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Date: 11 October 2013 Document No: 0482-RPT-001 Rev 0

Woxna Graphite Restart Project

Preliminary Economic Analysis (PEA)

Prepared For:

Flinders Resources

Compiled By:

GBM Minerals Engineering Consultants

Effective Issue Date:

11 October 2013

Qualified Persons:

Chris Stinton, B Sc, CEng MIMMM Geoffrey Reed, B App Sc, MAusIMM Bryan Pullman, B.Sc.(Eng.), P.Eng APEGA Henning Holmström, M Sc, PhD, MAusIMM

Contributing Engineers:

GBM Minerals Engineering Consultants Limited Reed Leyton Consultants Ltd Golder Associates AB Tailings Consultants Scandinavia

Preliminary Economic Analysis (PEA) - 0482-RPT-001 Rev 0

Revision History

Date Rev Reason Prepared Checked Approved

11/10/13 0 Approved for SEDAR CS/GR/BP/HH CS/IF IF


TABLE OF CONTENTS

SECTION 1 SUMMARY ................................................................................................................ 18

1.1 Background ............................................................................................................. 18

1.2 Introduction ............................................................................................................. 18

1.3 Reliance on Other Experts ..................................................................................... 18

1.4 Property Description and Location ....................................................................... 18

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

1.6 History ...................................................................................................................... 19

1.7 Geological Setting and Mineralization .................................................................. 20

1.8 Deposit Types ......................................................................................................... 20

1.9 Exploration .............................................................................................................. 20

1.10 Drilling ...................................................................................................................... 20

1.11 Sample Preparation, Analyses and Security ....................................................... 21

1.12 Data Verification ...................................................................................................... 21

1.13 Mineral Processing and Metallurgical Testing .................................................... 22

1.14 Mineral Resource Estimates .................................................................................. 23

1.15 Mineral Reserve Estimates .................................................................................... 24

1.16 Mining Methods ....................................................................................................... 24

1.17 Recovery Methods .................................................................................................. 29

1.18 Project Infrastructure ............................................................................................. 29

1.19 Market Studies and Contracts ............................................................................... 30

1.20 Environmental Studies, Permitting and Social or Community Impact .............. 31

1.21 Capital and Operating Costs ................................................................................. 32

1.22 Economic Analysis ................................................................................................. 33

1.23 Adjacent Properties ................................................................................................ 33

1.24 Interpretations and Conclusions ........................................................................... 33

1.25 Recommendations .................................................................................................. 34

SECTION 2 INTRODUCTION ....................................................................................................... 35


2.1 Client ........................................................................................................................ 35

2.2 Terms of Reference and Purpose ......................................................................... 35

2.3 Sources of Information........................................................................................... 35

SECTION 3 RELIANCE ON OTHER EXPERTS .......................................................................... 36

3.1 Qualified Persons ................................................................................................... 36

3.2 Reliance on Other Experts ..................................................................................... 36

3.3 Verification .............................................................................................................. 37

3.4 Financial Interest Disclaimer ................................................................................. 37

SECTION 4 PROPERTY DESCRIPTION AND LOCATION ........................................................ 38

4.1 Location ................................................................................................................... 38

4.2 Mining Laws and Regulations ............................................................................... 39

4.3 Permits and Concessions ...................................................................................... 41

4.4 Acquisition of Land ................................................................................................ 43

4.5 Taxes and Duties .................................................................................................... 43

4.6 Environmental Considerations .............................................................................. 43

4.7 Flinders’ Property Locations ................................................................................. 44

SECTION 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND

PHYSIOGRAPHY ..................................................................................................... 45

5.1 Access ..................................................................................................................... 45

5.2 Physiography .......................................................................................................... 45

5.3 Climate ..................................................................................................................... 45

5.4 Water, Power and Telecoms .................................................................................. 46

5.5 Local Resources ..................................................................................................... 46

SECTION 6 HISTORY ................................................................................................................... 48

