovery. There is a proportional relationship on the composite material with recovery increasing with oxidation until the optimum level of 70% is reached after which the gold recovery levels off.


18.3.7.4. CIL Continuous Tests

A continuous CIL pilot plant was operated for 5 days at an average feed rate of 16.4 kg/day oxidised residue with a residence time of 30 hours. The CIL feed was the Albion oxidised residue from the Phase II test. The average Au recovery was 89% and the Ag recovery 75 – 80%. The CN consumption was 2.53 kg/tonne CIL feed (equivalent to 4.9 kg/tonne of IsaMill feed) and the lime 7.31 kg/t CIL feed.

18.3.7.5. Design Package

The purpose of the continuous Albion and CIL pilot trials was to provide design data to enable an estimate of the capital and operating costs to an accuracy level of +15%/ -15%. A Metsim mass and heat balance for the IsaMill, Albion oxidation and CIL circuits was developed.

18.3.7.6. Vendor Testwork

Samples from the Phase II HRL testwork were sent to Lightnin to size the agitators for the Albion and CIL circuits, (SPX Vendor Report 2007-7734 Xstrata CIL and Albion Agitators) and to Outotec to size the high-rate thickeners used in the circuits for water management, (Outotec Vendor Report Albion Tails 1 & 2, CIL Tails 1 & 2). Xstrata Technology provided an independent report for the IsaMill, (M4 Testwork Report Certej Sept 2007,). Larox carried out filtration testing on the flotation concentrate, (Larox 16714 Xstrata Certej).

18.3.7.7. West Amenability Test

The initial batch tests in 2006 had indicated that the West concentrate would require oxidation levels greater than 80% for the Albion-CIL process to be viable. For this reason the West concentrate was not included in the composite feed for Phases I and II. In 2007 further tests were carried out on the West flotation concentrate.

Two standard batch tests were carried out and the test at the higher oxidation level reported that that an Au recovery of 91% could be achieved at an oxidation level of 74%.

A large batch test was then carried in a 60 litre reactor and small slurry samples were progressively taken, HRL Technical Memo No. 0850. These were then cyanide leached so an oxidation vs. gold recovery profile could be constructed.

Table 18-13: West Concentrate, Sulphide Oxidation / Gold Recovery Profile

Sulphide Oxidation Level

Gold Recovery - % w/w

Silver Recovery - % w/w

43.9% 66.5 59.3 66.0% 84.0 45.7 67.0% 84.5 65.9 69.5% 89.6 65.7 82.5% 89.2 65.6

This data shows that after an oxidation level of 70% the gold recovery levels off, this is similar behaviour to the composite material comprising concentrate from the Central, East and Intermediate ore zones ().


The final residue from the large batch test was subjected to a CIL test and a gold and silver recovery of 90% and 79% was achieved, the latter significantly higher than for the timed samples from the batch test.

18.3.8. Deportment of Gold & Silver in Albion and CIL residues, Amtel, December 2007

During the Albion testwork programme at HRL samples of oxidised Albion residue, C33, and final residue following cyanidation of the oxidised Albion residue, C32, were submitted to Amtel for mineralogical examination. These were taken from selected batch tests which had reported lower than normal recoveries.

18.3.8.1. Gold Deportment

The gold occurrence was described in terms of the following forms and carriers: water soluble gold salts, colloidal size gold (<0.5µm in diameter) encapsulated by gypsum, colloidal gold enclosed in haematite and solid solution gold in residual free pyrite. Gold deportment in the two samples is summarised in the Table 18-14 below.

Table 18-14: Gold Deportment in Albion and CIP tailings

Carriers & Forms of Gold Sample C32, CIP Tailings, post Albion Oxidation

Sample C33 , CIP Feed, Oxidised Albion residue

Gold salts, (encapsulated in gypsum)

7% 10%

Colloidal size Gold, (encapsulated in gypsum)

27% 74%

Encapsulated by elemental Sulphur

6% -

Haematite 39% 10 Pyrite 21% 6

Total 100% - Assay 1.92 g/t Au 100% - Assay 3.1g/t Au

In conclusion it was observed in the samples examined, (which were selected for their poor results), that gold losses to the Albion CIP tails are due to gypsum encapsulation, precipitation of colloidal gold with haematite and un-reacted pyrite bypassing the Albion reactor and/or protected by haematite coatings.

18.3.8.2. Silver Deportment

The silver occurrence was described in terms of the same forms and carriers as gold and are summarised in the Table 18-15 below.


Table 18-15: Deportment in Albion and CIP tailings

Carriers & Forms of Silver

Sample C32, CIP Tailings, post Albion Oxidation

Sample C33, CIP Feed, Oxidised Albion residue

Silver salts, (encapsulated in gypsum)

5% 11%

Colloidal size Silver, (encapsulated in gypsum)

48% 80%

Haematite 27% 6 Pyrite 20% 3

Total 100% - Assay 9.9 g/t Ag 100% - Assay 9.2g/t Ag

Encapsulation by gypsum of cyanidable colloidal size Ag mineral grains and silver on the surface of sphalerite and, to a lesser extent, pyrite particles, was the major cause for silver loss accounting for 48% of the Ag in the C32 CIP tails.

In conclusion, silver losses to the Albion CIP tails were due to the same reasons as gold losses, with gypsum encapsulation being even more detrimental to Ag recovery. The photomicrograph Plate 7.3 shows a Gypsum crystal from Sample C33, Albion residue/CIP feed, encapsulating pyrite and Iron Oxide particles. In the sample of CIP tails these types of gypsum crystals were more abundant and had considerably more inclusions.

18.3.8.3. Summary of Recoveries

The continuous CIL gold and silver recoveries established on the Phase II Albion residue from the concentrate generated at the Hellas Gold laboratory using the 2.66t of drill core are shown in the table below;

Table 18-16: Overall Recoveries from Hellas and HRL testwork

East, Central, Inter. West

Flotation recoveries Au 90.3% 87.6%

Ag 90% 88.7%

Flotation conc. grades Au 17.9g/t 14.5 g/t

Ag 95g/t 105g/t

Albion recoveries Au 90.7% 90%

Ag 74% 79%

18.3.9. Comminution Testwork Programme

In 2005, samples from all four ore-types were tested at SGS for comminution characteristics and the data was used to size two standard Semi Autogenous Grinding (“SAG”) mill and ball mill circuits using JKSimMet, a computer software package used for the analysis and simulation of comminution and classification circuits in mineral processing operations as developed at the Julius Kruttschnitt Mineral Research Centre, Australia. The circuits modelled were SAB, Semi-Autogenous Ball Mill and SABC, as previous but including a pebble crusher for breaking down re-circulating SAG mill discharge screen oversize material. Based on the earlier laboratory flotation test work the target milling circuit product, feed to flotation, size suggested a P80 of 75 microns. Mineralogical studies by Amtel and further flotation testwork, see below, indicated that a


grinding circuit product size P80 of 120 microns provides better downstream metallurgy than the 75 microns originally envisaged.

In 2006 splits from the samples that were processed at Hellas Gold were also sent to SGS-Lakefield for further SAG testwork. A large composite was prepared from the ore-domains and this was subjected to a drop weight test and a continuous MacPherson autogenous grindability test. In addition a further 12 individual samples from the deposit were subjected to SAG Mill Comminution, (SMC), variability tests.

The MacPherson test confirmed the amenability of the ore to semi-autogenous testing. The drop weight test results and the SMC variability tests were used to design the comminution circuit.

Table 18-17: Comminution Circuit Parameters

Circuit Configuration SABC Circuit

SAG BM Overall

Nominal Dimensions 24’ x 9’ EGL 15.5' x 24' -

Internal Dimensions (m) 7.11 x 2.74 4.6 x 7.16 -

Installed Power (kW) 2300 2900 5200

Mill Speed, % Critical VPD 75 -

Throughput (Mtpa) 3.0 3.0 3.0

Avg. Gross Power (kW) 2076 2400 4476

Avg Gross Power (kWh/t) 5.6 6.5 12.1

Avg Product K80 (microns) 1096 116 116

The table shows the configuration for a SABC circuit. The circuit will be operable without a pebble crusher in a SAB circuit configuration with recirculation of any oversize material to the SAG feed. The SAG mill will operate with a ball charge of up to a 15%. The SAG throughput is expected to vary within the range of 295-485 t/h and the final grind will always be 122 microns or finer.

The Amtel 2007 report observed that the mineralogical analysis has indicated that gold and silver recoveries in the flotation plant can be increased by not over grinding to minimise the generation of fines; the suggested feed to float F80 is 120microns. Incorporating a gravity circuit in the grinding plant and using the latest generation of gold selective reagents in the flotation plant will also increase gold and silver recoveries. Space has been provided in the plant layout to include a gravity concentrator in the cyclone underflow stream. Controlling the primary grind and classification circuit is key to avoiding over grinding and thus the generation of fines.

Modern grinding circuits can be controlled to produce a product PSD within a fairly broad range and also to produce a product within a narrow variation as other variables change such as ore hardness and feed rate. A coarser grind will result in lower mill power draws. SGS McPherson re-ran the mill simulation and selection software with a P80 of 120microns and this did result in lower powered primary grinding mills thereby


reducing capital and operating costs although of course the major potential benefit will be improved gold recovery.

Additional independent laboratory test work at SGS Lakefield has confirmed that there is no decrease in gold flotation recovery at a grind P80 of 120 microns.

The base case is thus to design the grinding plant with a product P80 of 120 microns which will be achieved by the mills sizes above. Space will be provided in the grinding plant to retro-fit a pebble crusher and ancillaries in the unlikely event of unforeseen conditions developing.

The concentrate regrind target size is 45 microns which is based on mineralogical and laboratory test work.

18.3.10. Cyanide Detoxification Testwork

In Romania, any surface impoundment of water must have a Weak Acid Dissociable, (WAD), cyanide level of less than 10 ppm CNWAD. To discharge water from an impoundment to the river system the level must be less than 0.1 ppm total cyanide, (CNTOTAL). This latter level is lower than the current EC regulation which is 1ppm CNTOTAL.

Samples of cyanide tailings slurry from the Albion-CIL tests were tested at HRL using an Inco SO2/Air CN destruction technique (HRL Technical Memoranda No. 0827 and 0838). These were successful and CNTOTAL levels of <0.1 ppm were achieved in a single pass

More detailed CN detoxification tests were carried out at SGS-Lakefield (LR 11443, February 2007) using a variety of techniques. The Inco SO2/Air injection system was found to be suitable to achieve a surface impoundment of <5 ppm CNWAD. A large sample of CIL tailings was treated and the solution obtained tested by different techniques. Hydrogen peroxide was found to be the best method to achieve a final discharge of <0.1 ppm CNTOTAL.

18.3.11. Flotation Tailings & CIL residue Percolation Tests

Test samples of a flotation concentrate and tailings produced from the SGS-Lakefield locked-cycle testwork were tested at SGS for settling rates and a hydraulic conductivity test to determine their permeability for TMF design (LR 11675, Sept 2007). The low hydraulic conductivity showed that the sample had was effectively impervious.

A sample of CIL tailings from the Phase II Albion continuous trial was also sent to SGS-Lakefield (LR 11675-001 Rpt2) for identical hydraulic conductivity testwork and again the sample tested had an impervious relative permeability.

18.3.12. Certej Process Plant Summary

The Certej plant is designed to produce a gold-silver flotation concentrate from a disseminated and breccia hosted gold-silver mineralisation at a total treatment rate of 3 million tonnes of ore per annum. The average feed grade for Certej over the life of the mine is expected to be 2.0g/t Au and 11.4g/t Ag. The flotation concentrate will then be oxidised by the Albion process, (a combination of Ultra-fine grinding and atmospheric oxidation), and the gold and silver recovered as doré in a conventional CIL circuit as outlined in the block flow sheet in Figure 18-6, Figure 18-7 and Figure 18-8.


18.4. Plant Description

18.4.1. Overall Description

The process plant has been designed to process 3.0Mtpa of gold bearing ore and will comprise the following principal elements:

• Single stage crushing of run of mine ore using a jaw crusher.

• Semi Autogenous Grinding (“SAG”) followed by ball milling.

• Flotation to produce a gold-silver concentrate. The circuit will comprise rougher flotation, three stages of cleaning with a regrind mill and a 1st cleaner scavenger.

• Concentrate dewatering.

• Ultra-fine grinding of the flotation concentrates to 80% passing 9 microns in an M10000 IsaMill.

• Oxidation of 70% of the sulphides in the Albion circuit. This comprises 5 Continuous Stirred Tank Reactors, CSTRs, with injection of oxygen and pH control through the addition of milled limestone.

• Leaching of the gold and silver in a CIL circuit.

• Recovery of the gold and silver to doré after stripping the carbon in an AARL elution circuit, electro-winning and smelting of the cathodes

• Cyanide detoxification of the CIL tails to <5 ppm CNWAD, by the Inco method of SO2/Air injection. In exceptional circumstances further detoxification of the TMF’s combined overflows will be carried out in a separate detox plant and will ensure that any discharged water contains less than 0.1 ppm total cyanide, CNTOTAL, as per current Romanian regulations. (This level is lower than the current EC limit.)

The general processing route is outlined in Figure 18-6, the production of flotation concentrate in Figure 18-7 and the treatment of this flotation concentrate to recover doré in Figure 18-8. Figure 18-9 is a General Arrangement drawing showing the layout of the plant and the TMFs.

18.4.2. Crushing

The crushing section is located near the plant and incorporates the ROM pad, a primary jaw crusher, and a primary crushed stockpile. ROM ore is generally fed to the primary crusher by direct tipping from the mine haul trucks.

In normal conditions, the crushing plant will operate for three shifts. The crusher has a design capacity of 434 t/h and the crushed ore stockpile has a nominal live capacity of 6,800 giving 18 hours capacity for the SAG mill.

18.4.3. Comminution

The SAG and comminution work undertaken by SGS-Lakefield determined the size of mills and motors required for the plant throughput rate of 3.0Mt/a and a product size with a P80 of 75 and 120 microns. The latter coarser size of 120 microns is the Mesh of Grind (MOG) that the comminution circuit will operate at. The chosen SAG mill will have a nominal size of 7.32m diameter x 2.75m Effective Grinding Length and the ball mill,


5.33m x 7.92m. The SAG motor will be rated at 2.5 MW and have variable speed control and the Ball Mill will be driven by a fixed speed 3.5 MW motor. Pebble ports of 70-mm will remove pebbles to the screen oversize conveyor which will be re-circulated to the SAG mill feed. Ground slurry from the SAG mill is discharged through a grate discharge with a 38-mm opening onto a classifying trommel. The SAG mill trommel undersize will flow to the discharge sump from where the combined SAG and Ball mill product will be pumped to a cluster of 5 x 500mm hydro-cyclones for classification with 4 normally being fed. The circuit will be designed to give cyclone overflow slurry with solids P80 of 120 microns and 41% solids w/w.

The flotation concentrate regrind mill will be a tower mill (Metso Vertimill 650-VTM) using 12mm media and driven by a 500kW fixed speed electric motor.

The 1st Cleaner tailings and the regrind mill discharge will be pumped by one running, with one standby variable speed pump, to a cluster of five 380-mm diameter hydro-cyclones, normally four in operation with one on standby. Cyclone overflow will gravitate to the cleaner-scavenger flotation bank and the cyclone underflow will feed the regrind mill.

18.4.4. Flotation Circuits

The flotation section of the concentrator will be fed by gravity flow from the primary Ball mill hydro-cyclone cluster overflow. The use of a single stream or open-circuit will be adopted in the flotation section i.e. open circuit flow on the rougher bank with no recirculation from the cleaners back to the roughers.

All the flotation machines will be tank cells with individual air supply and electrically controlled air valves. The rougher flotation bank will comprise 5 x 130-m3 cells and will have a nominal volume of 650-m3 with level control dart valves between alternate cells. The cells will be installed in a step-wise fashion providing controlled gravity flow between adjacent groups of cells.

The rougher concentrate will be collected in a series of launders and pipes, and will report to the 1st Cleaner bank of 4 x 30m3 cells. The product from the 1st Cleaner will report to the 2nd Cleaner bank of 4 x 20m3 cells and then to the 3rd Cleaner bank of 3 x 20m3 cells. The 1st Cleaner tails will report to the regrind mill discharge sump with the ground mill outlet stream. The regrind mill product, the 380-mm cyclone overflow, will report to the cleaner-scavenger bank of 4 x 30m3 cells. The cleaner-scavenger flotation concentrate will be re-circulated back to the 1st cleaner feed. The cleaner-scavenger tailings will combine with the rougher flotation tailings and will be pumped to the tailings thickener. The cleaner circuit wil be optimised during plant operations to accommodate any changes in ore characteristics.

18.4.5. Grinding & Flotation Circuit Control

The SAG mill is equipped with load detection by bearing pressure and sound detection. An expert system has not been included at this time on the grounds of cost, but it can be added in the future. The SAG mill inlet dilution water is ratio controlled to the SAG mill fresh feed rate. The sump dilution water is controlled by the cyclone inlet pulp density and the sump level is controlled by the pump speed. The cyclones are equipped with


actuated valves. The particle size of the cyclone overflow is measured by an Outotec PSI 300 On-Stream Particle Size Analyser.

All the flotation cells groups have controlled air flow and level control. All concentrate and tailings sumps are variable speed controlled by the sump level.

The regrind cyclone feed density controls dilution water flow rate to the cyclone feed sump while the sump level is controlled by pump speed. The cyclone inlet pressure is monitored through the plant DCS.

18.4.6. Flotation Concentrate and Tailings Handling

The gold and silver bearing final concentrate slurry from final stage cleaner flotation cells will be pumped to the concentrate thickener feed box and then to a 10m diameter high rate thickener.

The thickener underflow density will be controlled at a nominal 60% solids by weight for feed to a stock tank and then to the Albion Process IsaMill circuit where it will be diluted to 45% solids by weight.

Combined flotation tailings will be pumped directly into the thickener feed launder. A high-rate thickener with a diameter of 17-m has been selected for this study.

Flotation tailings from the process plant at a density of 60% solids will be pumped to the Tailings Management Facility. The clear water from the thickener overflow will be pumped back to the process plant recycled water tank.

18.4.7. Description of the Albion Process Plant

The Albion Process is a method to treat refractory sulphide minerals to recover the valuable metals contained. The process consists of ultra-fine grinding of the mineral or concentrate, followed by oxidative leaching at atmospheric pressure in stirred tanks. The flow sheet utilises existing commercially proven technology in its unit operations. The process was developed by MIM Holdings (now Xstrata) and a former MIM subsidiary, Highlands Gold, to treat refractory base metal and gold ores. The technology has been tested extensively at pilot plant scale and is currently being evaluated for use in several gold mining projects throughout the world including Envirogold’s Las Lagunas gold Project in the Dominican Republic who were the first to obtain an Albion Process Technology license. A total of 57 operations have tested the Albion process at HRL and 11 have proceeded to full pilot plant trials.

The process to recover gold and silver takes place in three major steps;

• Ultra-fine grinding to less than 9 microns in an IsaMill.

• Oxidation in stirred tanks. Oxygen is added as the oxidising agent. The pH and Eh are carefully monitored, by the controlled addition of limestone.

• Recovery of gold from the oxidised pyrite by cyanidation and standard refining.

18.4.8. IsaMill Grinding

This is the first step in the Albion process and is where the flotation concentrate is reduced in size to a P80 of +/-9 microns. One important factor of the operation of the IsaMill is that the product size distribution curve is closely classified with little over grinding. This has been demonstrated in many plants worldwide. The largest is the


M10,000 unit (10,000 litre capacity) with a 3.0 MW variable speed drive motor which is in use at Centerra Gold’s Kumtor Gold Mine, Kyrgyzstan and at Anglo Platinum’s Western Limb Tailings Re-treatment project in South Africa. There were plans for three more units to be installed in 2006/7.

The IsaMill is a horizontal high speed stirred mill that achieves power intensities of up to 350 kW/m3. Inside the shell, rotating grinding discs are mounted on a shaft which agitates the grinding media and ore particles. The media that is used in the mill is generally ceramic beads or silica sand. The former is more expensive and is generally only used when there is a requirement to maximise the throughput for a particular application. The M10,000 unit selected for Certej is sufficiently large so that silica sand may be used but until a suitable sand media is located near Certej the IsaMill will operate with ceramic media.

18.4.9. Albion Oxidation Plant

Refractory gold bearing sulphides, such as pyrite, liberate both iron and acid when oxidised. In the alkaline process, this iron and acid is continually neutralised and precipitated from the leach by the addition of limestone. The continual removal of the leach reaction products from solution means that the leach progresses rapidly and very high levels of sulphide oxidation can be achieved. Ultra-fine grinding of the minerals to minus 10 microns or less prevents passivation by these precipitated leach products, as the leaching minerals are consumed before a sufficient layer of precipitates can form.

Oxygen is injected into the pulp at a controlled rate to achieve oxidation and neutralization of the sulphates produced is achieved by the addition of finely ground limestone which is added via a ring main.

The general leach reaction is:

FeS2 + 15/4O2 + 9/2H2O + 2CaO = FeO.OH + 2CaSO4.2H2O

The degree of oxidation is variable and from the testwork to date an oxidation rate of 70% has been determined as the optimum level. This means that 30% of the pyrite entering the Albion leach circuit will exit in the form of a sulphide.

The chemistry involved is well understood and the key to ensuring that the reaction is maintained is to monitor and control the pH (acidity) and the Eh, redox potential. This will be achieved by pH and Eh monitors and oxygen measurement units in each tank which will control the oxygen and limestone addition rates.

The reaction takes place in Continuous Stirred Tank Reactors, (CSTR), and to achieve a flow approximating plug-flow five CSTR units are selected. Each will have a diameter of 11.5-m and a height of 14.9-m and a live volume of 1,227-m3. They will be powered by 110 kW dual impeller agitators. This gives an overall residence time at normal throughput of 27 hours as determined by the test work programme.

The Albion leaching process will be carried out at about 97°C to achieve the best kinetics, the temperature will be developed by the oxygen and limestone associated reactions and tank cooling is supplied by air which is added at a controlled rate to the reactor tanks. Oxygen is sparged through each tank at 3,170Nm3/h through 25 Slamjet


SLJ75 oxygen spargers at 4.8barg. Cooling air is introduced through lances from the top of each tank.

The oxidised slurry exiting the final Albion tank will be cooled in a column before proceeding to the solid/liquid separation step and CIL.

18.4.10. Albion Tails Thickener

The Albion residue will be thickened in a 20-m diameter high-rate thickener. The recovered water will be recycled to the Albion tanks to compensate for the evaporative losses and the water consumed in the oxidising, exothermic reactions.

18.4.11. Gold Recovery Plant

This is a standard type CIL plant which is in use throughout the world. The only difference is that it is designed to treat a high grade residue with characteristics slightly different from a conventional free milling type CIL feed stock. These minor differences are:

• A high viscosity and ultrafine slurry which will increase the power required to the agitators by about 25%

• A high reactivity and leach rate which reduces residence time accordingly

• Higher cyanide and lime consumption rates than industry normal values by approximately 100%, although this is very variable and depends on many physico-chemical factors.

The leach/CIL plant will be relatively small requiring 30 hours retention time. Although the grade of the flotation concentrate is around 17-18 gpt gold and 80-100 gpt silver, due to the formation of reaction products during the Albion leach the incoming solids grade will average 8 g/t Au and 43 gpt silver.

The CIL comprises a series of six carbon steel agitated tanks in series. Each tank has a live volume of 800m3

, is 10.1m diameter, has an overall height of 11.0m and a residence time of 5 hours. The slurry pH will be adjusted to pH 10.5 by the addition of lime from a ring main. Each tank is equipped with a dual impeller 75kW agitator with both impellers rubber lined hydrofoils. Oxygen is added as air by lances entering from the top of the tanks and discharging below the bottom impellers.

Each CIL tank will be equipped with an interstage carbon screen which will retain carbon while the slurry will flow via pneumatic dart valves and launders from one CIL tank to the next. Carbon will be transferred counter-current to the pulp flow by carbon interstage pumps. The carbon pump deliveries will be screened by wedge wire screens to avoid the slurry returning to the previous tanks in the drain. By-pass launders will be installed to by-pass any of the CIL tanks for maintenance. Cyanide solution will be fed to the CIL tanks with the cyanide dosing pumps.

Carbon will be withdrawn from the first and second CIL tanks by carbon transfer pumps and pumped to the loaded carbon screen where the slurry will be washed from the carbon. Screen undersize material will flow back into the same CIL tank, while the carbon collected on the screens will gravitate into the loaded carbon storage tank.


Regenerated and conditioned fresh carbon will be fed to the last or second last CIL tanks in the train. Slurry leaving the end of the train will flow to the carbon safety screen via the CIL tailings sampler.

18.4.12. CIL Tailings Thickener

The CIL tailings will be thickened in a high-rate thickener. This will provide cyanide-rich water for the pre-leach tank (reducing the need for additional water), reducing fresh cyanide consumption and reducing the operating cost of the cyanide destruction circuit. This has been sized at 17-m diameter.

18.4.13. Elution and Carbon Regeneration

The elution section will treat carbon from the CIL circuit to remove gold and silver and is based on the widely used AARL system. A single 11-tonne column has been selected which can operate on 2 elutions per day, 6 days a week during periods of high silver production. The nominal design is for slightly more than one elution per day. The gold and silver loading capacity of the activated carbon has been established to be 8kg/tonne.

Loaded carbon collected on the loaded carbon screen will fall through a loaded carbon sampler and into the loaded carbon storage tank. Carbon is pumped from the surge tank into the acid wash column.

Acid washed carbon will be transferred from the acid wash tank to the elution column by pressurising the former vessel with transportation water. The Certej elution circuit is based on the AARL system and will be operated at 130o C under pressure of 4.3 bar. Water for elution is supplied from the raw water supply tank. Elution solution will be pumped by the eluate circulating pumps through the elution column.

Before entering the column the solution will be heated in two stages: firstly in the elution recovery heat exchanger where the hot solution leaving the column will pre-heat the entering solution. The temperature of the elution solution will be raised to the required temperature using an LPG fired thermal fluid elution heater.

The first stage of cooling will be effected in the elution recovery heat exchanger, as mentioned above. The eluate will then gravitate to the electro-winning section.

Eluted carbon will be fed into the electrically heated horizontal regenerating kiln which can treat 1 tph. The carbon will be reactivated at a temperature of 750oC. Reactivated carbon will discharge from the kiln into the quench tank. Water from the transportation water tank will be used for quenching. The regenerated carbon pump will pump the carbon from the quench tank onto the regenerated carbon screen.

18.4.14. Electro-winning and Smelting

A standard system of electro-winning will be used to recover gold and silver from solution. Elution solution will flow from the eluate cooler to the flash tank where it will gravitate to two banks of two electro-winning cells inside the Goldroom where the gold and silver will be deposited on stainless steel cathodes. Spent electrolyte will flow back to the strip solution tank. Loaded stainless steel wool cathodes will be stripped manually in the cells using a high pressure water lance.

The sludge from cathodes and the cells will be filtered and washed in a plate and frame pressure filter. The filter cake will be mixed with fluxes and smelted in an induction


furnace. The gold and silver doré will be poured in cascading bar moulds, with displacement of slag to the final moulds.

Bullion bars will be stored in a vault prior to dispatch. A bullion balance and sample balance will be provided. Normal levels of security for gold rooms will be employed. There will be two stages of security barrier to access and leave the sensitive areas. Security cameras will be deployed in all areas of the gold room and will be monitored from the security control centre.

18.4.15. Cyanide Detoxification

The discharge of CIL tailings from the carbon safety screens will constitute a barren, alkaline slurry with elevated thiocyanate and cyanide content.

Anticipated levels of free cyanide are 360 ppm and a reduction of the cyanide level in the CIL tailings stream to less than 5 ppm CNWAD prior to deposition is required. This will be achieved by treatment in an INCO S02/Air cyanide destruction plant.

The S02/air cyanide reaction entails the oxidation of the cyanide ion to the non-toxic form “cyanate”, OCN-, and is given below;

SO2 + O2 + H2O + CN- = OCN- + SO42- + 2H+

This reaction is catalyzed by the presence of copper. Base metals that have previously complexed with the cyanide are precipitated as metal hyroxides.