6.1 Historical Deposit Ownership and Exploration ................................................... 48

6.2 Sampling and Chemical Analysis ......................................................................... 49

6.3 Density Determination ............................................................................................ 49

6.4 Petrology ................................................................................................................. 49

6.5 Historical Mineral Resource Estimates ................................................................ 49


SECTION 7 GEOLOGICAL SETTING AND MINERALIZATION ................................................. 51

7.1 Regional Geology ................................................................................................... 51

7.2 Local Geology ......................................................................................................... 53

SECTION 8 DEPOSIT TYPES ...................................................................................................... 54

SECTION 9 EXPLORATION......................................................................................................... 57

SECTION 10 DRILLING .................................................................................................................. 59

SECTION 11 SAMPLE PREPARATION, ANALYSES, AND SECURITY ..................................... 64

SECTION 12 DATA VERIFICATION .............................................................................................. 66

12.1 Drill Core .................................................................................................................. 66

12.2 Check Sampling ...................................................................................................... 67

12.3 Check Analyses ...................................................................................................... 69

12.4 Results and Discussion ......................................................................................... 70

12.5 Density ..................................................................................................................... 73

SECTION 13 MINERAL TESTING AND PROCESSING ............................................................... 76

13.1 Recent Test work .................................................................................................... 76

SECTION 14 MINERAL RESOURCE ESTIMATES ....................................................................... 80

14.1 Resource Data ......................................................................................................... 80

14.2 Mineral Resource Estimate .................................................................................... 89

14.3 Discussion ............................................................................................................... 90

14.4 NI 43-101 Compliance ............................................................................................. 90

SECTION 15 MINERAL RESERVE ESTIMATES .......................................................................... 98

SECTION 16 MINING METHODS .................................................................................................. 99

16.1 Introduction ............................................................................................................. 99

16.2 Geology and Mineral Resources ......................................................................... 101

16.3 Geological Block Model ....................................................................................... 102

16.4 Whittle Optimisation ............................................................................................. 106

16.5 Pit Slope Design Concept .................................................................................... 115

16.6 Open Pit Mining ..................................................................................................... 121


16.7 Open Pit Mining Operations ................................................................................ 137

16.8 Waste Rock Storage and Management Facility ................................................. 144

16.9 Mining Infrastructures .......................................................................................... 148

SECTION 17 RECOVERY METHODS ......................................................................................... 150

17.1 History and Background ...................................................................................... 150

17.2 Basis for Design .................................................................................................... 150

17.3 Improvements to the Plant ................................................................................... 156

17.4 Basis for the Estimate of Graphite Production .................................................. 157

17.5 Further Beneficiation ............................................................................................ 166

SECTION 18 PROJECT INFRASTRUCTURE ............................................................................. 168

18.1 General ................................................................................................................... 168

18.2 Electrical ................................................................................................................ 170

18.3 Tailings Management Facility (TMF) ................................................................... 173

18.4 Clarification Pond (Lake Uxatjärn) ...................................................................... 176

18.5 Emergency preparedness .................................................................................... 178

18.6 Water Management ............................................................................................... 178

SECTION 19 MARKET STUDIES AND CONTRACTS ................................................................ 179

19.1 Introduction ........................................................................................................... 179

19.2 Market Description ................................................................................................ 179

19.3 Natural Graphite Demand .................................................................................... 180

19.4 Natural Graphite Supply ....................................................................................... 180

19.5 Natural Graphite Market Price Trends ................................................................ 181

19.6 Pricing Basis for the Woxna Graphite Restart Project ..................................... 183

19.7 Production of Saleable Graphite ......................................................................... 183

SECTION 20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY

IMPACT ................................................................................................................... 185

20.1 Environmental Permitting Requirements and Status ....................................... 185

20.2 Further Permits ..................................................................................................... 186

SECTION 21 CAPITAL AND OPERATING COSTS .................................................................... 188


21.1 Basis of Estimate .................................................................................................. 188

21.2 Capital Cost ........................................................................................................... 192

21.3 Operational Cost ................................................................................................... 198

SECTION 22 ECONOMIC ANALYSIS ......................................................................................... 205

Methodology .......................................................................................................... 205 22.1

22.2 Economic Performance ........................................................................................ 206

22.3 Discounted Cash Flow Analysis ......................................................................... 208

22.4 Sensitivity Analysis .............................................................................................. 209

SECTION 23 ADJACENT PROPERTIES ..................................................................................... 212

SECTION 24 OTHER RELEVANT DATA AND INFORMATION ................................................. 213

24.1 Execution Schedule .............................................................................................. 213

SECTION 25 INTERPRETATION AND CONCLUSIONS ............................................................ 214

25.1 Mineral Resource Estimates ................................................................................ 214

25.2 Mining Methods ..................................................................................................... 214

25.3 Metallurgical Test Work and Processing ........................................................... 215

25.4 Organisational Structure ...................................................................................... 215

25.5 Tailings Refurbishment and Expansion ............................................................. 215

25.6 Production and Graphite Sale ............................................................................. 216

25.7 Infrastructure ......................................................................................................... 216

25.8 Financial Performance ......................................................................................... 216

25.9 Risks and Opportunities ...................................................................................... 217

SECTION 26 RECOMMENDATIONS ........................................................................................... 220

26.1 Mineral Resource and Reserve Estimates ......................................................... 220

26.2 Metallurgical Test Work ....................................................................................... 220

26.3 Mining .................................................................................................................... 221

26.4 Processing ............................................................................................................. 222

26.5 Organisational Structure ...................................................................................... 222

26.6 Tailings .................................................................................................................. 223


26.7 Production and Graphite Sale ............................................................................. 223

SECTION 27 REFERENCES ........................................................................................................ 224

APPENDIX A LOM Schedule........................................................................................................ 226


LIST OF TABLES

Table 1-1: Results of the Locked Cycle Flotation Tests ....................................................................... 23

Table 1-2: Expected Product Tonnages and Recovery ........................................................................ 23

Table 1-3: Resource Grade and Tonnage (at 7% Cg cut-off grade) .................................................... 24

Table 1-4: Golder Selection of Whittle Pit Shell for Kringelgruvan Deposit Ultimate Open Pit ............. 25

Table 1-5: Materials Contained in Pits .................................................................................................. 26

Table 1-6: Production Schedule Summary ........................................................................................... 27

Table 1-8: Initial Capital Cost (M USD) ................................................................................................. 32

Table 1-9: Operating Cost Summary .................................................................................................... 32

Table 1-10 - Summary Economic Analysis ........................................................................................... 33

Table 4-1: Woxna Graphite Project Tenure .......................................................................................... 39

Table 4-2: Woxna Graphite Project – Kringelgruvan nr 1 Mining Concession Co-ordinates ................ 44

Table 6-1: Drilling History of the Kringelgruvan Project ........................................................................ 48

Table 6-2: Woxna Graphite Project Historical Resources (after Claesson, 2002) ................................ 49

Table 9-1: Flinders Drilling of the Kringelgruvan Project ...................................................................... 58

Table 10-1: Kringelgruvan Drill Collar Co-ordinates (SWEREF99 TM Grid) ........................................ 60

Table 12-1: Drillholes and Core Examined by Reed Leyton Representative ....................................... 66

Table 12-2: Check Sample Intervals by Reed Leyton Representative ................................................. 67

Table 12-3: Drill Core Re-sampled for Check Analysis, Matched with Original Assays ....................... 70

Table 13-1: Results of the Locked Cycle Flotation Tests ..................................................................... 78

Table 14-1: Kringelgruvan Drilling Database Summary ........................................................................ 81

Table 14-2: Kringelgruvan ‘hg’ Domain (above 7 % Cg) and ‘grf’ Domain (lith code FGRF) .............. 85

Table 14-3: Kringelgruvan Domain Volume Validation ......................................................................... 85