The slurry will be pumped continuously from the CIL thickener underflow at 61% solids to the single cyanide destruction reactor via a down-comer. To achieve the correct % solids for contact recycled water from the CIL TMF will be pumped and mixed with the thickener underflow. Agitation will be used to contact the slurry stream with SO2 and O2 in order to oxidise the weak acid dissociable cyanide (CNWAD) contained in the feed. In this case, liquid gas will be used to provide the SO2 required, and compressed air will be used to supply the O2. The pH will be maintained with the addition of lime. The treatment pH is very important in this process, and lime will be available for pH control. A pH probe installed in the stirred volume of the reactor will be used to maintain the set point using automatic lime addition via a valve.

As there is insufficient dissolved copper in the tailings, copper sulphate solution, (that is used in the flotation circuit), will be used to provide the copper ions used as a catalyst for the reaction.

The detoxified CIL tailings will overflow the Inco reactor and be pumped through an automatic sampler, into the receiving tank from where it will be pumped to the CIL TMF.

18.4.16. Limestone Milling

Approximately 225,000 to 250,000 tpa of limestone will be required for neutralisation in the Albion process tanks. There is a deposit within the concession and analysis of six samples at ALS has shown that it contains over 95% of CaCO3 and has a neutralising capacity of over 960kg/tonne. There is a sufficient reserve of limestone for the life-of-mine.

The limestone will be primary crushed at the point of mining and will be delivered by truck to the plant and will be dumped either by truck or front end loader into the


limestone hopper which will feed a hammer mill. Limestone will be drawn out of the hopper by a variable speed belt feeder, the rate being controlled to a manual set point from the mill feed weightometer. The -12mm product from the hammer mill will directly feed a ball mill which will operate in closed circuit with hydro-cyclones to yield a product with a P80 of 75 microns. Ground slurry from the mill will gravitate to the limestone mill discharge sump, where it will be diluted with raw water and pumped by one of two variable speed mill discharge pumps to the limestone cyclones.

Cyclone underflow will gravitate back to the limestone mill, and the overflow at about 24% solids will be collected and pumped to the two 250m3 limestone storage tanks situated at the Albion Circuit which will provide a live capacity of about 8 hours. One of two available fixed speed pumps will deliver limestone slurry to the Albion circuit via a ring main, with excess limestone being returned to the stock tank. This continuous pumping of ground limestone slurry from the agitated tank is to prevent settling of the solids in the pipeline.

18.4.17. Oxygen Supply

Phases I and II continuous testwork indicated a grossed up oxygen requirement of 550 kg/tonne IsaMill feed requiring 173,250 tpa oxygen. Taking into account availabilities, the design capacity of the oxygen plant will be 520 tpd of oxygen.

The simplest method of obtaining a secure supply of oxygen is with a through-the-fence supply from a reputable supplier. Linde Romania has prepared a quotation to supply oxygen over an eleven year period. Linde would be paid a fixed monthly rental for the plant and a unit fee for each tonne of oxygen consumed. The unit fee is almost completely dependent on the local power cost

The oxygen plant will be a large consumer of energy and will have an installed power of approximately 8.5MW taken from a 6kV supply from the main plant sub station.

The budget quotation from Linde Romania was used in which comprises a Monthly Demand Charge plus a Unit Product Price.

The Unit Price for the supply of gaseous oxygen of €0.025 per Nm3 is based on a power cost of €0.0481/kWh and is almost entirely dependent on the cost of power and the cost model has been adjusted to use a power cost of €0.052/kWh calculated from the A33 tariff published in April 2007. Should there be a decrease in the Romanian unit cost of power there will be a corresponding decrease in the unit cost of oxygen.

18.4.18. Plant Services

Electrical power to the plant will be supplied from the existing 110 kV overhead transmission line from Paulis to the old Certej processing plant. This transmission line will be extended to the new plant site area to feed a high voltage switch yard and transformer station. The 110 kV substation will contain two transformers that will step down the 110 kV to 6.6 kV for distribution throughout the plant site. The transformers will be sized for future load growth with redundant capacity to allow for plant operation with one transformer out of service. 6.6 kV- 400 V dry fan cooled power transformer substations will be provided to distribute power at 400V motor voltage.


High pressure air required for the plant operation will be provided from a central compressor house located adjacent to the plant. Compressed air will be piped to take off points around the plant. Suitable dryers and local storage tanks will be installed for instrumentation air supply.

Gland water will be supplied from a fresh water supply tank and pumped around the plant to the required areas of usage; the primary consumers are the high pressure centrifugal pumps feeding the primary grinding circuit hydro-cyclones and the tailings pumps.

The Process Control System (PCS) will use Distributed Control System (DCS) type controllers and will be used for the majority of process control, drive control, data collection and start-up and shut-down sequences. The plant will also be controlled and monitored by operators via local stations in the main plant, the crusher building control room and gold room floor.

Field instruments, variable speed drives and valves will be wired back to the PCS via panels located in MCC rooms throughout the plant. Closed circuit television will also be used to monitor the operation of process equipment throughout the plant.

18.4.19. Tailings Management Facilities and Water Reticulation

18.4.19.1. General

Residues will arise from the flotation plant and the CIL plant and two separate TMFs to hold flotation and CIL residues respectively will be constructed. Suitable safety factors will be applied to each TMF to allow for possible future expansion.

A TMF facility selection process identified possible sites within the concession area surrounding the Certej open pit. These were formally assessed with respect to political, social, environmental and technical characteristics. Based on the selection criteria and to ensure least disturbance in the area, Macrisului Valley , situated to the north of the open pit site and in the same catchment was been selected as the preferred location.

18.4.19.2. Flotation TMF

The Certej flotation tailings dam will be a homogeneous rock fill structure witha capacity of 27 million m3. Raising of the dam will be in several stages using the centreline method of construction. This method is recommended by the International Commission of Large Dams (ICOLD) bulletins and Romanian standards.

The construction of the dam will be initiated with a coffer dam to control water during construction. When the coffer dam is complete, the main sections of the starter dam will be constructed for the first staged lift. A diversion gallery will be constructed from upstream of the dam, extending downstream following the slope of the valley ending at a location beyond the downstream final toe of the dam. This will control seepage during and after decommissioning.

In the subsequent stages, the dam will be raised in line with the increase of the deposited tailings volume.

When the tailings elevation becomes close to 687 metre above sea level (masl), a saddle dam will be erected in order to separate the two valleys at higher elevations and to


confine the storage to the western valley. The main dam body above the starter dam is pervious / draining.

The dam will be engineered for dry closure. The process of closure will commence prior to the cessation of mining, probably during the last two years.

It is proposed to develop a quarry in Valea Frumoasa to provide construction aggregates for construction of the plant and the TMF. This aggregate may also be sold to generate some early revenue for the project as building stone is currently in high demand for the numerous development projects in the vicinity.

Alternative locations for the TMF are being considered.

18.4.19.3. Cyanidation Plant TMF

The size of the CIL TMF is designed at 8.1 million m3. It will be situated adjacent to the flotation TMF sharing a common wall. It will be at a higher elevation to the flotation TMF of 750 masl. Any seepage will be collected in a large sump and pumped back to the CIL TMF.

The design of the water reticulation circuit for the Albion-CIL circuit is such that the CIL TMF will be zero discharge. This is made possible by the large evaporative losses in the Albion circuit which operates at 97° Celsius and has large volumes of cooling air which dissipate water to the atmosphere. In addition water is consumed in the chemical reaction for sulphide oxidation and large volumes of water are held in the settled solids in the dam.

A second stage CN detoxification plant will be constructed at the process plant to lower the water level in the unlikely event of exceptionally high precipitation in order to maintain effluent quality within the required levels. This CN detoxification plant will use hydrogen peroxide and will also be required following the mine closure. This will ensure that only water with less than 0.1 ppm total cyanide, CNTOTAL will be discharged as per Romanian regulations. This latter level is lower than the current EC regulation.

18.4.19.4. Transport of Flotation and CIL Tailings

There is a major ridge between the plant and the TMF running in a NW-SE direction. The plant is at an elevation of 610m and the ridge at 810m. Therefore a 990m tunnel will be excavated at an approximate elevation of 710m through the hillside through which the tailings and return water pipe lines will run.

The flotation tailings thickeners underflow will be pumped at a density of 60% solids w/w by means of four stages of centrifugal slurry pumps. Water will be returned from the Flotation TMF by two stage centrifugal pumps.

The CIL residue will also be pumped by dedicated pumps and pipelines from the first stage detox plant to the CIL TMF distribution system through the same tunnel.

18.4.20. Water Treatment Plants

It has been decided to locate one water treatment plant adjacent to the flotation plant treating water from the open-pit and ARD from the two dumps. The acidic water is treated by a water treatment plant comprising classical lime neutralisation using one


agitated tank plus a 15m diameter Reactor Clarifier to remove the precipitated sludge with the addition of 17g/m3 of flocculent.

A second plant will be situated below the flotation TMF treating water prior to discharge. The WTP has been designed to treat 150 m3/h of water but the normal flow will be +/-44 m3/h. The design and size of the WTP is the same as the WTP treating the ARD water at the plant.


Figure 18-1: Plots of Arsenic and Tellurium

Te zoning in Certej showing the clear distinction between the Central and Intermediate Zones

As Zoning in Certej showing the clear distinction between the East, Central and Intermediate Zones


Figure 18-2: Sulphide Oxidation vs. Au recovery for Initial Albion Tests

Sulphide Oxidation vs. Au Recovery for the Albion Process on different Certej Concentrates

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

Pyrite Oxidation

Au R

ecov

ery

CentralEastWestIntermediate


Figure 18-3: Comparison of Treatment Options; IRR

Comparison of Various Treatment Options for Certej Concentrate

-10%

0%

10%

20%

30%

40%

50%

$400 $450 $500 $550 $600 $650

Gold Price US$/oz.

Com

para

tive

Post

-tax

IRR

Pressure Oxidation/ CIL

Bacterial Leaching/ CIL

Geocoat/CIL

Albion Process/ CIL

Activox/ CIL

Roast on-site/ CIL

Roast off-site/ CIL


Figure 18-4: Schematic Diagram of Continuous Albion Testwork Equipment


Figure 18-5: Gold and Silver Recovery vs. Sulphide Oxidation for West Concentrate

0

10

20

30

40

50

60

70

80

90

100

0% 10% 20% 30% 40% 50% 60% 70% 80% 90%

Sulphide Oxidation

Gol

d an

d Si

lver

Rec

over

y - %

w/w

Gold Extraction Silver Extraction


Figure 18-6: Schematic Diagram of Certej Process Flowsheet


Figure 18-7: General Schematic Diagram of Certej Processing Route

CERTEJ OVERALL BLOCK DIAGRAM

ALBION SECTION

CONCENTRATOR SECTION

Jaw Crushing

SAG & Ball Mills

Sulphide Flotation& Regrind

Flot Tails TMF

Flot Tails Water Treatment

ROM 440t/hBy Trucks

IsaMill

Albion Process Oxidation

CIL, Elution & Electrowinning

CIL TMF

INCO Detox

Emergency Peroxide Detox

Discharge <0.1ppm CNt

Doré

Open Pit Mine

Discharge to Environment

Mures River

Waste & Low Grade Rock Dumps

ARD Water Treatment

Mine Water

Raw Water

Recycled Water

Recycled Water

Discharge to Environment

Future Gravity Gold Circuit

Concentrate Filtration & Handling


Figure 18-8: Processing Route for the Treatment of the Flotation Concentrate

ALBION SECTION

22t/day8000 g/t Au+Ag

CIL & ELUTION

ALBION PROCESS AND ANCILLARIES

ISAMILL LIMESTONE PLANT

LIMESTONE QUARRY & Jaw Crusher

Jaw Crusher

Hammer Mill

Ball Mill630kW

8 hr Storage Tanks

Tunnel

QUICKLIMESlaking Plant & Ring Main

IsaMill MP100003000kW

Albion Leach5 x Tanks11.5m dia x 15m ht

OXYGEN PLANT

By Supplier520 tpd Over Fence

20m Albion Thickener

Albion Cooling Air Blowers

CIL & DETOX Blowers

Feed Tank & Pump

Slurry Cooling Tower

17m CIL Thickener

ACID WASH & ELUTION11t

CIL6 Tanks30 Hr40%w/w

Neutralisation Tank

GOLDROOMElectrowinning & Furnace

DETOX 1INCO Cyanide Detox

3 Stage Tails Pumps

LIQUID SO2Receipt & dosing

SeepagePump

Albion Feed Tank & Pump

IsaMill Media 2mm Keramax

FLOTATION CONCENTRATE

Loaded Carbon

CarbonRegeneration

Barren Carbon

DORE

CYANIDEMixing & Ringmain

Copper Sulphate

LPG STORAGEBy Supplier

DETOX 2Tank + 10m Reactor Clarifier

H2O2

DISCHARGE TO ENVIRONMENT<0.1ppm CNtotCIL TMF

Copper Sulphate

HYDROCHLORIC ACID

CAUSTIC SODA

Tank Vent Systems

40tph

30tph

7100t Plant Stockpile

P80 9 Microns

40%w/w27 Hr97CpH5.5

60%w/w45CpH10.5

85tph

40tph

40%w/w

60%w/w

3km line3.4km line


Figure 18-9: General Site Layout


Plate 18-1: Typical Photomicrograph showing Arsenical Zoning of a Pyrite Grain


Coarse

0.2-1.6ppm Au; x=0.63ppm Au

This is the principal form of pyrite in the Central & West ore

And the second most important in the other 2 ore zones.

Zoned (euhedral)

0.3-1700ppm Au; x=47ppm Au

Significantly more abundant in the Intermediate ore zone.

Porous

0.2-150ppm Au; x=5.3ppm Au

Principal form of pyrite in East & Intermediate ore zones

Microcrystalline (µ-xline)

1.3-90ppm Au; x=58ppm Au

Significantly more abundant in the West ore than the other ore

zones

Fine grained (disseminated in rock)

3-340ppm Au; x=19ppm Au

Most common in East ore zone

Arsenopyrite

13-580 Au; x=121ppm Au

Significant in the East & Intermediate ore zones as seen on the

chemical plots

Plate 18-2: Photomicrographs showing the 5 Morphological forms of Pyrite described by Amtel


Plate 18-3:Gypsum Crystal in Oxidised Albion Residue


19. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

This section describes the estimation of the gold and silver resource at Certej through the use of geostatistical techniques.

19.1. Bulk Density

A total of 3,615 bulk density measurements were used in the current resource estimate which is considered above industry standard. During 1999, bulk density samples were collected from underground. Since January 2001 diamond core was routinely sampled for bulk density measurements. Following a detailed review of the stated positions of the bulk density samples, RSG Global made the decision to omit all density samples collected prior to January 2001 and European Goldfields concur with this decission. Only diamond core samples have been used to determine the bulk densities of the various lithologies, as a full audit trail is available for these samples and their position in three dimensional space is well established. Bulk density measurements were undertaken at the Cepromin laboratory in Deva, Romania, which is a commercial laboratory previously run by the Romanian Government prior to privatisation. Senior RSG Global staff audited the bulk density measurement process. Samples were collected and data recorded according to detailed mineralised zone location, lithology and style and intensity of alteration. Diamond core samples were prepared by ‘squaring off’ the ends of approximately 15cm billets of half core. Bulk density determination was by standard water immersion method with each sample coated in wax prior to immersion. Standard laboratory samples were used to calibrate the scales between each measurement. All samples were returned to site and the samples placed back into the core trays, without removing the wax coating, as a record. Results are supplied in hardcopy format with the bulk density measurement reported to two decimal places. A summary of the average bulk densities within the different lithologies across the deposit is given below. Table 19-1: Bulk Densities by Lithology

Lithology Bulk Density (t/m3) Cretaceous Sediments Hondol Andesite Neogene Sediments Baiaga Andesite Breccia Dealul Grozii Oxide

2.47 2.44 2.37 2.44 2.35 2.37 2.24

19.2. Database Structure and Content

The Certej database is stored in a GBIS front end linked to a SQL7 database produced by St Arnaud Data Management (SADM) of Perth, Western Australia. The database structure allows all generated data to be stored in the relational database.


The database structure allows all drillhole data to be stored including but not limited to assay data, geology, survey, bulk density, structure, downhole surveys and the equivalent channel data. Discussion of the full databse structure is outside the scope of this report. Drillhole and channel data can be exported using export routines developed specifically for EGL by SADM to a CSV file for import into Vulcan or other resource software. The exported data files have the following structure. Assay Project Concession

Site ID CJR001-028, CJSD001 - CJSD309

Sample ID Sample Number (unique)

Depth From Sample from depth

Depth To Sample to depth

Samp Type DDH, RC, SC, UC

AuPPM Fire Assay Gold 1 result

AgPPM AR/AAS

AsPPM AR/AAS

CuPPM AR/AAS

MoPPM AR/AAS

PbPPM AR/AAS

ZnPPM AR/AAS

Comments Sample Condition ie dry Collar Project Certej

SiteID Drillhole or channel ID

Site Type DH or channel

Work Type DDH, RC or channel

GridLocal ST70

EAST_ST70 East in Stereo 70 grid coords

NORTH_ST70 North in Stereo 70 grid coords

RL Reduced Level

EndDepth Depth of drillhole or channel length

DrillDate Date completed

Lease Concession Number Survey Project Certej


Depth Length of drillhole or channel

Dip Measured from 0degs to -90 for vertical

Azim_Mag Magnetic Azimuth

Azim_ST70 Azimuth in ST70 grid


Geology Project Certej


DepthFrom

DepthTo

Face

Age_Type Neogene

Lith1 Primary rocktype

Weath Oxidation from strong to sulfide or fresh

Sulf Dominant sulfides

Min1 Primary minerals in rock

Colour

ColourInt Lt to dark colour intensity

Grainsize FG to CG

Texture1 Sheared, porphyritic

Texture2 Secondary texture

Lith1Pcnt Lith 1 abundance as a percentage

Samp_Wet Sample condition

Comments Geologist Comments – fault, stope etc

19.3. Lithological Interpretation and 3D Model Creation

Lithological codes used during logging were simplified to make the interpretation easier to finally construct in 3D. East-West orientated 20 to 40m spaced sections displaying geology and assays were plotted at 1:1000 scale. Romanian geologists who have worked on the Certej Project for over 50 man years completed the litho-structural interpretation of the lithological boundaries. Plans spaced at 40m for the Certej deposit were also produced and the lithological boundaries were checked using these and adjusted were necessary to reflect the 3D nature of the contacts. Underground geological mapping collected by the Romanian state from inaccessible underground drives was also incorporated in this process to get a more accurate 3D interpretation of the geology. This interpretation was then used as the basis for the 3D geological model which was completed by local and expatriate geologists. The detailed interpretation was modelled for 3D continuity and the east and west main breccias were also included in the model. The rock types were simplified into the following for the 3D geological model;

• Neogene Sediments • Hondol Andesite (West) • Dealul Grozii Andesite (East) • Baiaga Andesite (Barren Central Andesite) • Cretaceous Sediments • Breccias (both Monomictic and Polymictic were grouped)

A bold outline was drawn around each Lithology on each of the east-west sections and these outlines were then digitised in Mapinfo software. The outlines were then snapped to the relevant drillholes and channels in Micromine and exported to Vulcan software for


final wireframe construction. Figure 19-1 is a west to east cross section through the deposit as modelled.

19.3.1. Mineralisation modelling

The gold and silver grades on the drillholes and channels were then used in conjunction with the litho-structural interpretation to models the mineralisation domains for the Certej Deposit. A series of 0.8g/t gold cut-off outlines were modelled and another set of 0.5g/t gold cut-off outlines were also modelled. The 0.5g/t gold cut-off outlines were considered too complicated to be constructed in to a three dimensional wireframe and were used by RSG and the EGL geologist as a guide for the 3D domains created and used in the current resource estimate. Several domains were modelled in 3D and the outlines were then snapped to the relevant drillholes and channels for 3D accuracy. Wireframes of the outlines were then produced for the estimation work. The geology and mineralisation domains were done in conjunction with EGL and these are considered accurate and representative for the Certej Deposit and reflect the original Romanian interpretation as best as possible.

19.4. Block Model Development

A block model was produced using the Vulcan mining software package. The block model contains sufficient variables to record the results of ordinary kriging (OK) grade estimates and other required parameters. The main block model for Certej was constructed orthogonal to the Stereo 70 grid system, with the model origin, extents, and block details shown in Table 19-2. The resource block model was developed using block dimensions of 25mE x 25mN x 10mRL with sub-blocking to 5mE x 5mN x 2mRL cubic dimensions for the purpose of providing appropriate definition of the topographic surface, geological and mineralisation zone boundaries. The block size was considered suitable with a nominal 40m by 40m drillhole spacing which has been infill drilled in areas to a nominal 20m pattern, with variably spaced underground channel samples throughout the deposit. The block model dimensions are presented in Table 19-2. Table 19-2: Block Model Dimensions

Origin Extent Number Block Size

Parent Sub-block

East North Elevation

344750 500650

0

2200m 1250m 680m

88 50 68

25 25 10

5 5 2

The lithological codes and block model parameters used in the mineral resource estimation are presented in Table 19-3 below. Table 19-3: Block Model Variables


Variables Type Description

lith integer Lithological Domains

domain integer Mineralisation Domains

weath integer Lith=10, weathered=20

void_pc double Underground Development as a percentage

resclass integer 1=measured, 2=indicated, 3=inferred

density double density

den_void double density x (100-void_pc)/100

au_ok double ordinary kriging - gold

au_ok2 double ordinary kriging - uncut

auok_kv double kriging variance

auok_ef integer estimation flag

auok_nc integer number of composites

auok_nh integer number of holes

auok_ad double average distance

ag_ok double oridinary kriging - silver

ag_okuc double oridinary kriging - uncut

agok_ef integer estimation flag

cu_ok double Copper Ordinary Kriged Grade

cu_okuc double Uncut Copper Ordinary Kriged Grade

cuok_ef integer Estimation Flag

pb_ok double Lead OK estimate

pb_okuc double Uncut Lead OK estimate

pbok_ef integer Lead ok estimation flag

zn_ok double Zinc OK Estimate

zn_okuc double Uncut Zinc OK Estimate

znok_ef integer Zinc Estimation flag

t0p6 double SMU block proportion >= 0.6g/t Au








g0p6 double Gold grade of SMU block proportion >= 0.6g/t Au






g1p8 double Gold grade of SMU block proportion >=1.8g/t Au



Variables Type Description

den_0p6 double density x t0p6 x (100-void_pc)/100








au_id2 double Gold ID2 estimate

auid2_ef integer Gold ID2 estimation flag

ag_id2 double Silver ID2 estimate

agid2_ef integer Silver ID2 estimation flag

cu_id2 double Copper ID2 estimate

cuid2_ef integer Copper ID2 estimation flag

pb_id2 double Lead ID2 estimate

pbid2_ef integer Lead ID2 estimation flag

zn_id2 double Zinc id2 estimate

znid2_ef integer Zinc id2 estimation flag

The 3D lithological model was used to code the block model for lithology. The topographical surface was defined from a digital terrain model (DTM), which contains the most current pit pickup for the existing open pit at Certej. The DTM also takes into account the mine waste which was surveyed and the base of the waste is calculated by intersecting with the pre-mining topographical surface. The DTM was based on a helicopter DTM from the helicopter borne geophysical surveys carried out in 2001 but adjusted where data of greater accuracy is obtained such as drillhole collar locations and pit survey.

A wireframe model of the underground mine development at Certej has been represented in the main and sub-blocked models as void space calculated as the proportion of each block intersecting the wireframe model. This proportion is then used to adjust the density of each block in order to simplify volume calculations in each resource run. The volume of underground void space, calculated as 228,854m3 based on the wireframe model, is exactly reproduced in the block model. A new oxidation surface was logged and created for all drilling data at Certej and the model was coded with this surface into oxidised and fresh material. Statistics were also completed on the bulk density data for the oxidised and fresh material and these were also coded to the model. Bulk density assignment was completed via block model scripts using the mean densities for the major rock types calculated during the statistical analysis of the bulk density database. The oxide was combined and a single density was assigned. Bulk density values were assigned the sub-blocked model on the basis of the sub-block rock type coding shown in Table 19-4 below. Table 19-4: Block Model Bulk Density Assignment


Lithology Block Model Code Number of Data Bulk Density (t/m3)

Cretaceous Sediments

Hondol Andesite

Neogene Sediments

Baiaga Andesite

Breccia

Dealul Grozii

Oxide

lith = 10, weather = 20






weath = 10

1229

862

481

181

84

704

19

2.47

2.44

2.37

2.44

2.35

2.37

2.24

19.5. Statistical Analysis

A detailed statistical analysis of the gold and silver captured within the mineralisation envelopes in preparation for resource estimation is summarised below:-

• Analysis of sample lengths and generation of composites • Statistical analysis of the composite gold, silver, copper, lead and zinc data within

the modelled envelopes • Application of upper cuts. Assessment of clustering and determination of

declustered grade statistics • Correlation analysis between gold, silver, copper, lead and zinc grade data • Compositing and Data Coding

The drillhole and channel sample databases coded with geology and estimation domain data were composited as a means of achieving uniform sample support. For the purpose of the estimation, the drillhole database and the channel sample database were combined to form a combined Certej sample database. All statistics, variography, and estimation from this point on have been completed using the combined data set, unless explicitly stated. Several composite runs were undertaken to investigate the appropriateness of various composite lengths and methods. A regular 3m run length (down hole) composite was selected as the most appropriate composite interval to equalise the sample support at Certej. The decision to produce 3m run length composites was based on the following factors:-

• The majority of the data (98%) collected at Certej was collected at 1m. • It reduced the overall variability of the data introduced by the large amount of

channel sampling data thus stabilising the total sampling variance. The use of 3m composites reduced the coefficient of variance of the total data by over 25 percent.

• The presence of underground workings (voids) in the database. The production of longer composites resulted in the production of shorter composite tails in regions of highest gold/silver grade. This is desirable as the tails will tend to skew the data.

• The 3 metre composite is compatible with the SMU block size with approximately 1 composite height per vertical block dimension of 2.5 metre (bearing in mind that most drill holes are at around 50 to 60° dip). Any greater composite length would have smoothed the data too much and reduced the ability to model the grade geostatistical characteristics selectively.


The effects of compositing to 1m, 2m, 3m and 5m unit lengths, on the statistical characteristics of the resultant composite gold grades were investigated and are presented in Table 19-5. As expected, a reduction in the overall variance is noted with an increase in composite length. The statistics provide little evidence of distortion of the mean grades for the different compositing lengths and as such, for the reasons discussed above, the 3m composite datasets were used for all subsequent statistical analyses, variography and grade estimations. Table 19-5: Summary Statistics for Composite Data Composite Data Defined

within the Mineralisation Envelopes Gold (g/t)

In-situ 1.0m 2.0m 3.0m 5.0m

Number

Minimum

Maximum

Mean

Median

Std Dev Variance

Coeff Variance

32,333

0.005

832.00

1.50

0.77

7.02

49.22

4.69

32,777

0.005

832.00

1.49

0.76

7.15

51.10

4.79

16,595

0.005

590.85

1.49

0.82

5.97

35.62

4.01

11,186

0.005

411.37

1.50

0.85

5.27

27.80

3.52

6,883

0.006

248.64

1.50

0.89

4.27

18.26

2.85

19.5.1. Statistical Summary by Estimation Domain

Descriptive and distribution statistics of the data points coded within the modelled domains were generated. Summary statistics of the gold grade data by zone are presented in Table 19-6. Table 19-6: Certej - Gold 3m Composites, Summary Statistics

Main zone All East Intermediate Central West Central (below

pit) RSG Domain All

Data Far East

East East HW

East FW1

East FW2

West Far

West Central

Count 10,92

3 1928 4,415 560 234 47 3382 233 124

Minimum 0.005 0.01 0.013 0.005 0.07 0.033 0.007 0.005 0.267

Maximum 411.3

7

198.4

4

138.2

4 15.01 7.15 3.51

411.3

7 7.39 7.72

Mean 1.49 1.95 1.58 1.40 0.84 1.42 1.22 0.80 1.52

Median 0.85 1.29 0.88 0.97 0.68 1.30 0.68 0.57 1.13

Standard Deviation 5.25 5.06 3.14 1.55 0.66 0.79 7.79 0.92 1.15

Variance 27.52 25.56 9.84 2.40 0.43 0.62 60.67 0.85 1.32

Coefficient of Variation 3.52 2.59 1.99 1.11 0.79 0.55 6.37 1.16 0.76

It should be noted that there is some overlap between the RSG domains and the main zones defined by mineralogy. This is because the RSG domains are based purely of continuity of gold grade which can be seen between the Central and Intermediate zones and the Intermediate and East zones in some areas. This is consistent with the Intermediate zone being a transition type zone with mixed mineralogical characteristics of both the East and Central zones.