Table 14-4: Block Model Parameters for Kringelgruvan ....................................................................... 87

Table 14-5: Search Parameters for Kringelgruvan ............................................................................... 88

Table 14-6: Estimation Parameters for Kringelgruvan .......................................................................... 88

Table 14-7: Kringelgruvan Mineral Resource Estimate @ 7 % Cut-Off ................................................ 89

Table 14-7: Kringelgruvan Mineral Resource Estimate @ 7 % Cut-Off ................................................ 89


Table 14-8: Kringelgruvan Graphite Combined Resource Grade and Cumulative Tonnage at Various

Cut-Off Grades ...................................................................................................................................... 90

Table 14-9: Kringelgruvan – Drill holes and Intervals Used in Resource Calculation .......................... 90

Table 16-1: Pit Optimisation Parameters ............................................................................................ 100

Table 16-2: Kringelgruvan Deposit 2013 Mineral Resource Estimate ................................................ 101

Table 16-3: Density by Material Type in Geological Block Model “vie_voxna_2013apr_75.csv” ....... 103

Table 16-4: Kringelgruvan Geological Block Model Parameters ........................................................ 103

Table 16-5: Geological Block Model Key Variables ............................................................................ 104

Table 16-6: Kringelgruvan Geological Model In-situ Tonnes and Grade ............................................ 104

Table 16-7: Kringelgruvan Geological Model In-situ Tonnes and Grade Using a 7 % cut-off Grade . 104

Table 16-8: Kringelgruvan Geological Model In-situ Tonnes and Grade Per Mineral Type ............... 105

Table 16-9: Topography and Geological Boundaries ......................................................................... 105

Table 16-10: Kringelgruvan Datamine Geological Model Tonnes and Grade (no cut-off applied) ..... 106

Table 16-11: Optimisation Block Model Tonnes ................................................................................. 107

Table 16-13: Kringelgruvan Whittle Pit by Pit Summary of Results .................................................... 111

Table 16-14: Selected Pit Shell and Whittle Results for Kringelgruvan Deposit ................................. 114

Table 16-15: Open Pit Design Parameters ......................................................................................... 122

Table 16-16: Summary of Total In-pit Resources ............................................................................... 125

Table 16-17: Summary of the Year 1 - 2 Production Schedule .......................................................... 128

Table 16-19: Kringelgruvan 10 Year Mine Production Schedule ........................................................ 133

Table 16-20: Kringelgruvan LOM Production Summary ..................................................................... 135

Table 16-23: Preliminary Mine Equipment List ................................................................................... 142

Table 16-24: Mining Workforce Estimate ............................................................................................ 142

Table 16-25: Estimated Mine Staffing Requirement ........................................................................... 143

Table 17-1: Crushing Criteria .............................................................................................................. 157

Table 17-2: Fine Ore Bin Parameters ................................................................................................. 157

Table 17-3: Primary Milling Criteria .................................................................................................... 158

Table 17-4: Flash Flotation Criteria ..................................................................................................... 158

Table 17-5: Rougher Flotation Criteria ................................................................................................ 159


Table 17-6: Scavenger 1 Flotation Criteria ......................................................................................... 159

Table 17-7: Scavenger 2 Flotation Criteria ......................................................................................... 159

Table 17-8: Scavenger Cleaner 1 Flotation Criteria ........................................................................... 159

Table 17-9: Scavenger Cleaner 2 Flotation Criteria ........................................................................... 160

Table 17-10: Scavenger Cleaner 3 Flotation Criteria ......................................................................... 160

Table 17-11: Rougher Cleaner 1 Flotation Criteria ............................................................................. 160

Table 17-12: Rougher Cleaner 2 Flotation Criteria ............................................................................. 160

Table 17-13: Regrind Mill Criteria (all SMD’s) ..................................................................................... 161

Table 17-14: Concentrate Dewatering Criteria ................................................................................... 161

Table 17-15: Bagging Plant ................................................................................................................. 161