Similar overall mean gold grades are reported for the east and west domain datasets, but the highest grade composite (411.37g/t Au) accounts for over 10% of the mean grade for the west domain. The influence of this composite and other high grades in the east and west domains is further demonstrated by the high to very high coefficient of variance (CV) values. While moderate CVs of 0.76 and 0.55 are reported from the central domains and east FW2 domain respectively, the mean grades of 1.52g/t Au and 1.42g/t Au are only based on a total of 1,124 and 47 composites respectively. The mean silver grades ranged from 5g/t Ag for the far east and east hanging wall (HW) domains to 13g/t Ag, for the main east domain. Most of the remaining domains have an average silver grade near 10g/t Ag (except the far west domain that has an average grade of 3g/t Ag). High CV values of above 1.4 are evident for all domains.

19.5.2. Investigation of High Grade Outliers

The behaviour of the tail of the metal distributions for the estimation domains was assessed in order to identify any sampling errors or separate populations, and to ascertain the risk of local and global overvaluation. The effects of the highest grade composites on the mean grade and standard deviation of the gold, silver, copper, lead and zinc datasets for each of the estimation domains were investigated by compiling and reviewing plots displaying point series of the mean grade and standard deviation of the datasets as successive high-grade data are removed in descending order. These plots were generated for each of the estimation domains. The resultant plots were reviewed together with probability plots of the sample populations and an upper cut for each dataset was chosen coinciding with a pronounced inflection or increase in the variance of the data. A list of the determined upper cuts applied and their impact on the mean grades during the OK gold grade estimations is provided in Table 19-7. Table 19-7: Outlier Analysis - Gold (g/t)

Domain Number

Data Maximum Mean

Std Dev

CV Upper

Cut Cut

Mean

Cut Std Dev

Cut CV

Number Data Cut

% Change

in Mean

East 4,415 138.24 1.58 3.14 1.99 27.00 1.54 2.15 1.40 4 98%

Far East 1,928 198.44 1.95 5.06 2.59 36.00 1.87 2.48 1.33 1 96%

East HW 560 15.01 1.40 1.55 1.11 15.01 1.40 1.55 1.11 0 100%

East FW1 234 7.15 0.84 0.66 0.79 3.50 0.82 0.54 0.66 1 98%

East FW2 47 3.51 1.42 0.79 0.55 3.51 1.42 0.79 0.55 0 100%

West 3,382 411.37 1.22 7.79 6.37 25.00 1.05 1.73 1.65 7 86%

Far West 233 7.39 0.80 0.92 1.16 4.00 0.78 0.83 1.07 2 98%

Central 124 7.72 1.52 1.15 0.76 5.00 1.49 1.01 0.68 2 98%

19.5.3. Analysis of Data Clustering

Possible preferential sampling of the higher and/or lower grade areas of the estimation domains was investigated by de-clustering the cut composite grade data for each metal type. The cell size used for each domain was configured for each population. De-clustering weights were calculated as 1/n, where “n” is the number of composites in


each cell. A summary of the raw (cut) versus de-clustered (cut) mean grades is presented in Table 19-8. The de-clustered mean gold and silver grades are significantly less than the raw mean grades, in particular for east domain. This indicates that clustered sampling in high grade areas has biased the mean significantly and therefore a de-clustering approach should be taken when interpolating grade. Table 19-8: Comparison of Raw and Declustered Mean Grades by Estimation

Domain

Block Decl. Cell

Dimensions

Au (g/t) Ag (g/t)

Raw Declus % Dif Raw Declus % Dif

East

Far East

East HW

East FW1

East FW2

West Far

West

Central

55 x 55 x 5.5

50 x 50 x 5

60 x 60 x 6

60 x 60 x 6

60 x 60 x 6

50 x 50 x 5

50 x 50 x 5

50 x 50 x 5

1.54

1.87

1.40

0.82

1.42

1.05

0.78

1.49

1.41

1.80

1.30

0.82

1.49

1.03

0.77

1.45

91%

96%

93%

99%

105%

98%

98%

97%

13

5

4

10

7

10

2

10

12

5

4

10

7

9

2

9

89%

88%

101%

102%

98%

86%

100%

97%

19.5.4. Correlation Analysis

Correlation analysis was completed between the composite gold, silver, copper, lead and zinc datasets for the combined estimation domains. Coefficient values can range between 0.0 and 1.0, with 0.0 indicating no correlation, and 1.0 indicating perfect correlation. Little to no correlation is evident between gold and the other metals. Copper appears to be moderately correlated with silver and the other base metals, while the best correlation is evident between lead and zinc.

19.6. Variography

Variography is used to describe the spatial variability or correlation of an attribute (gold, silver, sulphur etc). The spatial variability is traditionally measured by means of a variogram, which is generated by determining the averaged squared difference of data points at a nominated distance (h), or lag. The averaged squared difference (variogram or γ(h)) for each lag distance is plotted on a bivariate plot where the X-axis is the lag distance and the Y-axis representing the average squared differences (γ(h)) for the nominated lag distance. Fitted to the determined experimental variography is a series of mathematical models which, when used in the kriging algorithm, will recreate the spatial continuity observed in the variography. For the Certej study, all variography is based on the generated 3m run length composites and generated for each of the metals. Detailed grade variography was completed for the combined west domains, the far-east domain, the shallow combined east domains (east domains 2 to 5), the steep combined east domains (east domains 2 to 5) and the central domains.


The gold variograms are characterised by a moderate nugget in the west domain and moderately low nuggets in the east domains of between 26% and 46% of the variance. For most domains, the majority of the non-nugget variance is at or less than the drillholes spacing. Silver variograms are characterised by moderate nuggets of between 23% and 48% of the variance. The majority of the non-nugget variance is at distances greater than the average drill spacing. Table 19-9 lists the variogram models used for estimation and for gold in the uniform conditioning step (for selective mining unit emulation). In all cases a multi-structured nugget and spherical model has been applied. The resulting models were checked against the known behaviour of the mineralisation for each of the zones and metals, and were found to support these models. Table 19-9: Variogram Model Parameters

Gold

Domain Co Rotation

C1 Range 1 (m)

C2 Range 2 (m)

X Y Z X Y Z X Y Z

West 6-7 1.5 0 0 70 1.18 30 20 11 0.6 155 110 56

East 1, 3-5 Shallow 0.77 0 0 -35 1.04 25 25 9 0.56 78 78 25 East 1, 3-5 Steep 1.17 0 0 -55 2.37 19 19 19 0.97 73 73 73

Far East 2 1.8 0 0 -80 1.9 28 28 16 1.75 110 110 54

Central 8 0.46 0 0 0 0.28 12 12 12 0.26 35 35 35

Silver

Domain Co Rotation

C1 Range 1 (m)

C2 Range 2 (m)

X Y Z X Y Z X Y Z

West 6-7 0.28 0 0 70 0.39 50 35 17 0.33 90 69 37

East 1, 3-5 Shallow 0.32 0

0 -35 0.4 29 24 24 0.28 100

70 57

East 1, 3-5 Steep 0.23 0 0 -55 0.42 57 29 11 0.35 110

85 59

Far East 2 0.48 0 0 -80 0.42 23 23 13 0.1 83 83 48

Central 8 0.46 0 0 0 0.28 5 5 5 0.26 17 17 17

19.7. Block Model Estimation and Validation

Ordinary Kriging (OK) was used to estimate gold and silver grades in the modelled mineralisation domains for the Certej project. Uniform Conditioning (UC) was subsequently used to produce a selective mining unit (SMU) resource estimate for gold. The decision to use UC was based on a review of a number of items, including:-

• The mineralisation constraints - broad zones of mineralisation delineated that include a significant volume of sub-grade material.

• The drillhole spacing - the drillhole spacing generally ranged from nominal 20m x 20m drill pattern, to a maximum of 80m section spacing.

• The variography - significant short scale variability (i.e. moderate to high relative nugget effects and dominant short scale structures) was modelled in the variography.

• The mining approach- a selective open cut mining was being targeted with the SMU considerably smaller than the current drill spacing.

Following assessment of the above items it was considered that a recoverable resource model was appropriate wherein a high quality panel estimate could be generated from


which a selective mining model could be estimated. UC, a widely used geostatistical estimation technique, is particularly applicable to the determination of recoverable tonnes, and grades for selective mining scenarios given the relatively sparse data, significant short-scale variability, and selective mining approach envisaged. In earlier resource estimations for Certej carried out in 2002 and 2005 the SMU approach using uniform conditioning was applied. In each of the subsequent resource estimations, the spacing of the drilling has been reduced, particularly in high grade areas. This has acted as a means of verifying the validity of the SMU approach. In each case where the spacing was reduced the grade from the parent OK model remained similar whilst the grade from the SMU model improved. In addition by looking at the proportion of higher grade intercepts within a block, the increase in grade reflected the increase in the number of higher grade intercepts. It is therefore concluded that the SMU approach using uniform conditioning applied to ordinary kriged parent blocks accurately reflects the natural variability of the ore and the ability to mine selectively in practical mining unit dimensions and is therefore valid and representative.

19.7.1. Ordinary Kriging

Ordinary kriged gold and silver grades were estimated in the 25mE x 25mN x 10mRL matrix of blocks comprising the east and west mineralised domains using a staged sample search outlined in Table 19-10 below. The OK estimation was based on the cut 3m composite data discussed in Section 5, and the variogram models discussed above. Table 19-10: Sample Search Parameters - Gold and Silver Ordinary Kriging

Domain Pass

Search Orientation Search Radii Number of Samples

Bearing (Z)

Plunge (Y)

Dip (X)

Major Axis (m)

Semi-Major Axis

(m)

Minor Axis (m)

Min Max Max / Hole

West

1

2

3

0

0

0

0

0

0

70

70

70

65

90

120

65

90

120

20

30

40

24

12

6

32

32

32

5

5

East 2-5

Shallow

1

2

3

0

0

0

0

0

0

-35 -

35 -

35

70

90

120

70

90

120

30

40

50

24

12

6

32

32

32

5

5

East 2-5

Steep

1

2

3

0

0

0

0

0

0

-55 -

55 -

55

55

75

100

55

75

100

20

30

40

24

12

6

32

32

32

5

5

Far East

1

2

3

0

0

0

0

0

0

-80 -

80 -

80

50

70

120

50

70

120

20

22.5

40

24

12

6

32

32

32

5

5

The sample search orientation parameters reflect the mineralisation domain interpretations provided by Deva Gold, while the staged estimation approach is designed to force relatively more input data to be used to estimate regions well informed by drilling, and allow less input data to be used in regions of lower density drilling. In addition, a maximum of 5 composites from any one drillhole were used to restrict the effects of data clustering for the first and second passes. The applied sample search parameters are the result of detailed search neighbourhood testing.


Soft data boundaries were used, enabling data from all domains to be used during the estimation of grades in the individual domains. Block discretisation was applied based on 5 points in the easting and northing dimensions respectively and 4 points in the RL dimension, for total of 100 discretisation points per block. Estimation statistics, including kriging variance, number of composites used, number of holes / channels from which data were collected, nearest and average distance of data from the block centroid and slope of regression were recorded for each block. Any sub-blocks within the 3-D limit of each whole block were assigned the whole block grade estimate.

19.7.2. Change of Support Estimates for Gold

In many resource estimations in mining, the issue of ‘change of support’ is of great significance. The variability encountered in a deposit when taking small samples is very different from the variability encountered when mining the same deposit using larger (selective) mining units. The change of support calculation attempts to model these differences. In the case of the Certej resource estimates, the recoverable resources were calculated by the UC method, using Isatis software. Uniform conditioning allows the estimation of the proportion of ‘ore’ above a particular cut-off, within panels large enough for robust estimation when it is unrealistic to try to achieve this by directly estimating such a small block size, taking into consideration the average drillhole and composite spacing. Estimation of the parent blocks was carried out by OK. When calculating the proportion of ‘ore’ above a cut-off, it is necessary to determine how selective the mining will be. The Selective Mining Unit (SMU) reflects the mining selectivity. The selective mining unit (SMU) block size would represent a single blast in lateral extent with two units per bench blast. When UC was carried out, the proportion of ‘ore’ and metal tonnes above a cut-off were computed by determining how many SMUs within a panel were above that cut-off. In this study, it was established that a 25m x 25m x 10m panel could be reliably estimated from the given drilling grid, and that a SMU of 6.25m x 12.5m x 2.5m would be used to represent mining selectivity. This means that within each panel there are approximately 48 SMUs, so if 24 SMUs had a mean grade above the cut-off grade, then that would equate to 50% of the panel being above the cut-off. The modelled grade variability (the variogram) was used to calculate the expected variability of SMU size parcels of ‘ore’ (which will be less than that of samples or composites). The modelled distribution of composites (weighted) was then adjusted to reflect the expected variability of mining units. The adjustment factor is called the change of support coefficient. This process assumes that the distribution of SMUs de-skews as the SMU size increases (which is expected) and adjusts the distribution appropriately. Other methods (not employed here) do not make this adjustment for de-skewing. Change of support was applied by UC for cut-offs of 0.2, 0.4, 0.6, 0.8, 1.0, 1.2, 1.4 1.6, 1.8 and 2.0g/t Au. Parameter details are in Table 19-11 and Grade tonnage curves for the UC estimate and the OK estimate are given in Figure 19-2.


Table 19-11: Uniform Conditioning Parameters

West Shallow

East Steep East Far East

Block Correction

SMU size 6.25mE x 12.5mN x 2.5mRL

Discretisation 10 x 10 x 1 10 x 10 x 1 10 x 10 x 1 10 x 10 x 1

Punctual variance (anamorphosis) 3.4 2.35 4.49 5.46

Variogram sill 3.01 2.37 4.51 5.45

Gamma(v,v) 1.8 1.21 2.17 2.56

Real block variance 1.6 1.14 2.32 2.91

Real block correction (r) 0.83 0.80 0.80 0.84

Kriged block correction (s) 0.83 0.80 0.80 0.84

Kriged block-Real block correction (rho) 1 1 1 1

Information Effect

Hole spacing Hole direction Sample

length

5 x 8 -

55→090 1m

5 x 8 -

55→270 1m

5 x 8 -

55→270 1m

5 x 8 -55→270

1m

Variance of Z* (Estimated Z) 1.1247 1.095 2.1344 2.7456

Covariance between Z and Z* 1.1318 1.0977 2.187 2.7747

Kriged block correction (s) 0.81 0.78 0.77 0.82

Kriged block-Real block correction (rho) 0.99 0.98 0.98 0.99

Kriged Panel Correction

Panel size 25mN x 25mE x 10mRL

Discretisation Punctual variance

(anamorphosis) Kriged panel correction (S)

5 x 5 x 4

3.40 0.40

5 x 5 x 4

2.35 0.46

5 x 5 x 4

4.49 0.48

5 x 5 x 4 5.46

0.52

There are instances where there will be less than one composite per SMU block hence the need for the change of support and why the model does not assign definitive grades to each SMU but rather the proportion of SMU units above each cut-off limit for each parent block. Dilution arising from this is taken in to account by applying an Information Effect and also to try and account for ‘ore’ selection errors during grade control. It was assumed that for the west domain, grade control holes will be drilled at -55º to the East, on a 5mE x 8mN pattern, and will be sampled in 1m intervals. For the remaining domains, it was assumed that the grade control holes will be drilled at -55º to the West, on a 5mE x 8mN pattern, and will be sampled in 1m intervals.

19.8. Mineral Resource/Reserve Classification and Reporting

The resource estimates for the Certej deposit have been categorised in accordance with the criteria laid out in the Canadian National Instrument 43-101 (CNI43). A combination of Measured, Indicated and Inferred Resources have been defined using definitive criteria determined during the validation of the grade estimates, with detailed consideration of the CNI43 (CIM) categorisation guidelines. The resource estimate(s) have been classified as a combination of Measured, Indicated and Inferred Mineral Resources based on the confidence level of the key criteria that was considered during resource classification as presented in Table 19-12 below.


Table 19-12: Resource Classification Criteria

Data Density and

Distribution

The sample data array varies considerable throughout the Certej

resource. Underground channels are spaced between 20 and 50

meters. Surface drilling represents a notional 20m x 20m pattern in

the well drill eastern region with the remainder drilled up to an 80m x

80m pattern.

High

Audits or Reviews Independent parties have not audited the current resource estimate. NA

Database Integrity RSG Global has completed substantial validation of the database. High

Geological

Interpretation

The interpreted lithological boundaries are considered robust and of

high confidence, while the defined mineralised domains and oxidation

profile are predominantly of moderate confidence.

Moderate

Estimation and

Modelling

Techniques

Estimates based on detailed statistical and geostatistical analysis.

Ordinary Kriging used to estimate gold, silver, copper, lead and zinc

in the West and East domains with Uniform Conditioning used to

produce selective mining unit gold estimates. Inverse distance

squared used to estimate grades in Central domains.

Moderate

to High

Cut-off Grades A notional 0.5g/t gold lower cut-off grade was used to develop the

estimation domain boundaries used to constrain the resource.

NA

Mining Factors or

Assumptions

Whole block estimates for all mineralised regions completed for 25mE

by 25mN by 10mRL size blocks, while notional 6.25mE by 12.5mN by

2.5mRL block dimensions are emulated in the selective mining unit

(SMU) estimate for the Upper Zone. The SMU estimate is based on

the availability of close spaced drilling.

NA

Metallurgical

Factors or

Assumptions

Not applied NA

Tonnage Factors

(Insitu Bulk

Densities)

Large amount of data collected by industry standard methodologies. High

Measured Resources are restricted to the andesite hosted eastern domain between development and up to 10m below the lowest sampled level. This region is defined by on average, 20 x 20m spaced drilling and represents the highest confidence area of the deposit with the least gold grade variability.

Indicated Resource is restricted to regions where a strong geological understanding exists, and the estimate is based on 40m x 40m or better drilling. Inferred Resource represents the lower geological confidence blocks that do meet the indicated or measured criteria. Categorised resources at Certej for the combined east and west estimation domains and the inferred central and far west domains are reported in Table 19-13. The tonnage and grades reported reflect the total material within the domain envelopes, excluding the volume of the modelled underground voids, with no lower cut-off grade criteria applied.


Table 19-13: Total Resources All Domains UC Estimate at 0.8 g/t cut Off

Resource Category Million Tonnes Au g/t Ag g/t Measured 3.9 2.3 5 Indicated 37.6 1.9 11 Measured and Indicated Total 41.5 2.0 11 Inferred (East-West) 3.4 1.6 4 Inferred (Far West/Central Domains) 3.8 1.4 8 Inferred Total 7.1 1.5 6

19.9. Validation and Comparison with Previous Model

Extensive visual and statistical validation of the grade estimates was completed. This process included:-

• Visual comparison of the block estimates and composite data in cross section, long section, plan and oblique views.

• Comparison of global mean grades based on the block model estimates versus the raw and de-clustered input composites data.

• Comparison of block model and composites mean grade data for sequential block increments in the northing, easting, and RL directions (gold only).

• Comparison of grade, tonnage and metal distributions for panels (25mE x 25mN x 10mRL) based on ordinary kriging and global change of support estimates, and selective mining units (6.25mE x 12.5mN x 2.5mRL) based on uniform conditioning and global change of support estimates (gold only).

The visual validation has determined that the OK block grade estimates replicate the source input data well in regions of higher density sampling, such as the eastern domains (Figure 19-3), but that moderate smoothing has occurred in the areas where the data density is lower, mainly in the west of the deposit. This will result in a slight reduction in grade and increase in tonnes in the estimate in these areas but not in the overall metal content. This may lead to an improved grade of ore from these areas which provide feed in the latter years of the current life of mine.

Table 19-14: Comparison of Raw and De-clustered Mean Grades by Estimation Domain (based on cut gold and silver data)

Block Au (g/t) Ag (g/t)

Raw Declus OK Est* Raw Declus OK Est*

East

Far East

East HW

East FW1

East FW2

West

Far West

Central

1.54

1.87

1.40

0.82

1.42

1.05

0.78

1.49

1.41

1.80

1.30

0.82

1.49

1.03

0.77

1.45

1.39

1.90

1.54

0.80

1.41

1.02

0.80

1.38

13

5

4

10

7

10

2

10

12

5

4

10

7

9

2

9

12

5

4

8

8

9

2

9

Table 19-14 indicates that there is close agreement between global mean grades based on the block model estimates and the input composites data. The OK grade averages


are shown here which are the same as those for the UC since the value for each parent block is the mean of all the sub-blocks in it (remembering that the sub-blocks are not defined with individual grades but as a proportion of each parent block above a cut-off). The whole block estimate was compared against the input data by plotting the average gold grade grouped by easting. An example plot is shown in Figure 19-4 for the East zone. These charts show that the grade trends based on the averaged cut composites (peaks and troughs) are adequately reproduced in the whole block estimate. Table 19-15: Comparison between 2005 and 2007 SMU Estimates Uniform Conditioning - Gold g/t

2005 - SMU - 6.25mN x 12.5mE x 2.5mRL 2007 - SMU - 6.25mN x 12.5mE x 2.5mRL

Cut-

off Au

(g/t)

Resource

Category

Tonnes

(Mt)

Grade

(g/t)

Metal

(Moz)

Cut-off

Au

(g/t)

Resource

Category

Tonnes

(Mt)

Grade

(g/t)

Metal

(Moz)

0.6

0.8

1

1.2

1.4

1.6

1.8

2

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

48.3

39.01

31.35

25.29

20.53

16.78

13.78

11.41

1.7

1.9

2.1

2.4

2.7

2.9

3.2

3.4

2.59

2.39

2.17

1.95

1.75

1.57

1.41

1.26

0.6

0.8

1

1.2

1.4

1.6

1.8

2

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

Meas+Ind

51.81

41.47

33.20

26.79

21.83

17.94

14.85

12.38

1.7

2.0

2.2

2.5

2.8

3.0

3.3

3.6

2.84

2.61

2.37

2.15

1.94

1.75

1.58

1.43

The 2005 to 2007 SMU estimate comparison shown in Table 19-15 indicates that the updated resource has a higher grade and more tonnes (See also Figure 19-5). The higher grade is due to the newer drilling, particularly in the west, being on average a higher grade. The increase in tonnage is due to the conversion of Inferred to Indicated and the new interpretation expanding the estimate area.

19.10. Devagold Polygonal Resource Calculation Comparison

Independent polygonal resource estimations were completed by Devagold and European Goldfields for the Certej Deposit utilising the same data supplied to RSG. Lithology, structure and alteration models were used to identify zones of grade continuity on section using 0.5 g/t Au cut off throughout the deposit and at 0.8 g/t. The lower grade modelling helped identify the trends within which the higher grade envelopes sit. All the domains are lithostructurally controlled and their main characteristics are listed in the table below. A summary of the domains is given Table 19-16 below. Table 19-16: Deva Gold Resource Estimate Domains

Area Lithology Style of Mineralisation

Dominant Secondary

East

Sacaramb Andesite Impregnation Vein

Polymictic Breccia\Neogene & Cretaceous

Sediments

Dominant

Impregnation Minor Veining

Intermediate Cretaceous Sediments\Brecciated


Dominant

Impregnation Minor Veining


The modelled zones are very well constrained using a maximum 20m extrapolation and based on 20 to 40 metre spaced sections. A resource estimate for gold was completed by using digitised outlines in Micromine software. The outlines were then validated to check if all required assays are captured and modified where necessary. A resource estimate for the model was completed which calculates the volume, tonnage using an average density of 2.4 and the weighted gold grade average of each outline on each section and this is summarised in Table 19-17 below. Table 19-17: Certej Polygonal Resource Estimation Comparison with SMU model Estimate Cut-off Resource Mt Au g/t Au oz Polygonal 0.8g/t 40.71 1.95 2.55 SMU 0.8g/t 41.47 2.0 2.61

The Deva Gold resource estimate is unclassified but the maximum interpolation of 20 metres is consistent with measured and indicated resources. The Deva Gold estimate compares well with the RSG uniform conditioned SMU model for Measured and Indicated resources with the tonnes estimation within 2% and the grade estimation within 2.5%. The selectivity of the hand produced outlines replicates the selectivity of the SMU model and the slight generalising of these outlines during digitisation replicates the ‘information effect’ applied by RSG to account for dilution and errors in mining selection.

19.11. Dump Resource Estimation

19.11.1. Description of Dump Topographic Survey and Surface Generation

The generation of solids for the polygonal estimate for the north and south dumps required a reliable upper and lower wireframe. These two wireframes were then pressed together to form an intersecting solid that became the north and south solids used in this polygonal estimate. All the wireframes built in this process have been generated in Micromine Version 11.0.4, a commercially available exploration and mining software package, with the head office base in Perth, Australia.

Cretaceous Sediments\Brecciated


Dominant

Impregnation Vein

Polymictic Breccia\Cretaceous Sediments &

Baiaga Andesite

Dominant

Impregnation

Central

Neogene Sediments\Brecciated Neogene

Sediments

Dominant

Impregnation

Polymictic Breccia\Cretaceous Sediments &

Baiaga Andesite

Dominant

Impregnation Vein

West

Hondol Andesite Vein Impregnation

Polymictic Breccia\Hondol Andesite &


Dominant

Impregnation Vein

Hondol Andesite Vein Impregnation


19.11.1.1. Existing Topography

The upper dumps topographic surface was generated by a commercial survey company, SC Donamar SRL (Ltd) who are authorised to conduct surveying in Romania. At the time of the survey, both the north and south open pit files were checked by the registered Romanian surveyor on site and then supplied to Deva Gold in both AutoCAD, dwg and dxf format. The original topography was also supplied in both AutoCAD, dwg and dxf format, by the contracting surveyor on site and this is based on the Romanian Cadastral 1980, 1:5,000 topographical sheets.

The two files were then imported into Micromine (MM) and digital terrain models (DTM) were generated from points.

The drill hole collars and channel RL’s have been cross-checked against the MM generated wireframe to confirm the accuracy of the DTM. On nearly all accounts the upper DTM proved to be reliable.

19.11.1.2. North and South Dump Wireframe and Solid

The process of generating the upper north and south dump wireframes is similar. An outline of the north and south dump was generated from field, channel and drill hole observations. This outline was then pressed onto the MM generated topographic DTM. The DTM was then cut by the outline and the wireframes inside the outline becomes the upper dump wireframe (Figure 19-6).

The process of generating the lower north and south dump wireframes are the same. On a section-by-section basis within a sectional window, typically +/- 20m. A polyline was generated by snapping to drill holes at the intersection of dump material and in-situ soil/bedrock and the outline that defines the outer dump limit. It was noted in this process that the 1980, 1:5,000 topographical surface proved to be unreliable and was only used as a guide in defining the dump limits. Each polyline was then connected by tie lines and from this structure the lower wireframe was built using the proportional length function in MM (Figure 5.24).

The two wireframes – upper and lower were then pressed together and a solid was generated. The wireframe was then validated for self intersections, closure and invalid topology. After passing this phase, the wireframe could be used in MM to calculate an in-situ volume.

19.11.2. Dump Resource Estimate Description

The approach taken to estimate Coranda’s north and south open pit in-situ dump tonnage and grade has been to deliver a polygonal estimate that is simple and provides a global approach for the two dumps as total extraction would not involve sophisticated grade control.

To achieve this s it was agreed that a polygonal estimate would be appropriate. The polygonal estimate is based on the pressing of an upper and lower wireframes together in Micromine. The upper wireframes was systemically built from reliable survey information in AutoCAD and the lower wireframe from down hole RC interval data. The upper and lower wireframes share the same outlines at surface removing errors when


they are pressed together. To calculate a volume, tonnes and grade, Micromine provides the polygonal wireframe estimate function.

19.11.2.1. Polygonal Wireframe Estimate

The polygonal wireframe estimate function in Micromine requires the input of several parameters. The main parameter being a closed wireframe, a data source that contains intervals as 3D coordinates (trace coordinates) and a drill hole and channel database. Other parameters also include SG and void factor, unique to this polygonal estimate.