Table 17-16: Flotation Reagent Requirements ................................................................................... 161

Table 17-17: Screen Analysis of Concentrates of the Locked Cycle Test .......................................... 163

Table 17-18: Summary of Available Production Reports Year 2000 .................................................. 164

Table 17-19: Coarse Flake Production Specification .......................................................................... 164

Table 17-20: Medium Flake Production Specification ....................................................................... 165

Table 17-21: Fine Flake Production Specification .............................................................................. 165

Table 19-1: Average Sale Price Estimation ........................................................................................ 183

Table 20-1: Woxna Graphite Project Tenure ...................................................................................... 186

Table 21-1: Operating Inputs .............................................................................................................. 188

Table 21-2: Supporting Documents .................................................................................................... 188

Table 21-3: Project Area Breakdown .................................................................................................. 189

Table 21-4: Currency Exchange Rate ................................................................................................. 190

Table 21-5: Initial Capital Cost (M USD) ............................................................................................. 193

Table 21-6: Direct Cost Centre Factors .............................................................................................. 195

Table 21-7: Indirect Cost Centre Definitions ....................................................................................... 196

Table 21-8 - Indirect Cost Factors ...................................................................................................... 197

Table 21-9: Deferred Capital Investment ............................................................................................ 197

Table 21-10: Operational Costs .......................................................................................................... 198


Table 21-11: Summary of Contractor Operating Costs ...................................................................... 200

Table 21-12: Summary Power Requirements ..................................................................................... 203

Table 22-1: Depreciation and Loss Parameters ................................................................................. 205

Table 22-2: Base case financial inputs ............................................................................................... 205

Table 22-3: Historical Book Assets ..................................................................................................... 206

Table 22-4: Economics Summary ....................................................................................................... 206

Table 22-5: Discounted Cash Flow Analysis ...................................................................................... 208

Table 22-6: Sensitivity Analysis Results ............................................................................................. 209

Table 24-1: Indicative Execution Schedule ......................................................................................... 213

Table 26-1: Indicative Resource Development Costs ......................................................................... 220


LIST OF FIGURES

Figure 1-1: Pit Designs .......................................................................................................................... 26

Figure 1-2: LOM Production Schedule .................................................................................................. 27

Figure 1-3: Final Status Map of the Site ............................................................................................... 28

Figure 4-1: Location of Flinders Resources Graphite Projects, Sweden .............................................. 38

Figure 5-1: Topography and Access of the Kringelgruvan Project Area .............................................. 47

Figure 7-1: Regional Geology of the Kringelgruvan Project Area ......................................................... 52

Figure 7-2: Local Geology of the Kringelgruvan Project Area .............................................................. 53

Figure 8-1: Graphite Mineralisation of the Kringelgruvan Project Area ................................................ 55

Figure 8-2: Graphite Mineralisation of the Kringelgruvan Project Area ................................................ 56

Figure 10-1: Drilling at the Kringelgruvan Project Area ........................................................................ 60

Figure 12-1: Coarse Duplicate Data for Cg ........................................................................................... 70

Figure 12-2: Pulp Duplicate Data for Cg ............................................................................................... 70

Figure 12-3: Paired Historical and Modern Analytical Data for Cg ....................................................... 73

Figure 12-4: All 1 424 Bulk Density Determinations ............................................................................. 74

Figure 12-5: Domain ‘HG’ 376 Bulk Density Determinations ................................................................ 74

Figure 12-6: Domain ‘GRF’ 540 Bulk Density Determinations .............................................................. 75

Figure 13-1: Locked Cycle Flotation Test Diagram .............................................................................. 79

Figure 14-1: Histogram of Raw Sample Lengths for Kringelgruvan ..................................................... 83

Figure 14-2: Mineral Resource Cross Section, Kringelgruvan .............................................................. 84

Figure 14-3: Mineral Resource Cross Section, Kringelgruvan .............................................................. 86

Figure 16-1: Screenshot of Imported Whittle Block Models Summary for Kringelgruvan .................. 108