• The process for the generation of the wireframes has been covered in Section 5

• 3D coordinates – trace coordinates are required in the interval file and this process is easily achieved under drillhole function in Mircomine

• Prior to running this process 1m RC intervals must be converted to 2m composites. This is easily achieved in the downhole coordinates function in Micromine. Compositing of channel intervals is not required as they are already represented as uniform 2m lengths.

Parameters used in Micromine, under polygonal wireframe estimate function include:

• Select wireframe - set of wireframes have been used (north and south dump)

• Data source - interval file containing 2m composites from the original RC and channel assay and survey files. Mid-points for each interval must be coded into the file (generated from the Micromine drillhole function).

• Input Data

o Interval file – 2m composite assay file containing RC and channel intervals, with calculated mid-points for each interval (generated from the Micromine drillhole function). If a mid-point falls within the wireframe then it is included in the polygonal estimate calculation.

o Filter – none applied

o Hole, From & To field – assigned from interval file

o Drill hole setup – assigned drillhole database (made up of the collar, survey, assay and lithology files applicable to polygonal estimate)

• Method - simple average has been us. This option calculates the average grade of points that fall within the wireframe, using a mid-point, no attempt is made by Micromine to assign a portion of the interval that falls inside the wireframe. As the channel samples are a consistent 2m intervals, it was not effective to crate small intervals.

• Metal Calculation - Tonnes & grade

• SG – 1.44 has been assigned. This is explained in more detail in Section 3.4, and considers a void factor of 0.6, for space and poorly packed material in the dumps. The 0.6 is multiplied by SG to give 1.44, our in-situ dumps SG.

• Grade Fields

o Fields that are considered in the polygonal estimate, eg. Au and Ag.


o No default grade was applied in the polygonal estimate. Therefore if an interval has a missing assay due to poor sample recovery, no value was assigned.

o No cuts were applied to either gold or silver, with maximums in interval file (2m composite file) for North dump at 9.97 g/t (RC interval) and South at 11.52 g/t (RC interval). A summary of interval (2m composite) file used in discussed in Section 3.3.4.

• Numerical exceptions

o Ignore characters in Ag & Ag fields

o Ignore blanks in Au and Ag fields

o < x = 0.5x – lower than detection (LDL) values are halved, eg -0.01 g/t

19.11.3. Dump Resource Results

Based on vertical RC drilling from 2004-2005 and 2008 at approximately 40m x 40m and an extensive surface channel programme over all berms and some benches of the Coranda upper north and south dump surfaces, a global Indicated resource with an in-situ tonnage of 7,022,434 at 0.53 Au g/t and 8.9 g/t Ag, with a contained metal of 119,600 oz Au and 2,002,100 oz Ag, has been estimated (Table 19-18). This resource can be categorised as Indicated under CIM guidelines.

Table 19-18 Polygonal Indicated Estimate for North and South Coranda Dumps.

Category Volume SG

Insitu

Tonnage

Au

g/t

Au

Ounces

Ag

g/t

Ag

Ounces

Certej North

Dump Indicated 2,141,307 1.44 3,083,482 0.53 52,500 8.29 821,800

Certej South

Dump Indicated 2,735,383 1.44 3,938,951 0.53 67,100 9.32 1,180,300

Total Indicated 4,876,690 1.44 7,022,434 0.53 119,600 8.87 2,002,100

19.12. Reserves

The reserves are based on the geotechnical and mining considerations given in section 20 below. All stated probable reserves are completely included within the quoted Mineral Resources. Further exploration drilling will take place prior to production which is expected to convert some of the probable reserves to proven reserves. Table 19-19 presents a summary of the Certej Mineral Reserves.


Table 19-19: Certej Mineral Reserves ‘000t Au Au Ag Ag

g/t Moz g/t Moz


Probable 32,811 2.0 2.12 11.4 12.0


Probable

7,829 0.7

0.18 14.0 3.5 Certej Existing Dumps

Probable

6,320 0.5

0.11 8.9 1.8

Total

46,960

2.41 17.3 This reserve was estimated and reported in accordance with the Instrument and the classifications approved by the CIM Council in November 2004. The reported reserves were compiled by RSG, which has sufficient experience, relevant to the style of mineralisation and type of deposit under consideration.

19.13. Conclusions on Geology and Resources

The geology of the Certej locality and deposit is well established and a detailed regional and local model for the mineralisation has been put together by European Goldfields and Deva Gold. Detailed drilling, data analysis and modelling have resulted in the development of a robust geological and grade model. The resource estimate is well supported with over 60,000 metres of drilling and 22,000 metres of channel sampling. A significant proportion of the channel and drilling data collection and QAQC was completed under the independent supervision of consultants RSG. The QAQC programme is above industry standards and detailed analysis has shown no bias in either assay precision or accuracy. All assaying has been completed at reputable internationally recognised independent laboratories and samples submitted with internationally recognised standards from Rocklabs of New Zealand. Sample quality from RC and diamond methods is exceptional and underground channels are still preserved today for any future reference. All sample pulps are stored in a dry and locked facility on site in Certej. All diamond core and chip trays of RC samples are also stored in locked facilities at Certej.


Geological wire framing and three dimensional modelling of domains has been carried out based on careful mapping of four mineralogical zones, East, Intermediate, Central and West. These have distinct mineralogical signatures and correspondingly distinctive gold deportment. Their boundaries correspond to litho structural features which have also been modelled in three dimensions. Sampling of all of these zones has been representative. Block modelling and interpolation using ordinary kriging and uniform conditioning in to a selective mining unit block model have resulted in a Measured and Indicated resource of 41.5Mt at 2.0g/t Au and 11g/t Ag grade. In addition, an Indicated resource of 7,022,434 tonnes at 0.53 Au g/t and 8.9 g/t Ag, with a contained metal of 119,600 oz Au and 2,002,100 oz Ag has been estimated in existing dumps using a polygonal wireframe approach. The resource and reserve has been subject to several audits by RSG during the course of the projects development. The resource and reserve are categorised to CIM standards and as such meet the requirements of the Canadian National Instrument 43-101 and can be viewed as current. Analysis has shown that the resource estimation approach using a selective mining unit (SMU) model calculated using uniform conditioning of ordinarily kriged parent blocks by way of uniform conditioning is valid and a representative way of modelling the Certej Mineralisation. Comparison of the resource to internally produced polygonal estimate validates the SMU estimate. The Resource estimate is robust and has been completed several times by RSG with very similar results as well as having been checked internally twice with a less than 3% difference in volume.


Figure 19-1: Certej Lithology Model – West East Cross Section


Figure 19-2: Grade Tonnage Curves for the UC and OK Certej Estimates


Figure 19-3: Plot showing East Zone Intercepts versus Grade Model


Figure 19-4: Comparison of Drillhole Composite and SMU Model Grades, East Zone.


Figure 19-5: Comparison of 2005 and 2007 Certej SMU Resource Estimates


Lower North dump wireframe, plus x-sectional polylines and tie lines snapped to drill holes.

Upper North dump topographic wireframe cut from DTM, plus drill hole collars.

South dump wireframe, plus x-sectional polylines and tie lines snapped to drill hole

Upper South dump topographic wireframe cut from DTM, plus drill hole collars

Figure 19-6 North and South Dump Wireframes


20. OTHER RELEVANT DATA AND INFORMATION

20.1. Mining Considerations

20.1.1. Geotechnical Analysis

Golder was commissioned by EGL to conduct a study of the overall slope angles for the proposed Certej open pit mine in Romania. Programmes of mapping, underground and surface sampling, diamond core drilling and RC drilling have recovered information on the prospect since 2002. The study is based on information available at the time of the initial review in November 2005, additional information from the geotechnical drilling programme completed in February 2006, as well as other information provided by the mine. Golder completed and presented the findings of the study in the document titled “Report on Slope Design Certej Open Pit Deva, Romania”, in March 2006. Golder is satisfied that the level of geotechnical information currently available is sufficiently robust to meet the requirements of a Feasibility Study. The information used in the geotechnical assessment of the area and, in particular, the information relating to the rock mass quality and intact rock strength within the body of the open pit, are as follows:

• the geological interpretation of the ore body throughout the area of the open pits • the database of geological and geotechnical logging of the exploration cored

boreholes • the rock mass as exposed in the faces of the existing open pit • the rock exposures in the locality of the proposed open pit • the geometry of the existing Coranda open pit slopes • the geomorphology, particularly the gradient of the natural slopes in the area • groundwater • geotechnical logging of geotechnical borehole cores • geotechnical re-logging of select exploration boreholes; and, • a programme of geotechnical sampling and testing of cores from the geotechnical

programme. The level of data was good and enabled confidence in accurately defining the parameters then used in the design of a geotechnical model for the mine comprising 3 areas, sub-divided into 15 sectors according to the angle of the structures, rock type and weathering, to design pit slopes ranging between 30° and 60° mitigating the risk of pit wall failure.

20.1.1.1. Major Structures

The major structures crossing the area of the proposed open pit are: • North–south and north-northwest – south-southeast (NNW-SSE) trending near

vertical faults or shear zones. These zones tend to dip towards the WSW to the west of the Baiaga andesite and to the ENE to the east of the Baiaga andesite

• East–west near vertical faults or shear zones, associated with the emplacement of the andesite bodies; and

• shear zones sympathetic to the bedding deformation primarily as a result of the emplacement of the Baiaga andesite.


The rock mass associated with the north–south and NNW-SSE trending structure was observed in the upper levels of the existing open pit to be more fractured than the surrounding rock, with block sizes of the order of 150 to 200 mm, compared with the nearby rock mass of 300 to 400 mm. East–west fault zones were not observed in the existing open pit and are associated with the emplacement of the andesite bodies. The shear/fault zones associated with the emplacement of the Baiaga andesite trend essentially parallel to the dip of the bedding in both, the Neogene and the Cretaceous sediments, dipping to all quadrants down the sides of the intruded andesite. These structures are frequently associated with polymictic and monomictic breccias lensoid in form extending up to 30m in width and associated with the major NNW-SSE trending shear zones.

20.1.1.2. Minor Structures

Face Exposures The data collected from the Coranda open pit, and from the adjacent exposures, indicates a dominant joint set trending 35º to 215º dipping steeply both towards 125º and 305º. Other joint sets present had dip and dip direction of 80º/240º and 65º/85º. The block sizes ranged from 250 to 1000 mm based on the exposures in the upper levels of the existing open pit, with the majority of the blocks in the range from 250 to 400 mm. A plot and a rosette plot of all the pit data is illustrated in Figure 20-1 and Figure 20-2 respectively and the major concentrations of a number of joint and discontinuity sets are presented in Table 20-1. Table 20-1: Summary structural data from existing open pit

Set All data ≥ 3m Continuity ≥ 5m Continuity

Dip Dip Direction Dip Dip Direction Dip Dip Direction

1 74 159 75 159 76 163

2 62 171 62 169 61 173

3 47 131

4 83 075 79 076 84 072

5 65 054 67 054

6 33 019 31 018

7 81 238 83 237

8 68 200

9 74 195

10 84 276

11 66 027

The dominant discontinuity trends are aligned NW/NNW to SE/SSE, east-west and NE to SW. These structural orientations are consistent, at least in the NW/NNW to SE/SSE major structural trend that dominates the region.


Core Data The core from both, the exploration and geotechnical drilling, has been orientated. Manipulation of the data allowed the data to be presented in stereonet form, and the data from all the geotechnical boreholes and a selection of the re-logged boreholes were combined and plotted as illustrated in Figure 20-3. A rosette plot of the structural trends is illustrated in


Figure 20-4 and the major concentrations of a number of joint and discontinuity sets are presented in Table 20-2. Table 20-2: Summary structural data from geotechnical and re-logged

exploration boreholes

Set Surface Equivalent All Data

Dip Dip direction

A 2 59 161

B 77 127

C 84 114

D 4 80 090

E 5 61 064

F 80 062

The dominant discontinuity trends are aligned NW/NNW to SE/SSE, and NE to SW. These structural orientations are consistent, at least in the NW/NNW to SE/SSE major structural trends that dominate the region. However, the east–west trend evident in the surface data is not apparent from the borehole data as the borehole orientations are mostly sub-parallel. Conversely the NE to SW face exposure set and face are not evident in the major structures.

20.1.1.3. Rock Material Strength

The main rock types likely to form the side slopes of the open pit perimeter are exposed within the existing faces of the Coranda open pit. These include andesite, the Neogene sediments (sandstone and conglomerate), the Cretaceous sediments (sandstone and shale), and monomict and polymict breccias. These rock exposures provide useful information on the rock mass quality, and the impact of the tectonic events and structure on the rock masses within the area. In addition to the rock exposures within the open pit, information is available from the road cuttings in the locality of the open pit providing further information on the rock mass quality. This is particularly useful for the andesite where the degree of alteration varies rapidly within the rock mass. The Neogene sediments are essentially sandstone, conglomeratic sandstone and conglomerate with clasts typically up to 15 mm diameter in the conglomeratic sandstone and up to 50 mm in the more conglomeratic sequences. The units observed in the upper levels in the north of the Coranda open pit are generally fresh (unaltered) although there are zones where alteration of the sandstone has reduced the competent rock to a very weak rock. Brecciated zones were also evident adjacent to north-south trending structures with clasts of graphite, originating from the Cretaceous. Where unaltered sections of the Neogene sequence are present, a resistant and strong rock mass is formed as exemplified by the hill located at 345640E/501330N. The Cretaceous sediments are an intercalated sequence of sandstone and shale, frequently graphitic, with typical unit thickness of 1 m for each rock type, although units of both, sandstone and siltstone, are known to reach thicknesses of 5 m. The sandstone is typically moderately strong and the shale varies in strength between moderately weak to moderately strong, the difference generally being associated with the degree of silicification of the shale.


Close to the contact with the andesite intrusion, the Cretaceous sediments have been affected by the force of the intrusion, resulting in deformation of the sandstone and the shale such that the more competent and brittle sandstone exhibits boudinage, and the more plastic shale becomes deformed around the more competent sandstone blocks. The zone of deformation may be up to 20 m in thickness and is considered moderately weak. All three andesite intrusions are similar in rock type. The hydrothermal fluids have induced alteration of the andesite bodies such that the andesite exists in these affected zones as a weak to moderately weak rock, whereas in its unaltered state, is a fresh, strong rock. Close to the contact with the Cretaceous sediments, the andesite is altered, typically to a depth of about 20 m from the contact. In this zone, the andesite typically has a strength in the moderately weak range. Emplacement of the andesite bodies has led to the formation of mineralised breccias, in addition to hydrothermal fluids produced by hydrothermal breccias. In these localised zones, the breccia is frequently disturbed and altered, particularly close to the emplaced andesite, such that the rock is relatively weak. Outside of these zones, the breccia is silicified and is more, typically, a moderately strong rock.

20.1.1.4. Existing Open Pit Geometry

Data on the slope geometry of the existing open pit provided guidance on the potential for the overall pit slope angles at the proposed final depths. This data was compiled from topographic surveying data, and also from specific sections developed by the mine through the open pit faces. The sections developed through the open pit slopes, at some stage during the development, are presented in Table 20-3 and illustrated in Figure 20-5 to Figure 20-7. Table 20-3: Open pit slope angles and heights in existing pit

Slope Height m

Overall Slope Angle

20 80º

40 56º

60 45º

75 38º

95 30º

20.1.1.5. Gradient of Natural Slopes

Some of the slopes forming the topography in the area are relatively steep. These slopes have developed through geological time and may provide an indication of slope angles that may be developed in the open pit, or, at least indicate a lower bound for rock masses in the locality. Measurements of natural slope gradients in the general area of the open pit indicate slopes ranging from 30º to 33º, locally up to 35º in the west pit area, and approximately


27º in the east pit area. The rock types for the areas where the measurements were made are a combination of the main rock types encountered in the exploration drilling.

20.1.1.6. Geotechnical Model

The open pit surface used in the geotechnical model analysis is some 1.3 km in length, in the east–west direction, and some 0.5 km in width, in the north–south direction. This pit includes western, central and eastern sub-pits / phases within the perimeter of the overall pit and, with the detail of the data facilitating considerable accuracy, has been subdivided into sectors encompassing the three sub-pits (West, Central and East) illustrated in Figure 20-8 which have been further sub-divided into sectors (6 for the West pit, 4 for the Central pit, and 5 for the East pit) as presented in Table 20-4 to Table 20-9 inclusive. The sectors represent discernibly different zones of geotechnical conditions. Table 20-4: West pit typical final pit slope heights

WEST SUB-PIT

Sector Alignment Average Crest

Elevation m

Base Elevation

m

Height

m

1 E-W 550 330 220

2 NNW-SSE 575 330 245*

3 SSW-NNE 535 330 205*

4 E-W 520 330 190

5 NW-SE 450 330 220

6 NNE-SSW 435 330 105

* Current surface elevation (may not be crest elevation within pit)

Table 20-5: West pit typical geology at final pit profile

WEST SUB-PIT

Sector Typical Geological Conditions

1 Almost exclusively within the Hondol andesite with possibly some polymictic

breccia and Neogene sediments associated with north-south near vertical

shear zones

2 Highly weathered sedimentary rocks near surface underlain by Hondol

andesite to final depth

3 Highly weathered sedimentary rocks near surface underlain by Hondol

andesite to final depth

4 Predominantly Hondol andesite with minor presence of breccia, sandstone and

undifferentiated sediments and shale

5 Predominantly Hondol andesite with breccia present with andesite in lower

half of profile

6 Hondol andesite throughout profile

Table 20-6: Central pit typical final pit slope heights


CENTRAL SUB-PIT


Elevation

m

Base

Elevation

m

Height

m

7 NE-SW 555 450 105

8 E-W 545 450 95

9 NNE-SSW 555 450 105

10 NW-SE 510 450 60 * Current surface elevation (may not be crest elevation within pit)

Table 20-7: Central pit typical geology at final pit profile

CENTRAL SUB-PIT


7 Moderately and slightly weathered Neogene and Cretaceous sediments

overlying fresh Cretaceous sediments and andesite at depth

8 Sandstone overlying interbanded breccia and sandstone, andesite and breccia

and finally breccia and sandstone to depth

9 Sandstone in upper third of profile overlying sandstone/breccia with breccia

towards toe of slope

10 Breccia near surface overlying sandstone that forms the main part of the

slope with shale and andesite towards toe of slope

Table 20-8: East pit typical final pit slope heights

EAST SUB-PIT


Elevation

m

Base

Elevation

m

Height

m

11 NNE-SSW 555 330 225

12 E-W 580 330 250

13 N-S 585 330 255

14 E-W 520 330 190

15 WNW-ESE 520 330 190 Table 20-9: East pit typical geology at final pit profile


EAST SUB-PIT


11 Interbanded sequence of sandstone, breccia, tuff and shale dipping out of

slope

12 Predominantly undifferentiated sediments and sandstone, conglomerate with

minor bands of andesite

13 Completely and highly weathered andesite with voids near surface overlying

generally unweathered andesite to depth with breccia and sandstone towards

toe of slope

14 Undifferentiated sediments, sandstone with minor conglomerate followed by

breccia and, shale and undifferentiated sediments to depth

15 Undifferentiated sediments, sandstone with minor conglomerate followed by

breccia and, shale and undifferentiated sediments to depth

20.1.1.7. Groundwater

There seems to be no evidence of major seepage from the existing pit faces, or any operational requirement to remove groundwater from the floor of the existing open pit, which is at an elevation of approximately 475 m. Inspection of the existing open pit faces indicates the presence of occasional seepage points, but these may be from surface water, or snowmelt, passing through the rock mass and seeping through the pit faces. During the programme of exploration drilling and the geotechnical drilling, there seems to be no noticeable water strikes recorded, including a number of boreholes drilled to a lower level than the base of the proposed open pit at an elevation of some 330 m. An adit was developed through the central area of the existing open pit, aligned essentially WNW-ESE, at an elevation of approximately 410 m. The adit portal is located towards the northwest limit of the Hondol andesite. The first 400 m of the adit is aligned NW-SE, is approximately 5 m high by 5 m wide, is shotcrete lined, and passes through the Hondol andesite. The unlined section of the adit has a sectional area of some 7 m2, is some 800 m in length, extends from 345700E to 346470E, and has spurs to the north and south off the main adit. One exploration borehole that passed through the adit had water to the lip of casing that stands some 1 m above the adit floor, with some groundwater seeping from the open borehole in the sidewall of the adit. Very little seepage collects in the central drain within the adit from the surrounding rock mass. A valley with stream flow runs approximately north–south immediately to the west of the western end of the proposed open pit perimeter adjacent to the adit entrance, some 5 to 10 m below the adit invert. The valley bottom in this area lies at an elevation of some 400 m. Low ground, at an elevation of some 410 m, also exists towards the southwest of the proposed open pit. Drainage has been designed to ensure the stream water does not enter the West pit. In summary, the indications are that the groundwater level within the hillside will have a base elevation at the level of the surrounding low points in the valleys adjacent to the


proposed open pit, a level likely to be around 370.m to 390 m. This is below the base of the Central pit (450 m) and 40 m to 60 m above the base of the East and West pits (330 m). Storage within the hillside draining to the base of the East and West pits is not anticipated to be high based on the evidence from the drainage into the adit beneath the existing open pit.

20.1.1.8. Exposed Rock Mass Conditions

Evidence of the strength and stability of the rock mass is indicated in the existing open pit faces and also in the underground adit. Existing Open Pit The slope faces within the Coranda open pit have been developed in the upper 100 m of the hillside that have been both weathered by sub-aerial effects and, to a much lesser extent, altered by hydrothermal solutions passing through the rock mass. Consequently the upper 20 m to 30 m of the rock mass tends to be of a consistently low strength correlating to the new pits which contain an estimated 9.7 Mt of weathered material. Below the upper surface zones, the rock mass strength is generally a function of the degree of alteration, but also the rock type and the rock mass structure. The shale horizons within the Cretaceous are generally more susceptible to deterioration than the sandstone and siltstone, but where the shale has been silicified, the shale becomes a competent unit. Individual faces have been formed at near vertical angles and have ravelled back to lower angles as a result of the breakdown of the weaker materials and the dislodging of individual blocks within the rock mass. The competent sandstone and siltstone within the Cretaceous have a strongly developed joint set that lies orthogonal to the bedding structure. As much of the bedding in the open pit dips into the slope face, then individual blocks have tended to slide off the face and these have ultimately controlled the face profile. A considerable amount of qualitative data was obtained but no detailed survey information is available on the individual face profiles and many sections of the faces have been obscured by backfilling as upper faces have been developed and the waste rock pushed over the crest. This detail data is not critical. Adit During the inspection of the adit, it seemed that:

• there is no evidence of instability within the section of the adit that has been driven through the Hondol andesite, in the west. This section of the adit is approximately 5 by 5 m and is shotcrete lined, but where limited exposure allowed, the rock mass was tight with the andesite of medium strength, and joint spacing ranging from 200 to 400 mm;

• the section of the adit passing through the Cretaceous and breccia, that lies between the Hondol andesite and the Baiaga andesite, has localised areas of roof instability where shears have passed through the sequence;


• the section of the adit that passes through the Baiaga andesite shows little evidence of instability with the rock mass tight and of medium strength, and joint spacing ranging from 200 to 300 mm;

• the adit passes through the Baiaga andesite into the Cretaceous units until the Sacaramb andesite in the east. The majority of the adit is stable, but there are a number of sections where the shears from the major structure pass through the sequence and have caused localised instability at a number of sections along the adit. The shear material has intra-formational shears and is relatively weak and friable; and,

• one localised collapse had effectively blocked the adit at the interface between the Cretaceous and the Sacaramb andesite and, as a result, it was not possible to inspect the condition of the andesite. However, the andesite apparently is similar in quality to that observed for the Hondol and the Baiaga andesites.

In summary, as exposed within the adit, all three andesite masses are similar in quality with the impact of the shears that pass through the Cretaceous sediments evident.

20.1.1.9. Mine Rock Mass Rating Model

The rock mass quality has been assessed using Bieniawski’s Rock Mass Rating (RMR) and Hoek’s Geological Strength Index (GSI). The classifications are based on intact rock strength, degree of fracturing, condition of the fractures and groundwater. The GSI is based on a variant of the RMR, where the weighting for the various parameters is different. In each case, the classifications range from 0 to 100, where 100 is the best quality rock mass. Borehole Assessments In order to assess the rock mass quality throughout the length of each of the geotechnical boreholes and the re-logged exploration boreholes, the RMR and GSI have been estimated throughout the complete length of the borehole. These rock mass quality data have then been combined over sections of the borehole to provide a summary table for each borehole. In the estimation, no adjustment was made for orientation of structure or the presence or absence of groundwater. Orientation of structure is dealt with in the overall stability analysis for major structures and for minor structures is considered in the assessment of block failure for bench stability. As the slope stability analyses include a groundwater profile, the determination of the rock mass quality assumes the rock mass is dry. Otherwise, the stability analysis would “double count” the impact of groundwater. The summary of the rock mass quality for each of the geotechnical boreholes is presented in Table 20-10. Table 20-10: Summary of rock mass quality and geology for geotechnical boreholes

Borehole Depth Down Borehole - m Rock Mass Quality - RMR

CJGD001 0 to 77m

77m to 208m

POOR (RMR = 26)

FAIR locally POOR and locally GOOD

(RMR 46, ranging from 30 to 70)

CJGD002 6m to 277m FAIR (RMR = 57) locally FAIR and locally GOOD

CJGD003 0 to 78m

78m to 127m

FAIR (RMR = 53)

FAIR (RMR = 43)


CJGD004 0 to 18m

18m to 270m

VERY POOR (voids and low strength)

FAIR (RMR = 55)

CJGD005 5m to 167m GOOD (RMR = 69)

CJGD006 0 to 16m

16m to 89m

89m to 153m

POOR (RMR = 25)

FAIR (RMR = 53)

FAIR (RMR = 52)

CJGD007 0 to 13m

13m to 85m

85m to 242m

VERY POOR to FAIR (RMR = 19 to 45)

FAIR (RMR = 52)

FAIR/GOOD (RMR = 60)

CJGD008 0 to 8m

8m to 95m

95m to 255m

POOR and GOOD (RMR = 13 & 47)

FAIR (RMR = 56)

GOOD (RMR = 61)

For the re-logged exploration boreholes, the rock mass quality data is presented in Table 20-11. The overall quality of the rock mass is FAIR locally POOR, particularly in the upper sections of some of the boreholes, and locally GOOD, generally, but not exclusively, in the lower sections of the boreholes reflecting the expected properties of the cored material and confirming the observations and measurements made in the ‘geotechnical boreholes’ Table 20-11: Summary of rock mass quality and geology for re-logged

exploration boreholes


CJGD110 0 to 180m

180m to 353m

Not cored

FAIR (RMR = 57)

CJGD117 0 to 40m

40m to 334.3m

Not cored

GOOD (RMR = 62)

CJGD160 0 to 78m

78m to 127m

FAIR (RMR = 53)

FAIR (RMR = 43)

CJGD183 0 to 90m

90m to 186m

186m to 458.2m

Not cored

GOOD (RMR = 69)

GOOD (RMR = 73)

CJGD184 0 to 39m

39m to 58m

58m to 245m

FAIR (RMR = 40 to 57) with voids

GOOD (RMR = 68)

GOOD (RMR = 65)

CJGD188 0 to 16m

16m to 89m

89m to 153m

POOR (RMR = 25)

FAIR (RMR = 53)

FAIR (RMR = 52)

CJGD189 0 to 13m

13m to 85m

85m to 242m

VERY POOR to FAIR (RMR = 19 to 45)

FAIR (RMR = 52)

FAIR/GOOD (RMR = 60)

CJGD193 0 to 8m

8m to 95m

95m to 255m

POOR and GOOD (RMR = 13 & 47)

FAIR (RMR = 56)

GOOD (RMR = 61)

CJGD194 12m to 103m FAIR (RMR = 56)



103m to 221m

221m to 310m

GOOD (RMR = 70)

GOOD (RMR = 68)

CJGD195 0m to 300m GOOD (RMR 63)

CJGD211 0 to 40m

40m to 320m

Not cored

GOOD (RMR = 63)

CJGD212 0 to 60m

60m to 137m

137m to 260m

Not cored

GOOD (RMR = 70)

GOOD (RMR = 72)

CJGD216 0 to 33m

33m to 101.4m

FAIR (RMR = 56)

GOOD (RMR = 65)

CJGD220 0 to 74m

74m to 110m

GOOD (RMR = 61)

GOOD (RMR = 63)

CJGD222 0 to 84m

84m to 89m

GOOD (RMR = 65)

FAIR (RMR = 50)

CJGD230 0 to 100m

100m to 270m

Not cored

GOOD (RMR = 75)

CJGD232 0 to 120m GOOD (RMR = 63)

CJGD238 0 to 11m

11m to 82m

82m to 114m

POOR (RMR = 29)

FAIR (RMR = 55)

FAIR (RMR = 54)

The mean RMR for the various rock types encountered in the geotechnical boreholes are presented in Table 20-12. The mean Rock Mass Rating values are from the geotechnical boreholes and exclude the upper, weaker section from each of the geotechnical boreholes. The depth of the borehole data excluded ranges from none (borehole CJGD003) to 59m (borehole CJGD001) with the “typical” range of exclusion for the boreholes from 10m to 20m. Table 20-12: Summary of rock mass for main rock types

Rock type Rock Mass Rating (Mean*)

Sandstone, conglomerate, undifferentiated

sediments and shale

FAIR (RMR = 52)

Breccia GOOD (RMR = 61)

Andesite FAIR (RMR = 59)

* Mean excludes weathered material

Rock Mass Shear Strength The strength of the rock mass has been determined from the Hoek-Brown failure criterion. This strength criterion was developed as it is practically impossible to carry out triaxial, or shear tests, on rock masses at a scale which is appropriate to surface excavations in mining engineering. The Hoek-Brown failure criterion is based on the GSI, which is related to RMR. The value of the GSI ranges from 10, for extremely poor rock masses, to 100, for intact rock, and is related to the RMR as follows:

GSI = RMR89 – 5 where RMR89 is the rating system published in 1989 by Bieniawski


The GSI is used together with the Uniaxial Compressive Strength (UCS) of the rock and a material constant (mi) to determine the shear strength of the rock mass. The GSI values were checked with inspection of the cores, and an examination of the core photographs, to allow modification of the GSI, if required, where the rock mass classification systematic method has not properly reflected the rock mass quality. The rock mass shear strengths determined are different for each geological unit and for each sector of the open pit. Inevitably, there is a wide range in the rock mass shear strengths. The mean shear strengths for the rock mass are summarised in Table 20-13 and Table 20-14. Table 20-14 includes shear strength data based on an arbitrary zoning through each sector of the open pit; a zone representing the upper part of the open pit slope; an intermediate zone; and, a lower zone towards the toe of the open pit slope. Table 20-14 includes summary information for the mean data according to rock type. These data, like the depth zoning data, include a wide range of shear strengths for the various geological units. Table 20-13: Rock mass shear strength for major depth zones within open pit

Rock mass shear strength

Upper zone*(15)** Intermediate zone*(13)** Lower zone*(7)**

c (kPa) (º) c (kPa) (º) c (kPa) (º)

250 31 625 33 725 33 *Arbitrary depth zoning for all sectors of open pit ** Number of sets of rock mass shear strength data

Table 20-14: Rock mass shear strength for main rock types within open pit

Rock mass shear strength

Andesite (16) ** Sediments (16) ** Breccia (3) **

c (kPa) (º) c (kPa) (º) c (kPa) (º)

475 34 450 29 700 38 ** Number of sets of rock mass shear strength data

20.1.1.10. Slope Stability Assessment

Based on the geotechnical and exploration boreholes relevant to each of the sectors around the perimeter of the proposed open pit, a geotechnical/geological model, or typical cross section of the geology, was prepared. For each zone within the section, shear strength was determined and allocated in the slope stability analysis. Slope Stability Models Representative cross sections were developed for each of the sectors, and shear strengths applied to each of the rock mass zones within each section. These sections have been simplified but are considered to be a reasonable representation of the


conditions in each sector, although each sector is typically 200 to 300 m in length, and as such, there will be variation in the geological conditions and so the geotechnical conditions along the strike length of each sector. A groundwater profile was included in each of the slope stability models, reflecting the assumed drawdown profile as the open pit excavation develops. In each model, the groundwater profile was assumed to drain towards the toe of the slope, from a level within the open pit abutment at an elevation of some 370 to 380 m. Slope Stability Analysis The SLOPE/W software, developed by Geoslope, was used to perform the slope stability analysis. Each sector was analysed and the factor of safety (FoS) reported for both, the overall slope and the upper section of slope. In general, the upper section of the slope is weaker than the rock mass at depth and, as such, the FoS for this section needs to be considered in addition to that of the overall slope. In the profile, the model did not include individual bench faces, berms or ramps, although in some of the sectors, a berm was introduced into the model to effectively separate the upper, weaker section of the slope from the lower, stronger section, so that the impact of the weaker section on the overall slope stability is reduced. In the models, the slope was analysed using the circular failure mode, but where it was considered that a specific major structure, or discrete weaker unit, could impact on stability, then block failure analysis was conducted to assess the stability with failure along this particular structure. Slope Stability Results

The FoS results for each of the sectors is presented in Table 20-15 and the individual sector centre lines shown in Figure 20-8. The analyses performed are for overall stability without the introduction of ramp(s), or individual benches, although individual berms were introduced into the profiles for some sectors to meet the target FoS of 1.5, for overall, upper and lower sections of the slope profile. Table 20-15: Slope stability assessments for each sector


Sector

Slope

Height

m

Slope Profile

Factor of Safety

Upper Slope Overall Slope

1 170 60º lower & 53º upper, interface at

430m elev.

1.46 2.16

2 245 43º overall N/A 1.49

3 205 45º overall N/A 1.56

4 190 42º overall 1.46 1.49

5 120 30º in lower slope & 40º in upper,

interface at 365m elev.

N/A 1.50


375m elev.

1.60 1.87


525m elev.

1.95 1.96

8 95 45º lower & 30º in upper, interface at

515m elev.

1.58 1.53


500m elev.

1.53 2.01

10 60 50º overall, stability controlled by

weak horizon dipping out of slope

N/A 2.08

11 225 40º in upper slope (dictated by ore

grade requirements) with 51º in lower

slope, interface at elev. 400m

1.58 (lower

slope)

1.59 (main

part of slope)

12 245 55º lower and 30º in upper with 15m

berm at interface, elev. 485m

1.41 1.55

13 260 45º in lower and 40º in upper with

25m berm at interface elev. 465m

1.50 1.51



1.32 1.49



1.32 1.49

20.1.1.11. Bench Stability

Stereonet Analysis An analysis of the structural orientation data, in addition to the mapping carried out at the 510, 525 and 540 m levels, within the Coranda Open Pit, was used to provide stereonet plots for each of the sectors. Each stereonet shows, in addition to the contour plot of the structure, the following:

• A bench face, nominally drawn with a 70º profile; • The daylight envelope for the 70º bench face; • A 30º friction cone for the 70º bench face; and, • The toppling envelope for a nominal 45º overall slope and a bench face of 70º.


The daylight envelope indicates those joints that have the potential to slide out from the bench face, and the friction cone, assuming a typical friction angle for the joint and bedding planes of 30º, limits those joints within the daylight envelope that could potentially slide out of the face. The toppling envelopes are defined by a frictional limitation and are represented by the two planes, one representing the nominal overall 45º slope and the second, the nominal 70º bench face. There is also a strike limitation for toppling to occur, equivalent to ±30º. Thus, the toppling envelope for the 70º slope extends from plane 2 to the perimeter of the stereonet, and for the 45º, from plane 3 to the perimeter of the stereonet. In summary, the tendency for block failure in the open pit slopes, specifically the bench faces, is presented in Table 20-16. The structural analysis, on a sector by sector basis, is sometimes based on limited data, there being no representative surface data or possibly the data from only one borehole used to form the data set for an individual sector. However, where possible, the data from exploration boreholes in the vicinity of each of the sectors has been used to augment the data from the geotechnical and the re-logged exploration boreholes. The summary presented in Table 20-16 indicates that throughout the proposed open pit, there are few sectors where there is little potential for planar/wedge failures with joint concentrations relatively low (Sectors 1, 2, 3, 8, 9, 10, and 11), while there is no indication of planar/wedge failure potential in the remaining sectors. Similarly, there are few sectors where there is potential for toppling failure, either in the overall slopes or in the bench faces, with the most notable indication of toppling potential in Sector 9. The indicated typical bench profiles range from 70º to 80º, with the overall range from 60º to 90º. Table 20-16: Block failure assessment for each sector

Sector Planar/wedge failure Toppling failure

Bench face profile

Overall slope Bench face

1 Yes : very low

concentration at 53º

towards SSW

No Yes : Minor

concentration

70º - 80º

2 Yes : Minor concentration Yes : Minor

concentration

Yes : Minor

concentration

70º

3 Yes : Minor concentration

dipping at 45º out of face

No Yes : Minor

concentration

70º - 80º

4 No Yes : Minor

concentration

Yes : Intermediate

concentration

65º - 75º

5 No No No 70º - 90º



8 Yes : Minor concentration,

57º towards SSW

No No 70º - 90º

9 Yes : Minor concentration Yes : Minor

concentration

Yes : Intermediate

concentration

60º - 70º


Sector Planar/wedge failure Toppling failure

Bench face profile

Overall slope Bench face

10 Yes : Minor concentration No No 70º - 90º

11 Yes : Minor concentration No No 70º - 75º

12 No No Yes : Minor

concentration

70º - 75º

13 No No Yes : Minor

concentration

75º - 80º

14 No Yes : Minor

concentration

Yes : Minor

concentration

70º

15 No Yes : Minor

concentration

Yes : Minor

concentration

75º

It is likely that for bench stability the dominant structure will be the orthogonal bedding set that will affect mainly Sectors 12, 14 and 15, where the Cretaceous units dominate the East pit.

20.1.1.12. Geotechnical Conclusions

Golder completed an independent geotechnical study on information and data obtained between 2002 and 2006. The data was sourced from exploration drilling; geotechnical boreholes specifically drilled for the study; outcrop mapping; underground mapping in an adit running at depth through an axis of the orebody; and existing open pit faces. Golder is satisfied that the level of geotechnical information currently available is sufficiently robust to meet the requirements of a Feasibility Study and RSG concur with this statement. The main rock types that will form the final open pit slopes are Neogene sediments (conglomerate and sandstone), Cretaceous sediments (intercalated sandstone and shale), andesite and breccia. Alteration of all rock types has occurred in zones with competent sandstone reduced to a weak rock and the andesite from a strong rock to locally soil strength. Major structural trends are NNW-SSE and E-W with polymictic breccia and monomictic breccia associated with these major shear/fault zones. Jointing throughout the rock mass is evident, some associated with the major structure, where exposure in the open pit indicates more intense fracturing of the rock mass adjacent to the shears/faults. For the structure mapped within the Coranda Open Pit, the main joint structural trends are NW-SE, NE-SW and E-W with the most continuous structure dipping ≈60º towards the south. The main joint structural trends from the borehole data show a slightly different distribution with NNE-SSW, NW-SE and NE-SW trends. Major structure is the faulting and shears that trend NNW-SSE dipping predominantly towards the east but also towards the west.


The groundwater level within the hillside will most likely have a base elevation at the level of the surrounding low points in the valleys adjacent to the proposed open pit, a level of around 370m to 390m elevation. Storage within the hillside is not anticipated to be high based on the evidence from the drainage into the adit beneath the existing open pit. A programme of geotechnical boreholes has been completed in addition to the re-logging of a selection of exploration cores. A sampling programme of the geotechnical cores provided index data on the unit weight, moisture content and Uniaxial Compressive Strength of the rocks. The mean strength of all samples tested was 20.3 MPa with a maximum of 57.4 MPa and a minimum of 0.54 MPa. The majority of the samples have a strength of less than 20 MPa with an overall mode of 5 MPa to 10 MPa. The individual rock types, in general, show a similar trend with modes of 15 MPa to 20 MPa for andesite, 5 MPa to 10 MPa for the sandstone/conglomerate and 10 MPa to 15 MPa for the breccia. There were insufficient data for a mode to be defined for the undifferentiated sediments/shale. The data from the logging and laboratory testing have allowed assessments of the rock mass quality for each of the cores to be carried out. The overall rock mass quality, based on RMR, of the rock mass is FAIR locally POOR, particularly in the upper sections of some of the boreholes, and locally GOOD, generally, but not exclusively, in the lower sections of the borehole. The proposed open pit has been sub-divided into 15 sectors and a typical section prepared for slope stability modelling. The shear strength of the individual units comprising each section has been determined from the rock mass quality data using the GSI system and used as an input to the slope stability analyses. The analyses indicate that to achieve a target safety factor of 1.5 for the overall slopes, there is a range of stable slope configurations ranging from 55º (Sectors 1, 6 and 7) to 42º (Sector 4). An analysis of the joint structural data indicates that throughout the proposed open pit, there are few sectors where there is little potential for planar/wedge failures with joint concentrations relatively low (Sectors 1, 2, 3, 8, 9, 10, and 11), while there is no indication of planar/wedge failure potential in the remaining sectors. Similarly, there are few sectors where there is potential for toppling failure, either in the overall slopes or in the bench faces, with the most notable indication of toppling potential in Sector 9. The indicated typical bench profiles range from 70º to 80º with the overall range from 60º to 90º. It is likely that for bench stability the dominant structure will be the orthogonal bedding set that will affect mainly Sectors 12, 14 and 15 where the Cretaceous units dominate the East pit.

20.1.2. Pit Optimisation

RSG was commissioned by EGL to conduct an open pit optimisation analysis, based on the gold and silver grade estimates, using the Measured and Indicated mineral


resources, geological data, geotechnical data and densities for the proposed Certej open pit mine in Romania. The Whittle Four-X pit optimisation software (Whittle), developed by Gemcom Enterprise Mining Systems, was used for this task. The analysis is based on information available on site as of September 2007. RSG completed and presented the findings of the analysis in the document titled “Certej Gold Silver Project, Romania”, dated November 2007. The content in this section has largely been extracted from this document. RSG is satisfied that their technical report confirmed that the exploitation of the Certej Gold Silver deposit is economically viable.

20.1.2.1. Mining Approach

The Certej Project will be mined by conventional open pit, selective mining exploitation method. The mining method is a conventional open cut drill, blast, load and haul operation, using hydraulic excavators to carry out selective flitch mining, similar to many small open pit gold mining operations throughout the world, and off-highway dump mining trucks. Previous mining operations at this area, and other mining operations in similar conditions, indicate this type of equipment is appropriate and the most reliable option. The base case assumption is that mining and equipment maintenance will be carried out as owner operated, given the comparatively low local labour rates, the existing mining expertise within the country, and the requirement for selective mining. Drilling and blasting will be performed on 5 m high benches, with blasted material excavated in two discrete flitches, each nominally of 2.5 m height.

20.1.3. Mineral Resource Model

Whittle deals with the amount of metal in a block, not grade. Therefore, the grade of a block is estimated from the tonnage and the metal in the block. Consequently, the metal content of a block was estimated using the grade estimate derived from the UC grade tonnage estimate for gold, and the OK grade estimate for silver. The bulk density data, as presented in the mineral resource section, has been used to assign a mean density for each of the lithological units used to create the mineral resource block model in Whittle. The following densities were used:

• Cretaceous sediments 2.47 t/m3; • Hondol andesite 2.44 t/m3; • Neogene sediments 2.37 t/m3; • Baiaga andesite 2.44 t/m3; • Breccia 2.35 t/m3; • Dealul Grozii andesite 2.37 t/m3.

There was insufficient density data for the existing dumps and consequently a notional density value was used:

• Existing dump 1.80 t/m3.


20.1.3.1. Applied Geotechnical Considerations

Pit slope design is based on recommendations from an assessment of ground conditions by Golder, as summarised in the document titled “Slope Design, Certej Open Pit, Deva, Romania”, dated March 2006, and presented previously in this study. The information used in the geotechnical assessment of Certej and, in particular, the information relating to the rock mass quality and intact rock strength within the body of the open pit, are as follows:

• geological interpretation of mineralised rock throughout the area of the open pits;

• database of geological and geotechnical logging of the exploration cored boreholes;

• rock mass as exposed in the faces of the existing Coranda open pit;

• rock exposures in the locality of the proposed Certej open pit;

• geometry of the existing Coranda open pit slopes;

• geomorphology, particularly the gradient of the natural slopes in the area;

• groundwater;

• geotechnical logging of geotechnical borehole cores;

• geotechnical re-logging of selected exploration boreholes; and,

• a programme of geotechnical sampling and testing of cores from the geotechnical programme.

The recommended inter-ramp slope angles are between 30º and 60º, as presented in Table 20-17. These parameters assumed dry slope condition and, as a result, the cost of depressurisation holes has been included in the mining operating cost. Figure 20-8 shows a plan view illustrating the various slope sectors. Table 20-17: Slope profile summary

Sector Inter Ramp Slope Angle

1 60º lower & 53º upper, interface at rl 430 m

2 43º overall

3 45º overall

4 42º overall

5 30º in lower slope & 40º in upper, interface at rl 365 m



8 45º lower & 30º in upper, interface at rl 515 m


10 50º overall, stability controlled by weak horizon dipping out of slope

11 40º in upper slope (dictated by ore grade requirements) with 51º in lower slope, interface at rl

400 m

12 55º lower and 30º in upper with 15 m berm, interface at rl 485 m

13 45º in lower and 40º in upper with 25 m berm, interface at rl 465 m




20.1.4. Economic Input and Process Recoveries Parameters

Whittle requires revenue, mining and processing costs for the open pit optimisation analysis to be specified. These are used to determine the economic final pit. EGL provided all the optimisation input parameters. Nonetheless, RSG reviewed the processing and mining costs and found them acceptable for the open pit optimisation analysis. These input parameters cover a wide range of disciplines, and as a result, a number of specialists have been involved in their estimation. Table 20-18 presents a list of the main Whittle input parameters, and the specialists responsible for their estimation. Table 20-18: Whittle input parameters source

Input Parameter Source

Commodity price EGL

Mining Costs EGL with a review by RSG

Owner’s mining associated costs EGL with a review by RSG

Metallurgical and Processing EGL with a review by RSG

General and Administration cost EGL with a review by RSG

Geotechnical and Hydrology Golder with a review by RSG

Governmental EGL

Table 20-19 presents a summary of the main Whittle economic input parameters. All references to monetary values are denominated in United States of America dollar (USD), unless specifically stated otherwise.

Table 20-19: Whittle economic input parameters and metallurgical recoveries

Item Unit Value

Gold price USD/oz 425

Silver price USD/oz 7.0

Average mining cost USD/t 1.23

Processing cost to produce gold and silver doré USD/t milled 10.00

Mine supervision USD/t milled 0.35

General and Administration cost USD/t milled 0.33

Geotech / De-watering cost USD/t milled 0.05

Crusher feed USD/t milled 0.11

Processing metallurgical recovery

- Main Au

Ag

- West Au

Ag

%

81.9

76.4

78.8

70.1

State government royalty % of revenue 0.8

Refining cost % of revenue 1.85

Mining recovery % 95.0


20.1.4.1. Pit Optimisation Setup

The open pit optimisation analysis was carried out for a wide range of gold prices, from as low as USD 100/tr.oz to a maximum of USD 1,000/tr.oz. Figure 20-9 illustrates the pit size in relationship to the gold price, based on Measured and Indicated mineral resources only. The Whittle financial analysis was performed based on the following assumptions:

• mill feed production of 3.0 Mtpa; • sufficient waste is removed each period to enable the required processing rate to

be maintained, therefore, the operation is mill limited; and, • discount rate of 10% real.

Whittle uses the revenues, costs and ore process rate to estimate project cash flows. The cash flow is reported both, undiscounted and discounted. These cash flows are calculated pre-tax, with all costs and revenues in real terms, with no inflation or escalation. The effect of applying the discount rate is to reduce the impact of future cash flow and to emphasise the importance of cash flow generated in the early years of the project life. Typically, the three cash flows generated are:

• Undiscounted Operating Cash Flow; • Best Case Discounted Operating Cash Flow – each incremental pit is removed

prior to advancing to the next adjacent incremental pit. The cash flow schedule is the equivalent of mining multiple push backs; and,

• Worst Case Discounted Operating Cash Flow – each bench is mined out prior to moving to the next bench, using the optimisation block height as the default bench height. The cash flow schedule is the equivalent of top down ‘flat’ mining.

The optimum pit shell is usually chosen by inspecting these cash flows and selecting the pit shell with the maximum total cash flow. The maximum undiscounted cash flow is the pit shell where the incremental pit is breaking even, and is, therefore, the maximum economic pit in today’s revenue/cost terms. Depending on the size of the pit shell, when the discount rate is used, this pit shell with the maximum discounted cash flow is generally somewhat smaller. However, this smaller ‘final pit’ will be more profitable. An actual mining schedule will most likely lie between the two extremes of Worst Case and Best Case, and is modelled using an average of these two extremes. The cash flows, as described above, should be used for pit optimisation comparison purposes only. No Net Present Value (NPV) should be derived from these cash flows.

20.1.4.2. Pit Optimisation Results

Based on Measured and Indicated mineral resources only and at gold and silver prices of USD 425 and 7.0/tr.oz respectively, the optimum pit shell considering the maximum average discounted cash flow is 16. This computer optimised pit contains some 31.5 Mt of mill feed at an average grade of 2.1 g/t of gold and 13.1 g/t of silver for approximately 1.7 Moz of recoverable gold metal and 9.9 Moz of recoverable silver metal. Some 102.2 Mt of waste are contained within this computer generated pit shell for a waste stripping ratio of 3.3:1 (t:t). The undiscounted operating cash flow is USD


263 million. The Worst Case discounted cash flow is $137 million, whilst the Best Case discounted cash flow is $172 million. To ascertain the likely discounted cash flow derived from a more realistic mine production schedule, RSG estimated the average discounted cash flow as USD 154 million. Figure 20-10 illustrates a summary of the pit optimisation results.

20.1.5. Sensitivity Analysis

The physical pit envelope is such that any remaining ore is located in the east of the pit and is, effectively uneconomically mineable due to the terrain which would require the stripping into a hillside. The increase in waste stripping in relation to additional ore is clearly shown in Figure 20-11. The base case assumed a gold price of USD$425/oz and a silver price of USD$7/oz. A variation was run at USD$600/oz for gold and USD$7/oz for silver. Pit 16 (selected) appeared to be optimal, at USD$425, for a Base Case Scenario using the Maximum Undiscounted Cashflow as the selection criteria. The same pit is optimal at USD$600 using the Maximum Average Discounted Cashflow as the selection criteria. Pit 16 lay mid way between the best and worst cases. RSG considered Pit 16 a robust selection.

20.1.6. Engineered Pits

RSG was commissioned by EGL to conduct detailed pit design for the selected optimum pit shell 16. The content in this section has largely been extracted from the RSG document titled “Certej Gold Silver Project, Romania”, dated November 2007. Detailed pit slope parameters, comprising of batter angles and berm widths, used for the pit design are presented in Table 20-20. These pit slope parameters were based on the description of inter-ramp slope angles presented earlier in this section. Table 20-20: Pit slope parameters

Sector

Batter Angle

Berm Width m

Sector

Batter Angle

Berm Width m

1U 75º 6 9U 55º 16 1L 80º (rl 430m) 6 9L (rl 500m) 80º 6 2 60º 6 10 75º 6 3 60º 6 11U 60º 8 4 60º 8 11L (rl 400m) 70º 6 5U 60º 8 12U 55º 16 5L 55º (rl 365m) 16 12L (rl 515m) 75º 6 6U 75º 6 13U 60º 8 6L 80º (rl 375m) 6 13L (rl 465m) 60º 6 7U 75º 6 14U 50º 18 7L 80º (rl 525m) 6 14L (rl 470m) 70º 8 8U 55º 16 15U 50º 18 8L 65º (rl 485m) 6 15L (rl 470m) 70º 8


Sectors 1 to 6 pertain to slope angles in the western region of the pit. Similarly, Sectors 7 to 10 pertain to the central region of the pit, and Sectors 11 to 15 pertain to the eastern region of the pit. In addition, Golder recommended an extra wide berm of 25 m for Sector 13 at rl 465 m, as well as a 40 m wide berm for Sectors 14 and 15 at rl 470 m to separate particular upper weaker zones from lower stronger zones. However, based on a review by RSG, such wide berm of 40 m was not considered necessary, and for Sectors 13, 14 and 15, a berm width of 18 m was maintained from rl 470 m to the natural surface. Catch berms were inserted every 20 m, in vertical wall height. A 24 m wide, dual access ramp was incorporated in the detailed mine design. This is more than sufficient for a 62t off-highway dump truck based on allowing a safe operating width of 3 truck widths plus a windrow. The ramp gradient is set at 1 in 10, or 10%. A 12 m wide, single lane access ramp is utilised for the deeper parts of the pit. A minimum mining width of 30 m was assumed for both pits. This width suits the excavator loading radius and allows sufficient room for truck turning and positioning. Table 20-21 presents a summary of the material breakdown as contained within the final pit designs, excluding Inferred mineral resources. There are Inferred mineral resources within this engineered pit designs, which are classified as waste, totalling some 0.8 Mt at 1.6 g/t for gold and 8.9 g/t for silver. Table 20-21: Engineered pit material inventories

Area

Total Mt

Waste Mt

Ore Mt

Gold g/t

Silver g/t

Gold Moz

Silver Moz

SR t:t

East Pit 95.1 72.2 22.8 2.1 11.0 1.5 8.1 3.2

West Pit 38.1 28.1 10.0 1.8 12.4 0.6 4.0 2.8

Total 133.1 100.3 32.8 2.0 11.4 2.1 12.0 3.1

The engineered pits generally honour pit shell 16 well. Waste tonnes decreased by 1.9%, from 102.2 to 100.3 Mt, while ore tonnes increased by 4%, from 31.5 to 32.8 Mt. Overall excavated material was reduced by 0.4% to some 133.1 Mt. The net effect of the engineered pits on the contained metal is +1.3% for 2.1 Moz of contained gold and -10.4% for 12.0 Moz of contained silver. This reduction in waste material and increase in ore reduces the 3.3:1 strip ratio of the computer generated pit to 3.1:1 for the engineered pit.

The East pit is the largest containing some 71% of the total excavated material for the Certej project. It is also the pit with the best ore processing metallurgical recoveries. Therefore, commencing excavating activities with the East pit provides the best project cash flow in the earlier years. Considering the large size of the East pit, two starter pits and 1 subsequent cutback were developed, using some of the smaller Whittle pit shells associated with a high gold price, as a guide. Table 20-22 presents a summary of the material breakdown as


contained within the East pit, excluding Inferred mineral resources. The West pit was engineered following a single stage strategy, that is, without intermediate cuts. Table 20-22: East Pit Material Inventories

Area Total

Mt Waste

Mt Ore Mt

Gold g/t

Silver g/t

Gold Moz

Silver Moz

SR t:t

Starter East 35.0 26.6 8.4 2.3 8.1 0.6 2.2 3.2

Starter Central 6.0 4.4 1.6 1.8 21.8 0.1 1.1 2.8

Cutback 54.1 41.2 12.8 2.0 11.6 0.8 4.8 3.2

Total 95.1 72.2 22.8 2.1 11.0 1.5 8.1 3.2

Figure 20-12 illustrates the starter pits, final cutbacks available to the scheduling process.

20.1.7. Updated RSG Engineered Pits

EGL re-classified the material in the RSG Pit 16 block model at US$650/tr.oz. gold US$7.5/tr.oz.silver which reassigns 7.8 Mt of material, previously deemed waste at US$425/tr.oz gold and US$7.0/tr.oz. silver, to ore at 0.7 g/t gold and 14.0 g/t silver. This material has been deemed mineable at US$650/tr.oz gold US$7.5/tr.oz.silver and is fed to the plant during the last 4 years of the schedule. The waste tonnage mined decreases correspondingly. A further 6.3 Mt at 0.5 g/t gold and 8.9 g/t silver is deemed mineable by EGL at US$650/tr.oz gold US$7.5/tr.oz.silver in existing dumps. Resources are detailed in the previous section and a recovery factor of 0.9 has been applied to this. This is more conservative than the 0.95 factor used for in-pit reserves since the dumps were placed on irregular topographic surfaces This material is cheaply mined as it excludes blasting and an uphill haul from a pit. This material has been fed to the plant during the last 4 years of the schedule.