Figure 16-2: Whittle Resultant Pits (RF 0.85 – Magenta, RF 1.0 – Blue) ........................................... 112

Figure 16-3: Graph of Kringelgruvan Graphite and Waste Tonnage Versus Revenue Factor and

Discounted Net Value ......................................................................................................................... 112

Figure 16-4: Kringelgruvan Graph of Results Pit Discounted Net Value and Tonnes of Graphite by

Revenue Factor ................................................................................................................................... 113

Figure 16-5: Kringelgruvan Spider Graph Sensitivity of Pit Value to Variance of Costs and Graphite

Price .................................................................................................................................................... 114


Figure 16-6: Kringelgruvan RF 1.0 Whittle Pits................................................................................... 115

Figure 16-7: Inter-relationships Between Bench Geometry, Inter-ramp Slope Angle, and the Overall

Slope Angle (Wyllie & Mah, 2004) ...................................................................................................... 116

Figure 16-8: Catch Bench Geometry (originally from Call, 1986) (Kuchta & Hustrulid, 2006) ........... 117

Figure 16-9: Ovanakers Kommun Overview Map ............................................................................... 119

Figure 16-10: Final Kringelgruvan Pit Designs ................................................................................... 122

Figure 16-11: Final Pit Designs Compared with the RF 1.0 Whittle Shells ........................................ 123

Figure 16-12: Comparison of the Final Pit Design with the RF1.0 Whittle Shell (207.5 m Elevation) in

Plan ..................................................................................................................................................... 124

Figure 16-13: Comparison of the Final Pit Design on Section (as labelled) ...................................... 125

Figure 16-14: Distribution of Tonnage by Bench in the West (left) and East (right) Pit Designs ........ 126

Figure 16-15: RF 0.7 Whittle Shells Used as the Starting Point for Mining in the East and West Pits

............................................................................................................................................................ 127

Figure 16-16: Kringelgruvan Status Map – Up to End of Year 2 ........................................................ 127

Figure 16-17: Kringelgruvan 2 Year Mining Schedule ........................................................................ 128

Figure 16-18: End of Year 1 Status Map ............................................................................................ 129


Figure 16-20: Kringelgruvan Production Schedule to Year 3 ............................................................. 131


Figure 16-22: Kringelgruvan 10 Year Production Schedule ............................................................... 133

Figure 16-23: End of Year 10 Status Map .......................................................................................... 134

Figure 16-24: Kringelgruvan LOM Schedule ....................................................................................... 135

Figure 16-25: End of Mine Status Map ............................................................................................... 136

Figure 16-26: Site Overview ................................................................................................................ 136

Figure 16-27: Open Pit Drill and Blast Geometry (courtesy of DynoNobel.com) ................................ 137

Figure 16-28: Typical Blast Pattern Design ........................................................................................ 139

Figure 16-29: Woxna Graphite Mine Technical Services Department Organogram .......................... 144

Figure 16-30: Location of the Waste Storage Area ............................................................................ 148

Figure 16-31: Power Line and Pit Interaction ...................................................................................... 149


Figure 17-1: Process Block Diagram .................................................................................................. 151

Figure 18-1: Processing Plant the Woxna Project Area ..................................................................... 168

Figure 18-2: Open Pit at the Woxna Project Area .............................................................................. 168

Figure 18-3: Tailings and Water Management at Woxna ................................................................... 174

Figure 18-4: Upstream Tailings Dam .................................................................................................. 175

Figure 18-5: Clarification Pond Expansion .......................................................................................... 177

Figure 18-6: Water Conditioning Plant ................................................................................................ 177

Figure 19-1: European Average Price of Natural Graphite by Type from 2003 to 2012 (USD/t) - (2) 182

Figure 19-2: European Average Price of Medium Size Flake Graphite by Grade from 2003 to 2012

(USD/t) - (2) ......................................................................................................................................... 182

Figure 21-1: Plant Operating Cost Proportions ................................................................................... 200