Table 20-23: Updated Engineered pit material inventories

Total

Mt Waste

Mt Ore Mt

Gold g/t

Silver g/t

Gold Moz

Silver Moz

SR t:t

RSG Total 133.1 100.3 32.8 2.0 11.4 2.1 12.0 3.1 Reassigned

Material -7.8 +7.8 0.7 14.0 0.2 3.5 -0.8

Pit Adjusted

133.1 92.5 40.6 1.8 11.9 2.3 15.5 2.3

+ Dump Material

6.3 6.3 0.5 8.9 0.1 1.8

Total 139.4 47.0 1.6 11.5 2.4 17.4

20.1.8. Mine Production Schedule

The mine production schedule is based on the final engineered pits and starter pits designed by RSG. ISM generated the mine production schedule in Microsoft Excel, based on a bench by bench mining approach, within each cut / pit. The production scheduling


process consists of developing a mine plan that is economically feasible and, at the same time, practical and robust. In order to minimise the waste pre-strip and maximise the ore exposed, several criteria needed to be satisfied. These include:

• minimum width of cutbacks, including access for mining for every bench; • mining fleet flexibility for campaign mining; • average ore haulage travel distance of 2 km, or less, over the Life of Mine; • identification of high grade areas; and, • minimum number of working faces per period of time.

ISM considers that applying a truck and shovel mining method for the operation satisfies these requirements. The mine production schedule is based on the need to supply an eventual product feed of 3.0 Mtpa, while maximising the annual cash flow. This produced a compliant schedule as follows:

• almost 14 years of pit life;

• 47.0 Mt of ore feed at 1.6 g/t for gold and 11.5 g/t for silver; and

• 92.5 Mt of waste.

Annual mine drawings detailing the bench configurations, and sections presenting the pit profile against surface topography, are given as Figure 20-13 to Figure 20-26. The mining requirements for equipment, manpower, facilities and infrastructure have been developed from the mine production schedule presented in Table 20-24. This Table continues to report the 7.8 Mt of material now reassigned from waste to ore as waste. This ensures that a mining cost is applied to the material. Subsequently at the end of the schedule a rehandle cost is applied to this material to transfer it from the stockpiles to the crusher. Table 20-24: Mine Production Schedule

Area Item Unit Total P0 P1 P2 P3 P4 P5 P6 P7 P8 P9 P10 P11 P12

Total Rock Mt 133.1 9.0 14.0 14.0 14.0 14.8 15.0 7.9 7.7 10.0 10.0 10.0 6.1 0.7

East Total Mt 95.1 9.0 14.0 14.0 14.0 14.8 15.0 7.9 4.5 1.3 0.5

Waste Mt 72.2 8.2 11.9 10.0 11.4 12.5 11.3 4.2 2.0 0.5 0.2

Ore Mt 22.8 0.8 2.1 4.0 2.6 2.4 3.7 3.7 2.5 0.8 0.3

Au g/t 2.1 2.1 2.1 2.3 2.1 1.7 1.9 2.2 2.3 2.1 2.0

Ag g/t 11.0 6.6 6.5 9.2 14.7 20.4 12.5 8.0 9.7 8.6 7.1

West Total Mt 38.1 3.2 8.7 9.5 10.0 6.1 0.7

Waste Mt 28.1 3.0 6.4 6.6 7.6 4.2 0.3

Ore Mt 10.0 0.2 2.3 2.9 2.4 1.9 0.3

Au g/t 1.8 1.5 1.8 1.8 1.7 2.0 2.4

Ag g/t 12.4 12.1 11.1 12.5 12.4 14.7 8.5

Stock In Mt 3.4 0.8 1.0 0.7 0.7 0.1 0.2

Pile Au g/t 1.5 2.1 1.7 0.9 1.3 0.9 0.9

Ag g/t 9.0 6.6 9.2 12.2 8.3 7.9 10.4

Out Mt 3.4 0.7 0.4 0.6 0.3 0.6 0.9

Au g/t 1.5 2.1 1.7 1.7 1.1 1.1 1.1

Ag g/t 9.0 6.6 9.0 9.0 10.1 10.0 10.0

Bal. Mt 0.8 0.1 1.0 0.6 0.7 1.4 1.1 1.2 1.5 0.9


Au g/t 2.1 2.1 1.7 1.7 0.9 1.1 1.1 1.1 1.1 1.1

Ag g/t 6.6 6.6 9.0 9.0 12.2 10.1 10.1 9.9 10.0 10.0

Mill Ore Mt 32.8 2.8 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 2.7 0.3

Au g/t 2.0 2.1 2.5 2.1 1.7 2.1 2.4 2.1 1.9 1.9 1.6 1.7 2.4

Ag g/t 11.4 6.5 9.3 13.9 18.1 12.5 7.9 9.9 10.6 12.0 11.9 13.2 8.5

EGL will be responsible for site preparation, haul road construction and maintenance, excavation and haulage of ore to the crusher, and waste to the dumps, oversize breakage, and equipment maintenance. ISM envisaged that conventional open pit mining techniques and equipment is utilised, consisting of hydraulic excavators (both bull clam and backhoe configuration) loading into off-highway rear dump trucks. Additionally, ISM has assumed pit benches can be developed, and will conform to the specific excavating equipment, but generally to 5 m vertical lifts.

Access roads up onto the site are developed during the pre-production phase, camp construction and mobilisation. The main arterial roads are generally built to a 25 m width. There are limited soil types, and/or soil making materials available on the site. Where these can be identified, they will be recovered during the pit preparation phase and stockpiled for future use with progressive dump and pit rehabilitation. Low grade ore is stockpiled during the initial production years. This ore helps balance the feed to the plant during periods of high waste stripping. High grade Run of Mine (RoM) ore is transported to the RoM area, and discharged directly into the crusher. A minimal RoM stockpile will be available for surcharge materials at the RoM pad area. Most of the waste rock from the excavation areas will be hauled to external dumping areas, initially on the north flank of the pit, and later on the south flank of the pit. Where possible, end tipping is utilised and the dump profile is progressively extended outwards. Generally, construction criteria is followed to ensure a good geotechnical design, which includes deposition over a large tipping front, and judicious placement of blocky materials into the base areas of the dumps. The practicalities of backfilling areas of the excavated East pit to minimise both, transport costs and visual impact on the local environment, were investigated. Consequently, some waste rock is end-tipped in the East pit in the latter years of mining. When dumps are redundant, the area will then be formed into a gentle sloping formation and covered, where possible, with soils or soil forming materials in accordance with environmental regulations. ISM has assumed that drilling and blasting is required to assist fragmentation and subsequent loading of all rock, waste and ore-bearing. The on-site staff at Deva will undertake mine planning, mine scheduling, grade control and performance monitoring.


20.1.9. Haul Distance Profile and Truck Fleet Dimensioning

Annual average distance haul and dump profiles (courses) were estimated by ISM for ore, waste and re-handle rock types, utilising the bench block plans and the mining production schedule. Each course consists of return average distances, at various gradients, from the loading location to the crusher station, dump sites, or stockpiles (


Figure 20-27). These courses are the basis for the truck fleet size and diesel consumption analysis. The Fleet Production and Cost software (FPC), developed by Caterpillar, was used to estimate truck cycle times and diesel fuel consumption rates for each return course. Although the average haul and dump distances are adequate to estimate equipment consumables and cycle times for a given period, a different approach must be taken to estimate the necessary number of trucks. Therefore, the longest one way course in a given year was used for this purpose, in order to attempt to maximise the utilisation of the loading equipment.

20.1.9.1. Assumptions

Based on the mine production schedule profile, ISM considered different combinations of excavator-truck capacities. After considering the main equipment operating costs, excavators’ loading cycles, and designed pit geometric constraints, ISM selected an excavator with bucket capacity of some 8.5 m3, loading 62t off-highway truck, for all load, haul, and dump activities. Therefore, the operating specifications and performance of the Caterpillar 775E (CAT 775) off-highway truck, with nominal capacity of 62 t, were used to simulate all haul and dump activities. Truck cycle times and diesel fuel consumption rates are based on the specific performance charts of this model, as illustrated in Figure 20-28. Nonetheless, for the same course, similar capacity trucks by other providers should be within ±10% of the CAT 775 estimates. The rimpull and retarding charts assisted with the estimation of the average travel speed of a truck. A maximum travel speed limit of 30 kph was set for each course. The analysis also considered whether a truck was travelling loaded, or not. The truck operator’s efficiency, as defined in FPC, ‘is based on the one way haul distance for each course. The longer the haul distances, the higher the operator’s efficiency rating. This excludes any wait time. Table 20-25 presents the truck operator’s efficiency rating as a function of the haul distance. Table 20-25: Truck operator’s efficiency profile

Haul Distance Operator's Efficiency m %

152 77 305 80 610 86

1,067 90 1,524 92 2,438 95

> 2,438 95

20.1.9.2. Haul and Dump Courses

The Certej open pit operation was divided into 2 distinctive areas for the estimation of average one-way courses per rock type, namely East and West. Each course was defined in terms of true distance, haul road gradient and tonnes for ore and waste rock types. The same average one-way course is assumed for the return to the excavator.


Haul truck productivities were calculated on an annual basis from first principles, and were guided considering the following:

• generalised haul profiles were determined for each unique rock type and general destination; that is, waste to exterior dump, high grade ore to RoM pad and/or stockpile;

• speed limits were determined for the various phases of haul, and average speeds established for empty and full runs for in-pit, ramp, main and restricted haul runs;

• the travel time was simulated for each haul profile. Adding the fixed and travel times determined the total cycle time. The average trips per shift and shift productivities were compiled for each time period.

20.1.9.3. Excavator and Front-End-Loader Parameters

Rock loading activities are carried out by two types of loading unit, 2 x 8.5m3 excavators, in backhoe configuration, which are primarily used for in pit mining, and an 8.4 m3 Front-End wheeled Loader (FEL) used for crusher feed and loading trucks for stockpile re-handle. FEL productivities are based upon their primary role of crusher feed and then extended to their secondary role of truck loading from stockpile. Table 20-26 presents the main loading units parameters. Table 20-26: Loading unit parameters

EXCAVATOR FRONT-END-LOADER

Item Unit Value Item Unit Value

Bucket Capacity lcm 8.5 Bucket Capacity lcm 8.4

Bucket Fill Factor % 96.0 Bucket Fill Factor % 97.2

Availability % 90-75 Availability % 90-75

Loader Cycle Time min 0.44 Loader Cycle Time min 0.65

First Bucket Dump min 0.05 First Bucket Dump min 0.10

Truck Exchange Time min 0.7 Truck Exchange Time min 0.7

Total Loading Time min 2.07 Total Loading Time min 2.75

Total Dump Time min 1.2 Total Dump Time min 1.2

Loader Passes per Truck # 4 Loader Passes per Truck # 4

Truck Effective Capacity dry t 59.7 Truck Effective Capacity dry t 59.8

The overall availability of both, excavator and FEL, starts at 90%, and is reduced by 2.5% every year until it reaches 75% in Year 6. Then, this availability of 75% is kept constant for the remaining of the life of mine. This is done to simulate the decreasing productivity of the loading units over time, as the fleet ages.

20.1.9.4. Results

Table 20-27 presents the results of the truck fleet size and diesel consumption estimation analysis. In summary, the estimated annual number of trucks required was based on the longest average course per material type per mining face, thus allowing the loading units to work at full capacity, when required. Hourly truck diesel consumption rate for the LoM was estimated at some 79 litres.


Truck productivity for ore and waste material types haulage, for each mining area, was based on the estimated cycle times and effective work hours per year. The average truck production rate for the LoM is some 270 t, for average cycle times of 14 minutes. The LoM average truck operator’s efficiency is 92%, for an average, one-way, course distance of 1.6 km. No such detailed estimate of unit requirements has been developed for the transport of the stockpiled low-grade material and dump material to the process plant. Hauls will be flat, waste volumes reduced to zero and re-handle costs for these materials will be far less than those estimated for the pit haulage Table 20-27: Truck fleet analysis summary

Item Unit P0 P1 P2 P3 P4 P5 P6 P7 P8 P9 P10 P11

Material Mt 9.0 14.0 14.0 14.0 14.8 15.0 7.9 7.7 10.0 10.0 10.0 6.1

Distance km 1.9 0.8 1.6 1.0 1.2 1.6 2.3 2.3 1.9 1.3 1.6 2.3

Cycle time min 13.6 8.2 13.3 9.7 10.8 13.6 18.2 18.5 15.8 12.7 15.1 20.2

Production t/hr 272 449 277 381 341 271 203 199 234 291 245 182

No. required # 10 10 10 10 10 12 12 12 12 12 12 12

Operator eff. % 95 86 92 88 91 92 95 95 95 90 92 95

Avg. Speed kph 17 12 14 12 13 14 15 15 14 12 13 14

Fuel rate lt/hr 72 61 78 70 80 85 88 86 85 82 85 88

20.1.10. Mining Equipment

Mining equipment was selected based on operating conditions and experience from working in similar conditions. Mine equipment requirements were calculated based upon the annual mine production schedule, the mine work schedule and equipment shift production estimates, formulated from first principles in Microsoft Excel, and adjusted as necessary for practicality. A summary of the total mine fleet is presented in Table 20-28 and illustrates the number of maximum units required on the property for the mine life. Table 20-28: Mining equipment

Mining and Support Units Max No. of units

Rotary percussion drill rigs (120 -200mm diam. Holes) 3

8.5 m3 hydraulic excavator 2

62t Off-highway rear dump trucks 12

500 kW wheel loader 1

300 kW track dozers 3 300 kW wheel dozers 1

150 kW motor grader 1

120 kW integrated tool carrier 1

Water truck 1

Fuel and lube truck 1

The mining fleet has been formed with equipment from world renowned manufacturers. Quotes have been specifically requested for the Certej project. There is sufficient equipment to perform the all of the duties envisaged to operate the engineered pit as enumerated in the following points:


• construct additional roads, including upgrading of existing haul roads, as needed to support the mining activity;

• mine and transport the ore to the RoM crusher or satellite stockpile; mine and transport waste rock from the excavation areas to the waste storage areas, and includes dozing and dump maintenance;

• re-handle of ore rock from the stockpile areas to the crusher and from the identified historical ore dumps;

• maintain the entire mine working areas, in-pit haul roads, waste storage areas, stockpiles, satellite stockpiles and external haul roads; and

• build and maintain in-pit and on-dump drainage structures.

20.1.10.1. Mine Work Schedule

The mine is scheduled to work 365 days per year, and will operate three 8-hour shifts per day, for a total of 1,095 available shifts each year, for the mine life. ISM has utilised the 3 x 8 hour system, taking into account the cost of labour, the haulage distances and the increased risk to the safety of operations in working to a full 12-hour shift. A 4-crew roster system is required to maintain this schedule.

20.1.10.2. Equipment Shift Production Estimates

Mechanical availability refers to the equipment’s mechanical availability, which represents the amount of time averaged over the life of the equipment that the unit is mechanically available to work. ISM has used a declining mechanical availability factor, from 90 to 75%, over the life of the unit, to simulate the variations in productivity over time, as the unit ages. Mining utilisation represents the actual time during the shift that the equipment is working “productively”. Table 20-29 presents the operating time per shift for the major mining equipment. Out of a total 8-hour shift, the equipment will accumulate 420 minutes (7.00 hours) metered time, and 347 minutes net productive operating time based on a job efficiency of 82.6% (50 minutes per hour). Table 20-29: Major equipment shift operating time Item Minutes

Scheduled Time per Shift 480

Less Scheduled Non Productive Times

Travel Time/Shift Change/Blasting 10

Equipment Inspection 10

Lunch/Breaks 30

Fuelling, Lube & Service, others 10

Net Scheduled Productive Time (Metered Operating Time) 420

Net Productive Operating Time per Shift 347

Job Efficiency (50 min Productive Time per Metered Hour) 82.6%

Productivities per unit-shift (prior to any allowance for mechanical availability and utilisation of availability) were first calculated for all of the major mining equipment, and these were then used to determine the following operating requirements for each year:


• Operating shifts = Required Production / Productivity per unit shift; • Total fleet = Operating Shifts / available shifts x MA x UA; • Fleet Utilisation = Operating Shifts / Total fleet x Available Shifts; and, • Operators = Total Fleet x Crews x Fleet Utilisation.

Where MA refers to the equipment’s mechanical availability and UA refers to the utilisation of available equipment. It is impossible to fully utilise any unit for 100% of its available time, as time is always lost for other activities, such as:

• assigning operators to units; • sickness; • industrial problems; and • blasting activities.

Auxiliary equipment, such as dozers, graders and water trucks, are an essential requirement to the mining operations, but typically are never heavily scheduled.

20.1.10.3. Major Equipment

Mine equipment requirements were calculated to assure sufficient production capacity to meet the mine production schedule. A combination of excavators in backhoe and shovel configuration have been selected as the primary production equipment to load 62 t off-highway dump trucks. The shovel configuration is the preferred option in areas of non selective mining, such as waste removal, whereas the backhoe configuration is more adept for any selective mining. Drilling and Blasting Large drills capable of drilling between 120 and 200 mm diameter holes in a 5 m pass have been selected because of low unit costs and longevity considerations. ISM has assumed bulk blasting techniques, with little selective drilling beyond the detail of the model mining units, when blasting within the mineralised zone. Excavation 8.5 m3 hydraulic excavators, in bull clam and backhoe configuration, capable of loading 62t haul trucks in four passes are considered the most appropriate loaders for this operation. Diesel powered units are the recommended option for Certej due to the high cost of power distribution of the localised electrical network. Haulage Mechanically driven 62 t haul trucks have been selected as the appropriate hauling unit for both, the hydraulic excavators and the 600-kW Front-End wheeled loader used in the stockpile areas. Road Construction and Maintenance The network of haul roads, minor roads, ramps, working areas and waste tipping areas are maintained to a high standard of road repair by the fleet of 150.kW grader and 300-kW wheel dozers. The grader will concentrate upon the main arterial haul roads, whilst


the wheel dozers will clean up the areas around the excavating units and the tipping areas on the dumps. Stockpile Management The 500-kW Front-End wheeled loader is the preferred option for the ongoing re-handle from the stockpile to the crusher feed on the RoM pad. In periods of necessity, this loader can be utilised in the excavation areas to assist the production excavators.

20.1.10.4. Auxiliary (Support) Equipment

The major auxiliary equipment refers to the mine units which are not directly responsible for production, but which are scheduled on a regular basis. The primary function of the auxiliary equipment is to support the major production units, and provide safe and clean working areas. Equipment types include:

• 300 kW track dozers; • 300 kW wheel dozers; • 150 kW motor grader; • Water truck; and, • 500 kW wheel loader.

The minor auxiliary equipment includes all the smaller items of specialised and non-specialised mining equipment required for supporting the mining process, and includes:

• maintenance vehicles required to servicing and fuel the tracked equipment in the field, such as fuel trucks, lubrication and field service trucks, mobile breakdown repair vehicles;

• utility vehicles for transportation of labourers and field operatives around the mine site, and for general supervisor tasks and management;

• specialised equipment required in the maintenance area, such as mobile cranes, flatbeds for transporting tracked vehicles to the workshops etc. will be hired locally; and,

• mobile lighting plants required to illuminate specific operations, or areas of high safety risk, during the hours of darkness.

ISM envisages that daily checks and lubrication will be done at the machines, using the designated service/lube truck. Rubber tyred units will be serviced and maintained in the workshops. Track-mounted equipment will be serviced in the field. A hired low-bed tractor-trailer will be used to bring the track-mounted equipment into the workshops for maintenance.

20.1.11. Waste Rock Management and Stockpile Strategy

20.1.11.1. Run of Mine Stockpiles

High grade ore is transported to the Run of Mine (RoM) pad and discharged directly into the crusher. A nominal allowance of 15% is considered for ore misdirection, haulage congestion and disruptions to the haulage patterns, i.e. non operative crusher time. Two RoM stockpile pads were designed, with a combined approximate capacity of 1.5 Mt. RoM ore from the East pit is kept separated from RoM ore from the West pit. This is a necessity given the different rock characterisation of the two areas and permits a controlled blend to the mill if desired.


20.1.11.2. Historic Ore Dumps

The excavators will free dig this material in the last 2.5 years of the mining schedule and it will be transported the RoM pad and discharged directly into the crusher.

20.1.11.3. Waste Rock Disposal

The haul road system from the mining area to the Administration areas and Run of Mine pad is developed prior to full scale mining operations commence. Some of the initial rock excavation will be utilised as fill for the road development programme. Generally, construction criteria must be followed to ensure a good geotechnical design. Dumps were designed by ISM according to the following parameters and also taking cognisance of the performance existing dumps:

• Face slope 30º; • Bench height 30m; • Berm width 15-25m; and, • Overall slope ~20º.

Dump capacities were based on a swell factor of 30%. Some 18.7 Mt of waste rock, excavated from the West pit, will be used to partially backfill the East pit. Dump locations were determined by taking into account geologically prospective ground, the existing terrain drainage patterns, waste haulage profiles, and the space and infrastructure issues required for the planned operation. Therefore, two locations were identified for designing the dumps, namely North and South, given their location with respect to the pit. Positioning of the foothold of the dumps are designed to minimise the blockage or possible contamination from ARD of natural watercourses. Sterilisation drilling is planned at an appropriate time.

The North dump is located in a valley, and the height difference between the bottom and the top of the dump is some 75 m. The North dump covers an area of some 19 Ha and has a capacity of 0.4 Mm3, with no capacity to extend. The South dump is located in a valley, and the height difference between the bottom and the top of the dump is some 200 m. The South dump covers an area of some 60 Ha and has a capacity of 40.7 Mm3, with ample capacity to extend Some 9.7 Mt of waste rock excavated during the pre-production period will be used to build the RoM and stockpile pads, as well as other infrastructure items, as required. The proposed waste schedule is presented in Waste rock from the pre-production period is primarily used for providing fill to build the foundation of some of the surface infrastructure, for building the stock pile pads, and for building the main pit-to-crusher haul roads. The North dump is built from year 1. In year 2 the building of the South dump commences. Backfilling of the East pit commences in year 9, only after mining activities have completely ceased in it. Generally, based on the location of the dump sites, and the manner in which mining of the pits progresses, only one dump area at a


time is used. Consideration is being given to designing out the North dump due to its very small size. Table 20-30 continues to report the 7.8 Mt of material reassigned from waste to ore as lower grade material. This ensures that a mining cost is applied to the material. Subsequently at the end of the schedule a re-handle cost is applied to this material to transfer it from the stockpiles to the crusher.


Figure 20-27 illustrates the final pit design, dumps and associated roads for Certej. Table 20-30: Mine waste schedule

Area Unit Total P0 P1 P2 P3 P4 P5 P6 P7 P8 P9 P10 P11 P12

Development Mt 9.7 8.2 1.0 0.5

North Dump Mt 34.6 10.9 10.0 11.4 2.2

South Dump Mt 37.3 9.8 11.3 4.2 4.9 6.9 0.2

In-Pit (East) Mt 18.7 6.6 7.6 4.2 0.3

Total Mt 100.3 8.2 11.9 10.0 11.4 12.5 11.3 4.2 4.9 6.9 6.8 7.6 4.2 0.3

20.1.12. Manpower Requirements

20.1.12.1. General

Mine personnel includes all the salaried supervisory and staff working in the mine operations, maintenance, and engineering departments, and the operational labour required to operate and maintain the drilling, blasting, loading, hauling, and mine support activities. Positions requiring specific skills, or experience, not available in Romania will initially be filled by expatriates. In addition to performing their regular job functions, expatriate personnel will transfer knowledge and expertise to develop the capabilities of the local staff. Mine organisation was setup in a typical hierarchy at normal full production scenario as illustrated in Figure 20-29. Salaries were estimated for Romanian wages, except for the first three years of production, where the mine manager is an expatriate.

20.1.12.2. Salaried Staff

This includes the manpower for the mine management team, mine operations, mine maintenance, mine engineering, mine geology, mine administration and accounts.

20.1.12.3. Mine Operating Labour

Mine operating labour is classified into two main departments, Mine Operations and Maintenance. The majority of personnel in mine operations are equipment operators, and these are calculated from the unit equipment requirement tables on an annual basis. ISM assumed that there will be a fairly high level of cross training of the various operator types. Additional mine personnel are assigned to perform the following tasks:

• Service crew, who are responsible for the operation of the water trucks, and road construction crews;

• Blasting crew, who are responsible for loading and stemming the blast holes and initiation; and,

• Labourers, who are generally unskilled workers that assist with many of the mine support facilities such as moving pumps and pipes, and general mine clear up.

•


The complement for the mine maintenance crews has been calculated taking cognisance of the fleet numbers for both, the major and support equipment. General maintenance personnel (fuel and lubrication operatives, tyre personnel, and general labourers) have been estimated by ISM. An additional allowance in the manpower is required to cover vacations, training, sick leave and absenteeism. ISM has covered this contingency with an overall 15% increase in compliment.

20.1.13. Mining Programme

20.1.13.1. Pit Start-Up Strategy and Pre-Strip

Based on the optimisation work, 4 pits have been designed, namely Starter East, Starter Central, East and West. The Starter pits mine the least costly part of the deposit in the early production periods. The East pit is a single cut back of the Starter pits. This results in a reduction in waste stripping and better cash flow for these pits. The West pit is mined to its extents without any intermediate cuts. A pre-strip period of almost 9 months is envisaged during which 8.2 Mt of waste is removed and 764 kt of ore are stock piled. 2.8 Mt of ore are processed in Year 1 during the ramp-up period, later achieving the designed 3.0 Mtpa throughput in Year 2, until the end of the Life of Mine. The pit bottom of the Starter East pit is reached in Year 3, achieving a pit depth of some 140 m. Similarly, the pit bottom of the Starter Central pit is reached in Year 4, to a pit depth of some 60 m. Waste stripping of the East pit commences in Year 2, resulting in the exposure of ore by the beginning of Year 3. Final pit depth is achieved in Year 9 at some 220 m. Waste stripping of the West pit commences in Year 7, exposing ore in the same year, and reaching a final pit depth of some 200 m in Year 12. All pits are mined on a bench by bench basis. Figure 20-15 through to Figure 20-26 illustrate the end of period face positions.

20.1.13.2. Long Term Planning and Life-of-Mine

The mine schedule has been developed utilising hydraulic 8.5 m3 excavators and 62 t off-highway haul trucks. The LoM is almost 12 years at an overall strip ratio of 3.1 tonnes of waste per tonne of ore. 1.7 Moz of gold and 9.0 Moz of silver are recovered at average process metallurgical recoveries of 81% for gold, 75% for silver. A total of 32.8 Mt of ore are crushed and processed at average feed input gold grades of 2.0 g/t and 11.4 g/t, for gold and silver, respectively. The Long Term Plan envisages a maximum stock pile capacity of 1.5 Mt of ore in Year 9. The maximum mining rate is some 15.0 Mt in Year 5. Nonetheless, average LoM mining rate is some 11.5 Mtpa, starting at 14.0 Mtpa in Year 1, and then decreasing to some 10 Mtpa after Year 5. Then, it drops to 6.0 Mt in Year 11. The schedule of equipment is capable of these production scenarios.


20.1.14. Mining Reserves

All stated probable reserves are completely included within the quoted Mineral Resources. Further exploration drilling will take place prior to production which is expected to convert some of the probable reserves to proven reserves. Table 20-31 presents a summary of the Certej Mineral Reserves. Table 20-31: Certej mineral reserves

‘000t

Au Au Ag Ag

g/t Moz g/t Moz


Probable 32,811 2.0 2.12 11.4 12.0


Probable 7,829 0.7 0.18 14.0 3.5

Certej Existing Dumps

Probable 6,320 0.5 0.11 8.9 1.8

Total 46,960

2.41

17.3

This reserve was estimated and reported in accordance with the Instrument and the classifications approved by the CIM Council in November 2004. The reported reserves were compiled by RSG, which has sufficient experience, relevant to the style of mineralisation and type of deposit under consideration.