Figure 21-2: Proposed Organisational Chart ...................................................................................... 202

Figure 22-1: Sensitivity Analysis ......................................................................................................... 211


NOMENCLATURE

Abbreviation/Acronym Meaning

% Percentage

° Degree

ABA Acid Base Accounting

AP Acid Potential

ARD Acid Rock Drainage

Cg Carbon graphite

CSV Comma-separated values file format

g Grams

gpm Gallon Per Minute

g/t Grams/tonne or ppm

GBM GBM Minerals Engineering Consultants Limited

Golder Golder Associates AB

GPa Gigapascal

kg Kilograms

km Kilometres

kN Kilonewton

kW Kilowatt

LCM Loose Cubic Meters

LG Lerchs-Grossmann

LOM Life-of-Mine

m Metres

M Million

m³ Cubic metres

MOU Memorandum of Understanding

MPa Megapascal

MPA Neutralization Potential

Mt Million tonnes

M tonnes Millions of Metric tonnes

NP Acid Consumption

NPR Neutralization Potential Ratio calculation

Opex Operating expenses

Pa Pascal

PEM Potentially Economic Material

t/a Tonnes per annum


Abbreviation/Acronym Meaning

p/a Per annum

h/a Hours per annum

d/a Days per annum

PEA Preliminary Economic Analysis

ppm Parts per million

RF Revenue Factor

RL Relative Level

ROM Run-of-Mine

ROMt Run-of-Mine tonne

RT90 Grid system RT 90 2.5 gon V 0:-15 (Rikets nät)

SWEREF99TM Swedish Reference frame 1999

SEK Swedish krona

Tonnes Metric tonnes

tpd Tonnes per day

USD United States Dollar

Woxna Woxna Graphite AB

WSM World Stress Map


SECTION 1 SUMMARY

1.1 BACKGROUND

The Woxna Graphite Restart Project (the Project) consists of a partially depleted graphite mine,

graphite processing plant, related infrastructure, 4 mining concessions covering graphite deposits and

numerous adjoining exploration concessions located nearby the town of Edsbyn in the Ovanåker

Municipality, Gävleborg County, in the Kingdom of Sweden. The Project is owned by Woxna Graphite

AB (Woxna), which is 100 % owned by Flinders Resources Ltd (“Flinders”).

The Woxna mine operated from 1996 to 2001, when production was halted due to falling graphite

prices. Since then the site has been held on care and maintenance.

Graphite prices have strengthened since 2009 and today are approximately double the prices when

the Woxna mine last operated. The purpose of this report is to assess the preliminary economic

potential of restarting the Project. This Preliminary Economic Assessment (“PEA”) has been

completed with an accuracy of ± 30 %.

1.2 INTRODUCTION

GBM Minerals Engineering Consultants Ltd. (“GBM”) has been contracted by Flinders to undertake a

Preliminary Economic Assessment (PEA). The purpose of this PEA is to assemble contributions from

GBM and Woxna’s other appointed contractors, to estimate the capital cost of the restarting graphite

production at Woxna, operating costs and revenue to general financial analysis to assist the Flinders

Board to consider whether to restart the Woxna mine. The PEA is to an accuracy of ± 30 %.

1.3 RELIANCE ON OTHER EXPERTS

In compiling the PEA, GBM reviewed and relied on expert’s opinions, Qualified Persons statements

and Specialist Consultants reports and for information beyond their area of technical expertise.

Information has been provided by the Client and various sub-contractors, including Amelunxen

Mineral Processing Ltd (Aminpro), Golder Associates (Golder), Reed Leyton Consulting (Reed

Leyton), Tailings Consultants Scandinavia (TCS), Flinders Resources and Woxna Graphite AB.

1.4 PROPERTY DESCRIPTION AND LOCATION

The Woxna graphite project is located at 61°24'36.49"N and 15°37'03.28"E, some 10 km WNW of the

town of Edsbyn in the Ovanåker Municipality, Gävleborg County, in the Kingdom of Sweden. It


consists of four separate issued mining concessions each securing a graphite deposit, totalling

146.71 ha.