20.1.15. Conclusions on Mining

The Certej open pit design, equipping and costing follows traditional design procedures. The proposed technology and designs have been used for decades in operations in more rigorous conditions. The mining of the Certej deposit is considered to be a low risk means of extracting the identified mineral reserves. The engineered pit designs were based on Whittle optimisation to determine the general pit shape and depth. This comprised a main pit over the east, intermediate and central zones of the orebody and an adjacent smaller pit over the west zone. The final pit depth is some 220 m in the East pit, and some 200 m in the West pit. Optimised pit shells generated in Whittle were used as guidance to detail engineer the pits, including exact ramp positions and bench configuration. Average overall pit slopes including ramps are generally 40° to 60° according to geotechnical conditions but flatten to 30° in some places at the pit rim. Total ore probable reserves from the open pit is 32.8 Mt at 2.0 g/t for gold and 11.4 g/t for silver, at an approximate gold cut-off grade of 0.8 g/t Au, and 92.2 Mt of waste, for a LoM strip ratio of 3.1:1 (t:t). Some 9.7 Mt of waste are used for building roads and pads, 18.7 Mt are used for in-pit dumping in the East pit, and some 63.8 Mt are stored


between the North and South dumps. An additional 7.8 Mt of in-pit reserves at 0.7g/t gold and 14 g/t silver becomes economic to process at $650 per oz Au and $7.5 per ounce Ag. This lower grade reserve is classified as probable and will be processed at the end of mine life. Existing dumps from historic production contain probable reserves of 6.3Mt at 0.53 g/t Au and 8.9 g/t Ag. This material will be processed at the end of mine life with the lower grade in-pit material. Total Reserves of some 47Mt will support a mine life of 16 years. An owner operated open pit mining fleet has been selected to suit the required mining schedule, with the key equipment being 8.5 m3 hydraulic excavators and 62 t trucks, capable of excavating and moving up to 15 Mt of material per year. In peak years, the excess capacity of the re-handle loader could be used to supplement the excavators. A mining schedule has been prepared for 3 Mtpa of run of mine ore through the mill. Staffing is 231 people with a mining operating cost of € 1.22 per tonne of material mined derived from first principles. EGL consider the schedule and costs to be within +15% and -10% as is normal for a Feasibility Study.


Figure 20-1: Stereonet of rock mass structure, existing open pit


Figure 20-2: Rosette of rock mass structure, existing open pit


Figure 20-3: Stereonet of rock mass structure, geotechnical and re-logged boreholes


Figure 20-4: Rosette of rock mass structure, existing open pit – joints >3m


Figure 20-5: East pit section views


Figure 20-6: West pit section views


Figure 20-7: East and west pit long section views


Figure 20-8: Plan of Open Pit Showing Sectors, Geotechnical and Re-Logged Boreholes


Figure 20-9: Pit size vs. commodity price


Figure 20-10: Pit optimization results


Figure 20-11: Ore to waste relationship


Figure 20-12: Scheduling cutbacks

4

3

2

1

1. Starter East 2. Starter Central 3. East Pit 4. West Pit


Figure 20-13: Topography


Figure 20-14: Pre-Production


Figure 20-15: Pit year 1


Figure 20-27: General mine layout


Figure 20-28: Rimpull and retarding charts for CAT 775E (C282)

Rimpull (kilos)

Speed (kph)

Speed (kph)

Retarding (kilos)

Gear 1 Gear 2 Gear 3 Gear 4 Gear 5 Gear 6 Gear 7

Gear 1 Gear 2 Gear 3 Gear 4 Gear 5 Gear 6 Gear 7

5 10 15 20 25 30 35 40 45 50 55 60 65 70

0 5 10 15 20 25 30 35 40 45 50 55 60 65

50,000 40,000 30,000 20,000 10,000 0

20,000 10,000 0


Figure 20-29: Mine Organisation and Workforce

r MINE

MANAGEMENT r

*Mine Manager **1 **denotes expatriate

*Secretary 1 * denotes staff

MINE OPERATION r

r

MINE MAINTENANCE

r

MINE PLANNING AND GEOLOGY

r

*Mine Superintendent 1 *Maintenance Superintendent 1 *Chief Engineer 1 Shift Foreman 5 *Maintenance Engineers 2 *Planning Engineer Short Term 1 Drivers 5 Maintenance Foreman 5 *Planning Engineer Long Term 1 Labourers 4 Senior Technician 10 *Draughtsman 1 Explosive Loaders 4 Technician 10 *Chief Surveyor 1 Drillers 10 Assistant 14 *Assistant Surveyor 2 Driller Assistants 10 Welder 5 *Driver / Helper 1 Excavator Operators 10 Tyres 5 *Data Processing Engineer 1 Truck Operators 56 Refueling 5 *Chief Geologist 1 Bulldozer / Wheeldozer Operator 19 Stores 5 *Mine Geologists 2

Grader Operator 5 Clerk 3 *Computer Geologist 1

Wheel Loader Operators 5 TOTAL 65 *Samplers/helpers/Lab 10

Watering truck Operators 5 *Draftsman 1

TOTAL 139

*Clerk 1 TOTAL 25

TOTAL Personnel - 231


20.2. Tailings Management Facility


The mine is expected to produce approximately 3.4 million tonnes of tailings per year for 15 years over the Life of Mine.

The processing of the Certej ore deposit will be undertaken in two stages:

• ROM Ore flotation, producing gold concentrate and flotation tailings.

• Oxygen leaching (Albion process) followed by CIL leaching of the oxidized concentrate and gold and silver recovery on the eluted carbon by electrowinning. The products resulted in this second stage of ore processing are doré alloy and cyanidation tailings (CIL Tailings).

Two types of tailings will be produced by the process plant, flotation tailings and CIL tailings and these will be stored in two separate Tailings Management Facilities (TMF). The flotation tailings will be inert apart from low levels of residual sulphides not recovered during the flotation process. The CIL tailings will be the slurry treated in the Inco detoxification circuit to lower the levels of residual cyanide and under normal operations this TMF will be zero discharge. The Flotation TMF will be required to house 40.5 million tonnes of tailings and the CIL TMF designed to store 10.5 million tonnes of tailings.

The selected TMF sites are within the Macrisului Valley system which lies some 2 km north east of the process plant site (Figure 20-30).

On selection of the site, Professors D. Stamatui and M. Selarescu of the Technical University of Construction, Bucharest, were engaged by EGL to advise and assist the Romanian engineering consultants Cepromin on the design of the TMF.

Golder Associates (UK) Ltd (Golder) were engaged to undertake the review of the Cepromin design, to suggest modifications where necessary and complete the basic engineering and costing estimates for both the Flotation and CIL TMF respectively, and include the aspects of disposal discussed in the subsequent sections.

20.2.2. Initial Design

The initial design was undertaken by the Romanian Consultancy Cepromin, which developed a design proposed by two Professors from the National Technical University of Bucharest (NTUB), Professors Stamatiu and Selarescu. Golder was commissioned by DG to retain the overall concept of the design undertaken by Cepromin and the NTUB Professors, and develop it to a Bankable Feasibility level and which is presented in this section.


20.2.3. Data Review

The design process by Golder comprised the review of the following available data for the site: climatology, geotechnical, geochemistry, hydrology, hydrogeology, and seismicity. The Climatology data include air temperature, relative air humidity, total cloud coverage, precipitation and wind recorded between 1983 and 2006 at the Deva meteorological station. Precipitation data recorded from 1983 to 1996 at the Certej de Sus meteorological observation post and lake evaporation data have been recorded at the nearby Cinis Lake. The data were used in the water balance calculations and in sizing the storm water channels. A geotechnical assessment of the proposed TMF area was undertaken by Cepromin with the results summarised in the TMF design report. The results were used in determining the properties and depth of the bedrock and overlying material. A geochemical assessment for the Certej waste rock and tailings has been undertaken by Cepromin and included in their Waste Management Plan. The seismicity data for the site were obtained from the British Geological Survey (BGS). The Maximum Design Earthquake (MDE) is taken as equal to the Maximum Credible Earthquake (MCE) with a maximum bedrock acceleration of 0.14g, and an earthquake magnitude of 8.0. The Operating Basis Earthquake (OBE) is taken as the 1 in 475 year return period event corresponding to a maximum bedrock acceleration of 0.082g for a magnitude 8.0 earthquake. The OBE was used in assessing the stability of the embankments.

In December 2008 Golders carried out hydrogeological packer tests on available boreholes to determine water permeability values. They concluded that the rock is very strong and fresh, with little fracturing and has the required characteristics for foundation and construction of the TMF embankments. The permeability is low and the calculated seepage will have a low flow rate and a grout curtain beneath the dam will not be required.

20.2.4. Site Layout

The proposed location for the TMF is located approximately 2km north east of the Process Plant and the Open Pit, Figure 20-30. The location of both the Flotation TMF and the CIL TMF is on a tributary of Valley Certej, named Valley Macrisului Creek. The orientation of the valley is approximately north-south. The flotation TMF is to be built downstream of the CIL TMF.,

20.2.5. Design Criterion

A total of 40.5 million tonnes and 10.5 million tonnes of Flotation and CIL tailings respectively will be produced during the life of the mine. The production rate will be respectively 2.7Mtpa and 700,000tpa for the Flotation and CIL TMF. Assuming a final


placed dry density of 1.5 tonnes per cubic metre for the flotation tailings and 1.3 tonnes per cubic meter for the CIL tailings, the Flotation TMF and the CIL TMF have a required capacity of 27.0 and 8.1 million cubic metres respectively. The Flotation TMF includes the construction of one cross-valley embankment and one saddle embankment which will eventually join with the main embankment. The CIL TMF includes the construction of a cross-valley embankment. The design of the tailings management facilities at Certej incorporates diversion channels and drainage galleries for water management. These structures are briefly discussed in the sections below.

20.2.6. Dam Embankment Walls

The dam and any subsequent raises will be designed and operated to meet the most stringent criteria, such as the Maximum Credible Earthquake (MCE) and Probable Maximum Flood (PMF). Both these occurrences are equivalent to the 1 in 10,000 year return events. The design would satisfy an ICOLD hazard rating classification of high. The two dam main embankments have been designed to the highest international standards. The dam embankment designs for all the structures are similar and only vary in terms of dimensions. The main design features are presented in Figure 20-31 (General Layout), Figure 20-32 and Figure 20-33 (Flotation TMF Cross Sections) and Figure 20-34 (CIL TMF Cross Section) and are listed below.

• The Flotation TMF Main Embankment will be constructed with a starter dam 70m high. The starter dam slopes are 1.8H:1V on both the downstream and upstream slopes. The embankment will be subsequently raised to its final elevation of 160m high by three successive 15m lifts then seven 5m high lifts. The final elevation for the Flotation TMF will be 707mASL.

• The total volume of the flotation embankment fill is 7,063,000m3 and the volume of tailings retained is approximately 27.5 million m3.

• A Saddle Embankment will be built in a small depression immediately adjacent to the main embankment valley, and will ultimately be joined with the Main Embankment.

• The Saddle Embankment will be built with a starter dam 10m high up to elevation 680mASL and will subsequently be raised using the centreline method with five successive 5m lifts to its final height of 35m, for a final elevation of 707mASL. The starter Dam and the subsequent raises will all be built with a 8m thick clay core to prevent seepage in the secondary valley. A double filter will also be constructed on the downstream side of the clay core to prevent piping of the clay material into the rockfill forming the downstream slope of the embankment.

• The CIL TMF Main Embankment will be constructed with a starter dam (45m high at an elevation of 780mASL) and will subsequently be raised to its final elevation (95m high at an elevation of 825mASL) by six successive 5m high lifts using the


centreline method of construction and then six successive lifts of 2.5m using the upstream method of construction.

• The total volume of the CIL embankment fill is 1,139,999m3 and the volume of tailings retained is approximately 8,666,000m3. Benches will be incorporated in the CIL TMF Main Embankment . The benches will be 20m high, with an interbench slope angle of 1V:1.4H, and a bench width of 6m.

• All dam embankments will be constructed mainly from blasted rock originated from borrow areas/quarries to be developed within the dam impoundment area, and from excavated material originated from the construction of the diversion channels and the emergency discharge channels. Special features such as the clay cores, filters and erosion protection layer will be constructed using specified selected materials.

• A 2mm thick HDPE liner will be installed to initially to minimise seepage through the upstream face of the Flotation and CIL TMF Main Embankments during the first 18 months of deposition (time to allow the tailings beach to be formed against the embankment, to minimise seepage and damage to the filters). The HDPE liner will be anchored on the starter dam crest, valley sides and floor, and lie against the fine filter material.

• An erosion protection layer will be installed on the upstream face of both the Flotation and CIL TMF Main Embankments, on the surface not covered by the HDPE liner as well as for each raise. The layer will have a minimum thickness of 1.5m.

• For both Flotation and CIL TMF Main Embankments, a double filter (well graded sand and well graded sand and gravel) will be installed against the rockfill. This arrangement will be carried through for the full height of the starter dam and subsequently for all the raises. The filters will have each a minimum thickness of 1.5m.

• Beneath every raise for the Flotation and CIL TMF embankments, a layer of heavy duty geotextile will be placed on top of the tailings and will extend horizontally to the downstream side of the raise.

• The Flotation TMF will incorporate a Saddle Embankment which will be built in a small depression immediately adjacent to the main embankment valley.

• The Saddle Embankment will be built with a starter dam 10m high up to elevation 680mASL and will subsequently be raised using the centreline method with 5m lifts to its final height of 37m, for a final elevation of 707mASL. The starter dam and the subsequent raises will all be built with a clay core to prevent seepage in the secondary valley. Two filters will also be constructed on the downstream side of the clay core to prevent piping of the clay material into the rockfill forming the downstream slope of the embankment.

• Both TMFs are designed such that the maximum volume of pond water is reused by the process plant. Recirculation will be achieved by pumping from a barge floating in the pond area.


• Seepage will be collected in a seepage collection sump. This water will then be re-circulated via pumps back to the TMFs.

• A comprehensive diversion system consisting of drainage galleries and diversion channels will be constructed around the TMF impoundment to reduce the external catchment area to the TMF. The clean runoff water will be diverted around and underneath the TMF and discharged into the Valley Macrisului creek downstream of the Flotation TMF which will flow into the Certej creek and then to the river Mures.

• Emergency discharge channels will be constructed at the top of both the Flotation and CIL TMF on the abutments. It is anticipated that the emergency discharge channels will not be used during the life of the mine. However, the emergency discharge channels themselves are designed to accommodate the probable maximum flood at all times so that overtopping is avoided.

• The surface of the tailings will be shaped during the latter stages of the filling process to facilitate closure process of the TMF.

• A comprehensive range of monitoring equipment will be installed on all embankments in order to assess the performance of the structures during the life of the mine and post closure. A monitoring programme will be put in place and regular audits of the TMF will be undertaken.

20.2.7. Water Management

20.2.7.1. Water Balance

A water balance analysis has been undertaken to model the seasonal hydrological and the mine water parameters on a daily time step for the entire life of mine, and therefore estimate the process water supply, the containment requirement and the magnitude of discharge to the water treatment plant and/or to the environment. Eight different cases were analysed for both TMF by varying the climatic conditions. The results of the water balance are commented on below:

• The water balance for both TMF show accumulation of water during the life of the mine. If no recycling was to occur and under average rainfall conditions, the total amount of water present would be 11.8Mm3 in the Flotation TMF and 9.6Mm3 in the CIL TMF. With the normal recycling rate to the plant and under average rainfall conditions, the amount of water at the end of the Life of Mine (LOM) would be respectively 3.35Mm3 and 428,000m3 for the Flotation TMF and the CIL TMF respectively.

• There is a need to discharge/recycle water recycling from both facilities at a greater rate than the base case scenario to remove the accumulated surplus water once the pond reaches its maximum allowed volume. Under average conditions for the CIL TMF, secondary pumping to recycle water should begin in 2012, while for the Flotation TMF, pumping should begin as the mine commences operations.

• Additional Pumping requirements differ for the two facilities. For the CIL TMF, it was determined that an increased pumping capacity of approximately 20m3/hr


would be sufficient to clear the 1 in 10,000 year 24 hour storm, and a much lower rate (2 to 3m3/hr) would be adequate to effectively control average rainfall conditions. In the case of the Flotation TMF, an increased pumping capacity of approximately 100m3/hr would be required to regulate the 1 in 10,000 year 24 hour storm, whereas an increase pumping capacity of approximately of 35m3/hr should be adequate for average rainfall conditions. It is understood that these increases in pumping rates should not exceed the treatment capacity of the detox plant.

• For both TMF, a dry year would not empty the pond completely once the maximum volume has been reached. A succession of dry years during the entire Life of Mine would mean that no pumping would be possible from the CIL TMF, and that the entire make up water requirement for the plant would need to come from the river Mures. A succession of dry years would not have any consequences for the Flotation TMF, and the normal pumping rate could be applied for returning the water to the plant.

• For both CIL and Flotation TMF, a succession of wet years would mean that less abstraction of water would be required from the river Mures.

20.2.7.2. Diversion Works

The diversion scheme (diversion channels, drainage galleries and haul road drains) will be constructed in order to divert most of the storm water from the TMF sites and satisfy environmental and operational requirements, to minimise water treatment, and to ensure continuity of flow downstream.

The total catchment area increases since the area of tailings encroaches upon diversion channels and galleries.

Table 20-32: Catchment Area Summary

TMF Phase Area of Natural Land

(ha)

Area of Tailings and Ponds (ha)

Total Catchment Area (ha)

Flotation TMF 1 15.9 16.0 31.9

2 5.0 39.0 44.0

CIL TMF 1 11.9 8.6 20.5

2 3.6 26.1 29.7

Based on the TMF design life, size of the catchment area, storage capacity at the site and the topography it is considered that the diversion/scheme could be designed to accommodate the peak flow induced by a 1 in 10,000 year 24 hours duration storm event. The diversion works will comprise a drainage gallery, diversion channels, access road drains and lateral drains. The drainage gallery will be constructed on the Macrisului valley floor and the diversion channels will be constructed around the TMF and report to


the drainage galleries while the lateral drain will collect water from the side valleys and from the access road drains and divert them into the drainage gallery. The channels will be concrete lined (100mm thick) open trapezoidal and will be excavated into the fresh rock material. No stabilisation work is expected from the excavation of the channels. All the diversion works are designed to withstand the peak flow induced by a 1 in 10,000 year storm event and the intensity would therefore be 52mm/hr. Drainage Gallery

The drainage gallery will be constructed in two phases. The initial phase will extend to provide diversion for the first 3 years of the life of the facilities while the second phase will be built during operation until closure. The drainage galleries will be excavated in fresh rock and will have a circular section of 2.25m diameter. The minimum dimensions are 2m x 2m. The lining of reinforced concrete will be of thickness between 100mm, 500mm and 1000mm depending on the loading.

Table 20-33 Drainage Gallery Characteristics

Name Approx Length (m)

Diameter (m)

Average Slope (%)

Total Land Area

Diverted (m2)

Maximum Flow

(m3/sec)

Drainage Gallery

2,250 2.25 14% 1,000,000 14.55

Diversion Channels

The channels are located around the northern and western part of the CIL TMF with a total length of 1.5 km and provide diversion for approximately 300,000m2. The diversion channels will discharge the water into the drainage gallery via culverts. Lateral Drains

The lateral drains will be installed on the eastern side of the Flotation TMF inundation area. They will be formed by 630mm diameter HDPE pipes discharging into the drainage gallery. the sump downstream of the embankment into the river.

20.2.8. TMF Conclusions

Golder have reviewed the original Cepromin TMF proposals. The two main dam embankments have been designed to the highest international standards. The construction and operating costs have been estimated to an accuracy of +/- 20%. Golder recommend that further geochemical data for the tailings be completed prior to the detailed design phase with further testing to characterise the potential for acid generation of the flotation and the CIL tailings, as well as the waste rock material.


Figure 20-30: Location of the Flotation TMF and CIL TMF


Figure 20-31: Plan of TMF Embankments


Figure 20-32: Flotation TMF Main Embankment Cross Section


Figure 20-33: Flotation TMF Saddle Embankment Cross Section


Figure 20-34: CIL TMF Embankment Cross Section


20.3. Environmental Considerations

An environmental impact statement was produced in 2007 in accordance with the provisions of the Order of the Ministry of Environment and Water Administration No. 863/2002. This EIS report was produced by a consortium of Romanian agencies and institutes who prepared separate reports for the individual sections of the EIS. Their main findings are; Water ICIM of Bucharest concluded that by treating the acid rock drainage from the dumps the quality of water in the Hondol and Certej streams would meet all standards required and would improve the River Mures water quality. Water discharged from the TMFs to the local river system would be treated and comply with the admissible values laid down in NPTA 001/2005 and 002/2005. Wastes ECOIND of Bucharest defined and categorised all the wastes that would be produced during the three phases of the project, construction, operations and closure. Air AMEC (formerly AGRARO) assessed all airborne pollutions and dispersion patterns during the project life using various BAT computer models. They stated that the specific management actions would keep the air quality well within admissible levels. Biodiversity AMEC found that there were no risks to flora, fauna and animal life. Social Impact Studies on the ‘Social and Economic Environment’ and ‘Ethnic and Cultural Environment’ were produced. They concluded that the project would impact positively on the communities because of population stabilisation and the economic benefits. Archaeology Fifty two archaeological trenches were excavated by the Deva Museum and no pre-historical vestiges have been found in the mine site area. Risks and Hazards An analysis by OCON Ecorisc SRL concluded that the use of cyanide, explosives and the TMFs posed the greatest risks but this was very low due to the mitigating measures to be installed.


20.4. Financial Analysis


In order to assess the viability of the project and its sensitivity to key variables a cashflow model was constructed with the key operating and cost data that has been detailed in the previous chapters. This information, along with a set of financial parameters described below, is used to construct a Basecase financial analysis. A series of sensitivities are run on the Basecase model to show the impact of changes in costs and commodity prices.

There are some cost opportunities which could be achieved, and these are described later in this section.

20.4.2. Financial Assumptions

20.4.2.1. Metal Prices

The Au and Ag prices have been fixed throughout the life-of-mine at US$650/oz and US$7.50/oz respectively.

20.4.2.2. Exchange Rates

All operating and capital costs were calculated by AKES, Golders UK and European Goldfields in euros. The exchange rates used are detailed in their respective sections. In the financial model there is a provision to enter the US$:Euro exchange on a yearly basis. The rates currently used are US$1.30/€ in 2009, and a flat rate of US$1.30/€ for the LOM.

20.4.3. Production Schedule

The production schedule is based on the Mine Optimisation Study, (itself based on the 2007 RSG Pit Optimisation), the Hellas Gold flotation work and the HRL Continuous Pilot Plant tests.

A total of 32.8 million tonnes of ore are mined and processed through the Certej plant in years 1 to 12, and thereafter an additional 4 years of low grade ore and dumps, totalling approximately 14 million tonnes, will be processed. The financial model assumes 10.5% mass pull in years 1-11 and 4.5% (due to lower head grades) in years 12-16 of the project resulting in total concentrate of 4.1 million tonnes being produced by the mill and processed through the Albion plant.


Table 20-34: Mine Schedule for Financial Model (Years 1-12)

Totals(1-16) 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Mined Ore

East,central Tonnage '000s 22,844 764 2,101 3,951 2,603 2,361 3,671 3,737 2,516 822 318

intermediate Head Au gpt 2.09 2.13 2.06 2.30 2.15 1.70 1.88 2.16 2.29 2.11 2.01

Head Ag gpt 11.0 6.6 6.5 9.2 14.7 20.4 12.5 8.0 9.7 8.6 7.1

Western Tonnage '000s 9,967 210

2,291 2,892 2,403 1,850 321

Head Au gpt 1.82

1.47 1.76 1.78 1.71 2.03 2.41

Head Ag gpt 12.4

12.1 11.1 12.5 12.4 14.7 8.5

Low Grade Tonnage '000s 14,149 289 2,679

Head Au gpt 0.64 0.64 0.64

Head Ag gpt 11.7 11.7 11.7

Total Mill Feed Tonnage '000s 46,961 2,813 3,000 3,000 2,967 3,000 3,000 3,000

3,000 3,000 3,000 3,000 3,000

Head Au gpt 1.60 2.08 2.49 2.09 1.71 2.10 2.37 2.12

1.89 1.87 1.58 1.62 0.83

Head Ag gpt 11.51 6.5 9.3 13.9 18.1 12.5 7.9

9.9 10.6 12.0 11.9 13.1 11.4

Pit Waste Tonnage 100,317 8,236 11,899 10,049 11,397 12,470 11,329 4,167 4,948

6,886 6,790 7,597 4,218 332

Total waste & ore Tonnage 133,129 9,000 14,000 14,000 14,000 14,831 15,000 7,904 7,674 10,000 10,000 10,000 6,068 652

Contained Metal

East Au Oz. 1,553,207 188,096 240,332 201,827 162,638 202,236 228,343 204,770 105,893 19,073

Ag Oz. 8,198,440 589,015 892,706 1,344,498 1,725,242 1,209,580 765,938 954,642 593,917 122,902

Western Au Oz. 539,543 75,909 161,069 152,703 141,313 8,547

Ag Oz. 3,864,936 425,749 1,037,869 1,147,107 1,137,101 117,110

Low Grade Au 316,562 15,054 71,438

Ag 5,309,483 121,138 978,803

Total Au Oz. 2,409,311

188,096 240,332 201,827 162,638 202,236 228,343 204,770 181,803 180,143 152,703 156,368 79,986

Ag Oz.

17,372,859 589,015 892,706 1,344,498 1,725,242 1,209,580 765,938 954,642 1,019,666 1,160,771 1,147,107 1,258,239 1,095,913


Table 20-33(cont.): Mill Schedule for Financial Model (years 13-16)

2023 2024 2025 2026

Mill Feed-dumps/low grade Tonnage '000s 3,000 3,000 3,000 2,181

Head Au gpt 0.64 0.64 0.64 0.64

Head Ag gpt 11.7 11.7 11.7 11.7

Contained Metal

Low Grade Au Oz. 61,729 61,729 61,729 44,881

Ag Oz. 1,129,455 1,129,455 1,129,455 821,178

Total

Au Oz. 61,729 61,729 61,729 44,881

Ag Oz. 1,129,455 1,129,455 1,129,455 821,178

20.4.4. Operating Costs

20.4.4.1. Mine Operating Costs

As detailed in Section 6.10 the annual mining costs vary considerably as the pits deepen and the haulage distances increase. The average mining cost through LOM is €1.20/tonne mined. This excludes the cost of rehabilitation included in Section 6 as this is included as a capital item in the financial model. This includes re-handling costs, open-pit drainage and ARD pumping. We have assumed that the post reserve low grade and dumps ore will only incur a re-handling cost of €0.20/tonne.

20.4.4.2. Processing Operating Costs

The processing of the ROM ore is based on two distinct parts; the production of a flotation concentrate and the treatment of this flotation concentrate through the Albion process plant to recover gold and silver in the form of doré.

20.4.4.3. Flotation Operating Costs

The flotation operating costs were detailed in Section 7.8. The ROM ore feed is fairly constant at a nominal 3.0 Mtpa and the costs to produce a flotation concentrate do not vary significantly. They are composed of an average fixed cost of €1,918,712 pa and a variable cost of €3.57/ROM tonne. The variable costs are principally power, grinding media and reagents. The total flotation operating costs are €4.22/ ROM tonne.

20.4.4.4. Albion-CIL Operating Costs

The tonnage and grade of the flotation concentrate produced varies on a yearly basis as the feed type and head grades vary and this affects the Albion-CIL processing cost. The key fixed costs are labour (€381,363 pa), power (€386,035 pa) and oxygen plant rental (€2,400,000 pa), plus a small amount of consumables (€1,800 pa). Within variable costs of €47.97 /tonne concentrate treated, the bulk relates to consumables (limestone and


cyanide) and power. The Albion-CIL operating cost per tonne of concentrate is €58.54 for years 1-11 and €73.15 for years 12-16. The cost per tonne of ROM is €6.15 in years 1-11 and €3.25 in years 12-16.

20.4.4.5. TMF Operating Costs

The TMF operating costs totalling €0.20 /tonne concentrate are included in the process plant operating costs.


Table 20-35: Mine Operating Cost Summary

ITEM

UNIT

2010

2011

2012

2013

2014

2015

2016

2017

2018

2019

2020

2021

2022

*2023 to

2026

Total/

Unit

Distrib-utio

n €/t

Mined Drilling

k€

859

1,341

1,341

1,341

1,417

1,432

763

742

955

955

955

595

64

N/A

12,758

8%

0.09

Blasting

k€

1,472

2,357

2,448

2,382

2,486

2,573

1,586

1,504

1,848

1,853

1,814

1,238

138

N/A

23,699

15%

0.18

Loading

k€

934

1,459

1,459 1,459

1,542

1,559

854

832

1,062

1,062

1,062

637

68

N/A

13,991

9%

0.10

Hauling

k€

3,603

3,354

5,476

3,923

5,086

6,559

4,815

4,789

5,173

4,110

5,070

4,482

566

N/A

57,007

35%

0.43

Ancillary Equipment

k€

1,339

2,523

2,422

2,399

2,356

2,279

2,222

2,484

2,479

2,422

2,415

1,994

199

N/A

27,533

17%

0.21

Labour

k€

527

1,258

1,318

1,318

1,093

1,134

1,107

1,107

1,107

1,107

1,107

1,066

26

N/A

13,277

8%

0.10

General

k€

601

1,026

1,059

1,035

1,030

1,054

1,056

1,037

1,044

1,046

1,031

1,021

109

N/A

12,148

8%

0.09

TOTAL

k€

9,335

13,319

15,523

13,856

15,010

16,591

12,403

12,495

13,668

12,555

13,455

11,033

1,171

160,413

100%

1.20

€/t

Mined 1.04

0.95

1.11

0.99

1.01

1.11

1.57

1.63

1.37

1.26

1.35

1.82

1.80

1.20

€/t

Milled

12.21

6.34

3.93

5.32

6.36

4.52

3.32

4.58

4.39

3.91

5.60

5.96

3.65

0.20

4.89

Note: Year 2023 to 2026 involve the processing of lower grade stockpiled material and existing dump material and therefore only incur the re-handling cost of €0.20


Table 20-36: Concentrator Cost Summary

TOTAL EURO/annum 2011 2012 - 2013 2014 2015 - 2025 2026 TOTAL

Salaries and Wages 1,163,165 734,192 734,192 734,192 734,192 12,176,045

Consumables 6,611,927 7,044,860 6,967,163 7,044,702 5,145,110 110,305,640

Power 3,858,202 4,042,944 4,009,789 4,042,877 3,232,276 63,657,804

Maintenance Supplies 731,489 780,007 771,300 779,990 567,105 12,209,796

TOTAL 12,364,782 12,602,003 12,482,444 12,601,761 9,678,682 198,349,286

TOTAL Euro/ROM 4.39 4.20 4.21 4.20 4.44 4.22

Table 20-37: Albion Cost Summary

TOTAL EURO/annum 2011 2012 2013 2014 - 2021 2022 2023 - 2025 2026 TOTAL

Salaries and Wages 400,988 381,363 381,363 381,363 381,363 381,363 381,363 6,121,433

Consumables 11,751,532 12,529,584 12,532,156 12,513,056 6,138,765 5,073,042 3,688,882 161,964,496

Power 2,247,112 2,370,350 2,370,758 2,367,732 1,358,088 1,189,285 970,044 31,826,067

Maintenance Supplies 561,277 598,444 598,567 597,655 293,159 242,250 176,130 7,735,566

On-site licence fee 500,000 500,000 500,000 - - - - 1,500,000

Oxygen plant rental 2,400,000 2,400,000 2,400,000 2,400,000 2,400,000 2,400,000 2,400,000 38,400,000

TOTAL 17,860,909 18,779,742 18,782,844 18,259,807 10,571,375 9,285,940 7,616,418 247,547,561

TOTAL Euro/Con 60.46 59.62 59.62 58.05 68.51 72.83 82.16 60.80


20.4.4.6. General and Administrative Costs

The existing Deva Gold office located in the Minvest offices in Deva will continue to operate and will carry out the majority of the administrative duties for Certej such as accounting, translation and HR support. The total number of staff, 14, includes a full-time environmental officer and a lawyer for legal affairs. It is thought that there will be a need for 6 extra people to carry out the additional clerical, accounting, purchasing and HR duties required and these will be based at Certej in the admin block. The cost of these extra personnel will increase the Deva office support cost from €358k pa to €413k pa.

Safety wear has been included in the respective mining and process plant budgets. In the latter it is included at a cost of €100 per person pa.

The process plant manpower includes a safety officer and a training officer. Seven security staff are also included which is additional to the security dedicated for the gold-room duties. An IT expert is included in the process plant which is in addition to the IT support currently contracted out by Deva Gold.

Including a provision for supplies, utilities and miscellaneous costs the total estimated G & A costs will be €464k pa.

20.5. 13.5 Other Costs

Royalty

In Romania, recent legislation has increased the royalty on the production of precious metals to 4% of the total revenue. However, the Certej mining licence has a fixed royalty charge of 2% of mine production value, defined as net revenues after deduction of refining and smelting charges. This has recently been confirmed by the NAMR, and the royalty used in the financial model is calculated as 2% of net revenue.

Refining Costs

In May 2007 Brook Hunt carried out a survey for the sale of Certej doré to five European refiners, (Markets for Silver/Gold Doré, Certej Project, Romania). The estimated doré composition was provided by EGL and was based on the ME-ICP analysis of the loaded carbon from the Phase I testwork at HRL:


Table 20-38: Composition for Doré Market Survey

Element Proportion

Au 15%

Ag 62%

Cu 5.0 – 8.0%

Pb 1.0 – 2.5%

Zn 0 – 1.0%

Ni <4.5%

Te 0.029%

As 0.05% - 0.1%

Fe 0.05%

For this doré composition, the Valcambi quote at $96.7/oz. was the highest net revenue per ounce doré, for a gold and silver price of $600/oz. and $12/oz. respectively and assuming no penalties. The deductions imposed by the refinery covered freight and insurance, assays and handling, treatment and refining charges and payability. Using the quoted terms and a gold and silver price of $425/oz and $7/oz the Valcambi refinery deduction was 1.87% of the contained gold and silver value.

Albion Licence Fee

As agreed in the contract signed with Core Resources, there is a licence fee to use the Albion Process at Certej to be paid over 5 years. This fee is US$50k in Year -2 and US$120k in Year -1. In Years 1-3 the cost is US$650k pa.

Closure Costs

The closure costs of the TMFs have been estimated by Golders UK as €1.9 million in Year 16. A mine rehabilitation cost of €1.8 million was estimated. Total closure costs of €8 million have been provided for and have been built up as an annual bond payment of €0.5 million p.a. The balance of €4.3 million is estimated to cover the cost of decommissioning the plant and general mine site rehabilitation. This provision also meets the regulation within European Goldfields Mining permit that some 1% of total operating costs per annum be set aside as a rehabilitation bond.

20.5.1. Capital Costs

The three cost centres for the capital costs are the mining fleet, processing and infrastructure and the two TMFs.

20.5.1.1. Mine Capital Costs

The initial cost of the mining fleet, including all major and minor equipment and front end loader for the primary crusher, is €17.2 million. The sustaining cost for equipment replacement will be €6.5 million, and other sustaining capital of €900k. Mine prestrip of €8.5 million has been included in the financial model as a pre-operating expense in year -1.


20.5.1.2. Processing & Infrastructure Capital Costs

The total cost for the installation of the processing plant and associated equipment is estimated at €83.6 million.

The EPCM cost and the commissioning costs are estimated at €6.9 million and €1.0 million respectively giving a total construction cost for the processing plant of €91.5 million.

20.5.1.3. TMF Capital Costs

For the BFS, as designed by Golders UK, the initial cost will be €7.9 million and the sustaining capex for subsequent lifts until it reaches the design capacity will be €26.6 million.

20.5.1.4. Services and Other Costs

A total of €4.4 million for road construction has been included in the two pre-production years.

In pit dewatering plus dump ARD collection and pumping capital totals €1.9 million.

The cost of extending the power supply from the current position at the old Minvest Certej operation to a new substation adjacent to the new processing plant is estimated as €2.2 million.

The cost of raw water supply from the River Mures is covered in the processing plant capex.

20.5.1.5. Working Capital

Operating spares have been included in both the plant and mining capital estimates.


20.5.1.6. Total Construction Costs

The total initial capex for inclusion in the cash flow model is as follows.

Table 20-39: Capex Inputs to Cash flow Model

Area Item Cost € Millions

Mining Major Equip €15.2

Minor Equip €0.8

Site Prep. €1.2

TOTAL €17.2

Processing Construction €83.5

EPCM €6.8

Commissioning €1.1

Roads, Tunnel, Power, Services €8.4

TOTAL €117.0

TMF Initial Capex €7.9

TOTAL INITIAL PROJECT CAPEX €124.9

Mining Pre-Strip €8.5

TOTAL INITIAL PROJECT SPEND €133.4

20.5.2. Taxation and Depreciation

In Romania the current corporate tax rate is 16%. There are no current plans for the main political parties to increase this level.

Most methods for accounting depreciation are allowed in Romania provided a legitimate case can be made for their use. The most common for mining operations is to depreciate per tonne ROM ore extracted from the total resource. For tax depreciation at Certej, a 50% depreciation in Year 1 and 10% in the subsequent 5 years has been employed, in accordance with Romanian tax practice.

Under Romanian accounting legislation, losses may be carried forward 5 years.

Grants have not been included although Romania as a new EC member may be eligible as are new mining operations in other EC member states.

20.5.3. Project Financing

Financing is possible by a variety of measures and no final decision has as yet been taken.


20.5.4. Opportunities for Cost Reductions

The feasibility study incorporates the work done by a series of international industry consultants. In the second half of 2008, the dramatic fall in metal prices and global economic distress has led to a structural shift in the construction and procurement market, therefore European Goldfields’ management believe that there are a series of initiatives which are likely to result in cost reductions for the project. These have not as yet been incorporated in the financial model and therefore represent potential upside for the project.

20.5.4.1. Equipment

Since H2 2008 many capital projects have been cancelled or delayed and many operations have been shut down. There are now many opportunities for sourcing either second hand or new equipment at very competitive prices. The company has already identified opportunities relating to mills and the mining fleet as part of the equipment procurement process.

20.5.4.2. Contract Mining

€17.2 million of mining equipment has been included for owner mining equipment. A contract mining approach could save this upfront cost, although an increase in the operating costs will be the trade-off. Assuming the recovery of mining fleet capital (bought new) and a life of contract margin of 10%, mining costs would increase by 20% to €1.44 /tonne.

20.5.4.3. Pre-strip and Waste Mining

There is a significant road building programme under way in Romania currently, and waste rock for road foundations is in great demand. A road contract in the Deva area is being awarded and local management are in discussions with road contractors about selling quantities of waste rock from the current pit or historic stockpiles. As a minimum, there is the likelihood of selling the pre-strip material at cost for road construction use (ie zero cost to the project). There may be more favourable life of mine alternatives, whereby all project waste could be sold into this road building programme and general aggregates market in this region.

In addition, the Certej project perimeter includes other potential sources of high quality andesite which could be sold in addition to waste material.

20.5.4.4. Romanian Government Grants

The Romanian Government has a wide ranging programme for the stimulation of investment and economic activity, particularly in under-developed areas. The Certej project is eligible for a number of different grants under current legislation. In particular, Certej is eligible for the recent 5 year State Aid programme to promote regional development by the stimulation of investment and the creation of new jobs.


Under Government Decision 1165/2007 dated 2 September 2007, Romanian companies which have a registered office and perform their business in Romania are eligible for investment support, subject to the fulfilment of 2 main criteria:

• Investment of at least €30 million;

• Creation of at least 300 new jobs as a result of the new investment.

Monetary support under this programme is limited to a maximum of €28 million, and the mining industry is an eligible industry. The total budget for this programme is €500 million. No EU approvals are required for disbursement under this programme.

Since the Certej project is part of an eligible industry, and satisfies the key criteria, we would expect to benefit from up to €28 million cash inflow to the project during project construction.

In addition to this, broader structural development funding is available in Romania from the EU, as defined under Government Decision 651/2006, which sets out the regulation for State Aid over the period from 2006-2013. This programme provides State Aid to cover 50% of initial investment programmes plus a further 50% of labour costs in the first 2 years of operations. Disbursements under these programmes are subject to EU approval.

20.5.5. Project Returns and Sensitivity

Assuming a gold price of US$650/ Oz and US$7.50 /Oz for silver, the 16-year cash flow model was run for the base case:

The sensitivity of the Base case to the key variables of mining cost, total processing cost, Albion processing cost and capital expenditure cost is illustrated graphically in Error! eference source not found. for NPV and Error! Reference source not found. for IRR.

Table 20-40: Cash flow Model Indicators

Case Payback NPV 5% post tax

€m

IRR

%

Base case 4.3 years €88.5 21.3%

The impact of changes to commodity prices is set out in the tables below. Given that the current commodity price and the levels over the past two years have been significantly higher than those which have been used for the Base case analysis, we set out the changes to NPV and IRR by applying prices ranging from US$500 /Oz - $1,000 /Oz for gold, and US$8 /Oz – US$20 /Oz for silver:


Table 20-41: NPV Sensitivity to Commodity Prices

NPV 5% (in € millions)

$/oz Gold

500 600 700 800 900 1,000

Silver 8 (36) 50 131 212 293 373

10 (25) 60 141 222 302 383

12 (13) 70 151 232 312 393

14 (2) 80 161 241 322 402

16 8 90 171 251 332 412

18 18 99 180 261 341 422

20 28 109 190 271 351 432

Table 20-42: IRR Sensitivity to Commodity Prices

IRR (in %)

Gold 500 600 700 800 900 1,000

Silver 8 (22.8) 14.8 27.8 39.1 49.3 58.8

10 (1.0) 16.3 29.0 40.1 50.1 59.6

12 2.2 17.7 30.1 41.0 51.0 60.4

14 4.5 19.1 31.2 42.0 51.9 61.2

16 6.5 20.5 32.3 42.9 52.7 62.0

18 8.3 21.7 33.3 43.8 53.6 62.8

20 9.9 23.0 34.4 44.7 54.4 63.5

Table 20-43: Payback Sensitivity to Commodity Prices

Payback (in years)

$/oz Gold

500 600 700 800 900 1,000

Silver 8 16.0 5.6 2.8 1.9 1.6 1.3

10 16.0 5.3 2.7 1.9 1.6 1.3

12 9.4 5.1 2.6 1.9 1.5 1.3

14 8.3 4.7 2.5 1.8 1.5 1.3

16 7.6 4.4 2.5 1.8 1.5 1.3

18 7.0 4.1 2.4 1.8 1.5 1.3

20 6.7 3.8 2.3 1.8 1.5 1.3


20.5.6. Benefits to the Region and to Romania

The Certej project is one of the largest proposed direct inward investments into the region and also the whole of Romania.

20.5.6.1. Employment

The project is expected to employ directly over 300 employees from Certej and the immediate vicinity. Accumulating the total salary costs of the different cost areas, a total of €36.6 million is paid to and behalf of employees during the life of the project. Of this, almost €20 million is net salaries paid directly into the Certej and Deva communities. Applying a 3 times multiplier on direct employment, it is estimated that a further 900 jobs will be created in services to the mine, both direct and indirect. Given the high levels of unemployment in the area following the closure of the former state run mining operations, this is a key benefit for the local area. As described in the section above on Government Grants, the Romanian Government has implemented a series of incentive based investment measures in order to stimulate investment with a view to creating jobs in underdeveloped areas or areas which have suffered major economic contraction in the post communist era.

20.5.6.2. Fiscal Benefits

In addition to the direct and indirect employment benefits, the Romanian Government will enjoy significant fiscal benefits. The Romanian Government will receive a total of €63 million in royalties, corporation taxes and social contributions. Of this, €26 million represents corporation taxes, €20 million of royalties and €17 million of social contributions. This €63 million of fiscal benefits to Romania represents over 39% of the base case post-tax cash flow accruing from the Certej project to equity holders over the life of mine.

20.5.7. Conclusions

The Certej project yields a positive return at US$650 /Oz gold and US$7.50 /Oz silver. Lower commodity prices represent the single most significant risk for the project and appropriate price mitigation should be put in place to ensure that capital repayments and operating margins are protected. The areas of cost opportunity should be investigated and confirmed as soon as possible.


Figure 20-35: NPV Sensitivity Plot

Figure 20-36: IRR Sensitivity Plot

45,00055,00065,00075,00085,00095,000

105,000115,000

-15% -10% -5% 0% 5% 10% 15%

Euro

Sensitivity

NPV Sensitivity

Mining Cost

Total processing cost

Albion cost

Capex spend

12.00%

17.00%

22.00%

27.00%

32.00%

-15% -10% -5% 0% 5% 10% 15%

IRR

%

Sensitivity

IRR Sensitivity

Mining cost

Total processing cost

Albion cost

Capex spend


21. INTERPRETATION AND CONCLUSIONS

21.1. Geology and Resources

The geology of the Certej locality and deposit is well established and a detailed regional and local model for the mineralisation has been put together by European Goldfields and Deva Gold. Detailed drilling, data analysis and modelling have resulted in the development of a robust geological and grade model. The resource estimate is well supported with over 60,000 metres of drilling and 22,000 metres of channel sampling. A significant proportion of the channel and drilling data collection and QAQC was completed under the independent supervision of consultants RSG. The QAQC programme is above industry standards and detailed analysis has shown no bias in either assay precision or accuracy. All assaying has been completed at reputable internationally recognised independent laboratories and samples submitted with internationally recognised standards from Rocklabs of New Zealand. Sample quality from RC and diamond methods is exceptional and underground channels are still preserved today for any future reference. All sample pulps are stored in a dry and locked facility on site in Certej. All diamond core and chip trays of RC samples are also stored in locked facilities at Certej. Geological wire framing and three dimensional modelling of domains has been carried out based on careful mapping of four mineralogical zones, East, Intermediate, Central and West. These have distinct mineralogical signatures and correspondingly distinctive gold deportment. Their boundaries correspond to litho structural features which have also been modelled in three dimensions. Sampling of all of these zones has been representative. Block modelling and interpolation using ordinary kriging and uniform conditioning in to a selective mining unit block model have resulted in a Measured and Indicated resource of 41.5Mt at 2.0g/t Au and 11g/t Ag grade. In addition, an Indicated resource of 7,022,434 tonnes at 0.53 Au g/t and 8.9 g/t Ag, with a contained metal of 119,600 oz Au and 2,002,100 oz Ag has been estimated in existing dumps using a polygonal wireframe approach. The resource and reserve has been subject to several audits by RSG during the course of the projects development. The resource and reserve are categorised to CIM standards and as such meet the requirements of the Canadian National Instrument 43-101 and can be viewed as current.


Analysis has shown that the resource estimation approach using a selective mining unit (SMU) model calculated using uniform conditioning of ordinarily kriged parent blocks by way of uniform conditioning is valid and a representative way of modelling the Certej Mineralisation. Comparison of the resource to internally produced polygonal estimate validates the SMU estimate. The Resource estimate is robust and has been completed several times by RSG with very similar results as well as having been checked internally twice with a less than 3% difference in volume.

21.2. Mining

The Certej open pit design, equipping and costing follows traditional design procedures. The proposed technology and designs have been used for decades in operations in more rigorous conditions. The mining of the Certej deposit is considered to be a low risk means of extracting the identified mineral reserves. The engineered pit designs were based on Whittle optimisation to determine the general pit shape and depth. This comprised a main pit over the east, intermediate and central zones of the orebody and an adjacent smaller pit over the west zone. The final pit depth is some 220 m in the East pit, and some 200 m in the West pit. Optimised pit shells generated in Whittle were used as guidance to detail engineer the pits, including exact ramp positions and bench configuration. Average overall pit slopes including ramps are generally 40° to 60° according to geotechnical conditions but flatten to 30° in some places at the pit rim. Total ore probable reserves from the open pits is 32.8 Mt at 2.0 g/t for gold and 11.4 g/t for silver, at an approximate gold cut-off grade of 0.8 g/t Au, and 92.2 Mt of waste, for a LoM strip ratio of 3.1:1 (t:t). Some 9.7 Mt of waste are used for building roads and pads, 18.7 Mt are used for in-pit dumping in the East pit, and some 63.8 Mt are stored between the North and South dumps. An additional 7.8 Mt of in-pit reserves at 0.7g/t gold and 14 g/t silver becomes economic to process at $650 per oz Au and $7.5 per ounce Ag. This lower grade reserve is classified as probable and will be processed at the end of mine life. Existing dumps from historic production contain probable reserves of 6.3Mt at 0.53 g/t Au and 8.9 g/t Ag. This material will be processed at the end of mine life with the lower grade in-pit material. Total Reserves of some 47Mt will support a mine life of 16 years. An owner operated open pit mining fleet has been selected to suit the required mining schedule, with the key equipment being 8.5 m3 hydraulic excavators and 62 t trucks, capable of excavating and moving up to 15 Mt of material per year. In peak years, the excess capacity of the re-handle loader could be used to supplement the excavators.


A mining schedule has been prepared for 3 Mtpa of run of mine ore through the mill. Staffing is 231 people. EGL consider the schedule and costs to be within +15% and -10% as is normal for a Feasibility Study.

21.3. Project Conclusions

The Certej project yields a positive return at US$650 /Oz gold and US$12 /Oz silver. Lower commodity prices represent a significant risk for the project and appropriate price mitigation should be put in place to ensure that capital repayments and operating margins are protected. The areas of cost opportunity should be investigated and confirmed as soon as possible.


22. RECOMMENDATIONS

The Certej Definitive Feasibility Study, summarised in this report, indicates a viable project and it is recommended that it should now progress to basic and final engineering stage prior to financing and construction on approval of the relevant environmental and construction permits.


23. REFERENCES

• Certej Gold Silver Project, Romania, Technical Report. RSG Global Consulting Pty

Ltd (Now Coffey Mining). November 2007

• Certej Gold Project – Romania, Process Plant and Infrastructure Cost Engineering Study Report. Aker Kvaerner Engineering Services Ltd. February 2008.

• Markets for Gold/Silver Dore. Brook Hunt. May 2007.

• Certej Project: Macrisului Tailings Management Facility Bankable Feasibility Study Romania – Golders Associates (UK) Ltd

• Petrographic Evaluation of Twenty Nine Core samples from the Certej Prospect, Romania, TLC Report number 01062” prepared for Deva Gold, Terry Leach, August 2001.

• Geological Study on Gold and Silver Mineralisation of Coranda Mica, Coranda Open Pit and Dealul Grozii Deposits with a view to determine the Distribution of Gold and Silver within the Mineral Assemblages; Institute of Geological Prospecting, Bucharest, Romania, 1993.

• Mineralogical Investigation of polished samples from the Certej Deposit for Ammtec, Roger Townsend & Associates, 2001.

• Mineralogical Examination of Five Gold Ore Samples from the Certej Deposit, Romania; SGS Lakefield Research Limited report LR10476-001 M15506-SEP02, December 12, 2002.

• Deportment Study of Gold in Sample 662 and Sample 663 from the Certej Project, SGS Lakefield Research Limited report LR10476-005/M15015-JUL05, September 15, 2005.

• A Gold deportment study of the 670 LCT Concentrate and 670 LCT Tail from the Certej Project, SGS Lakefield Research Limited report 11471-001 Final report, March 8, 2007.

• QEMSCAN Analysis of a Certej Leach Residue and Concentrate from the Albion Process. Report from Intellection Pty Ltd of Australia, September 2006.

• Deportment of Gold and Silver in Certej Flotation Concentrates and Tails, Amtel report 07/34, June 28 2007.

• Deportment of Gold and Silver in Certej’s Albion Products, Amtel report 07/46, December 1 2007.

• Report on Certej Project Tailings Management Facilities Bankable Feasibility Study Romania. Golder Associates (UK) Ltd. March 2008.

• Proposed Grinding System for Certej Deposit based on Small-scale Data. SGS Lakefield Research. 2005

• Flotation Optimization of Four Samples of Gold Mineralization from the Certej Deposit. SGS Lakefield Research. Jan 2006.


• Proposed Grinding System for the Certej Circuit based on Bench-Scale Data. SGS Lakefield Research. Nov 2006.

• The Detoxification of Cyanide Leach Tailings. SGS Lakefield Research Feb 2007.

• Metallurgical Study on Refractory Gold-Pyrite Concentrate from the Certej Project – Hazen Project 10318. Hazen Research Inc. 2007.

• An Investigation into the Liquid-Solids Separation of the Certej Deposit Flotation Test Product. SGS Lakefield Research. September 2007.

• An Investigation into the Liquid-Solids Separation of the Certej Deposit CIL Tailings Product. SGS Lakefield Research. September 2007

• Technical Memorandum 0850: Albion Process Testing of Certej West Concentrate. HRL Testing. October 2007.

• Technical Memorandum 0874: Continuous Albion Process Leaching of Certej Concentrate Pilot Plant Stage 2 and Associated Process Design Test Work. HRL Testing. October 2007.

• Preparation of Bulk Flotation Concentrates from Four Samples of Certej Drillcore at the Hellas Gold Laboratory – Greece & Interpretation of Results. Hellas Gold. 2007.

• Additional Simulation for the Certej Circuit Based on Small-Scale Data. SGS Lakefield. Nov 2007.

• Report on Laboratory Testing of CYTEC Reagents on Certej Ore, May 2008, Mike Peart, Certej Industries CV, May 2008.


CERTIFICATE OF AUTHOR I, Mr Patrick William Forward, do hereby certify that:

1. I am General Manager, Exploration at European Goldfields (Services) Limited of 11 Berkeley Street, Level 3, London, UK, W1J 8DS. My residential address is 14 Cochrane Road, London, United Kingdom.

2. I am a graduate of the Imperial College of Science and Technology, London and

hold a B.Sc. honours degree Mining Geology (1989). 3. I am a Member of the Australasian Institute of Mining and Metallurgy

(Membership 225134) and an Associate of the Royal School of Mines. 4. I have practiced my profession for a total of 16 years since my graduation from

university. 5. I have read the definition of “qualified person” set out in National Instrument 43-

101 (“NI 43-101”) and certify that I am a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the preparation of Sections 3 to 17, 19, 20.2 to 20.5 and 21

to 23 of the technical report titled “Certej Updated Definitive Feasibility Study Summary Technical Report” and dated 26th of February, 2009 (the “Technical Report”) relating to the Certej property. I have visited the Certej property on several occasions during 2004, 2006, 2007 and 2008 for an approximate total of 150 days with my most rescent visit being November 17th to the 20th 2008.

7. I have had no prior involvement with the property that is the subject of the

Technical Report. I am not aware of any limitations imposed upon my access to persons, information, data or documents that I consider relevant to the subject matter of the Technical Report.

8. 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 the Technical Report not misleading.

9. I am not independent of European Goldfields Limited and Deva Gold pursuant to

section 1.4 of NI 43-101. 10. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been

prepared in compliance with that instrument and form.

11. I own securities of European Goldfields Limited in the form of share options and shares, and as such I have an indirect interest in the Certej property.

Dated this 26th day of February, 2009.

Patrick William Forward


CERTIFICATE OF AUTHOR I, Gordon Antony Jackson, do hereby certify that:

1. I am Consultant Metallurgist at TJ Metallurgical Services of 6 Coxburn Brae, Bridge of Allan, Stirling, United Kingdom. My residential address is 6 Coxburn Brae, Bridge of Allan, Stirling, United Kingdom.

2. I graduated with a B.Sc. Hons. (Eng) from Royal School of Mines, Imperial

College of Science, Technology & Medicine, London University in 1980. 3. I am a Fellow of the Institute of Materials, Minerals and Mining. 4. I have practiced my profession for a total of 28 years since my graduation from

university. 5. I have read the definition of “qualified person” set out in National Instrument 43-

101 (“NI 43-101”) and certify that I am a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the preparation of Section 18 (Mineral Processing and

Metallurgical Testing) of the technical report titled “Certej Updated Definitive Feasibility Study Summary Technical Report” and dated 26th of February, 2009 (the “Technical Report”) relating to the Certej property. I have visited the Certej property on many occasions during 2005, 2006, 2007, 2008 and 2009 for more than 330 days with my most recent visit being February 15th to the 20th 2009.

7. I have had no prior involvement with the property that is the subject of the

Technical Report. I am not aware of any limitations imposed upon my access to persons, information, data or documents that I consider relevant to the subject matter of the Technical Report.

8. 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 the Technical Report not misleading.

9. I am not independent of European Goldfields Limited and Deva Gold pursuant to

section 1.4 of NI 43-101. 10. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been

prepared in compliance with that instrument and form.

11. I own securities of European Goldfields Limited in the form of shares, and as such I have an indirect interest in the Certej property.

Dated this 26th day of February, 2009.

Gordon Antony Jackson

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