The Woxna mine site is located on the Kringelgruvan nr 1 mining concession which is the key project

area and the subject of this report. At Woxna there is a processing plant, a tailings facility, office

infrastructure and power and water services exist next to a partially exploited open pit. The company

appears to have all the necessary rights, permits and licences validly granted to access the

properties, conduct exploration activities and mine and process graphite at Woxna.

Three other mining concessions, Gropabo, Mattsmyra and Mansberg contain historic graphite

resources only are undeveloped and are not evaluated in this report

1.5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND

PHYSIOGRAPHY

The Project is accessible via 9.5 km of unsealed all-weather forestry roads from tarred Route 301.

The climate is comparatively temperate, considering Sweden’s northern latitude. The climate is typical

of Fennoscandia with cool summers and cold winters. At Edsbyn, some 10 km to the south west of

the project area, the monthly average minimum temperature ranges from -8 °C to +11 °C and the

range of average maximum monthly temperatures is -1 °C to +23 °C. Edsbyn receives 30 mm to 70

mm of precipitation per month with autumn and winter typically drier than the spring and summer.

Mining and processing operate all year round, with possible short and minor disruptions during winter

storms as a result of snowfall and very low temperatures.

Connected grid power, water and conventional telecoms are available at the Project. Mobile

telephone services are widely available.

Local services, in terms of machine and engineering plant maintenance, are available in Edsbyn.

Road, rail and service infrastructure is well developed. Sweden has a long history of mining and local

and specialised labour is widely available.

1.6 HISTORY

The Kringelgruvan mineralization was discovered in 1986 and subsequent trenching and drilling

programmes were carried out between 1987 and 1988. Mining leases and permits were obtained

allowing mine and processing plant construction between 1993 and 1995. Production began in 1996

and continued until 2001 when the project closed down because of declining graphite prices.

Historical mineral resource estimates were available, which together with information from drilling

conducted in 2012 was used by Reed Leyton to calculate the NI 43-101 resource estimates published

in 2012 and updated in this report.


1.7 GEOLOGICAL SETTING AND MINERALIZATION

The geology of Sweden consists of three main components: Precambrian crystalline rocks, the

remnants of a younger sedimentary rock cover, and rocks of Caledonian Orogen (490 Ma – 390 Ma).

The Kringelgruvan claim shows development of trace to massive graphite in metasedimentary and

metavolcanic host rocks which have been metamorphosed to sillimanite grade and intruded by felsic

units ranging from alkali pegmatite to granite.

The geology is dominated by steeply-dipping, calcareous quartz-rich meta-tuff, with interbedded

metasedimentary units and cross-cutting pegmatite.

1.8 DEPOSIT TYPES

The dominant graphite rock association at Woxna is contact metasomatic or hydrothermal deposits in

metamorphosed calcareous sedimentary or volcaniclastic protoliths, associated with prominent

pegmatite intrusions that are interpreted to be the heat source during contact metamorphism. The

pegmatite intrusions comprise quartz, orthoclase and phlogopite and intrude a metamorphosed,

highly strained stratigraphic succession dominated by sedimentary and volcaniclastic protolithologies,

which have undergone later brittle fracturing.

The graphite deposits occur beneath a thin blanket of Quaternary age moraine deposits. The

graphite and minor associated pyrrhotite are excellent conductors that allow for prospecting using

geophysical methods.

1.9 EXPLORATION

Exploration in the early 1980s proceeded under the direction of the Swedish Geological Survey and

subsequently by MIRAB (Mineral Resources AB), a Swedish exploration company, following their

acquisition of the exploration and mining leases from the Swedish State in 1992. A variety of

techniques were used to determine the mineralogy, including magnetic, radiometric and

electromagnetic methods. Follow-up diamond drilling took place during 1988-1989 and by Flinders in

2012.

1.10 DRILLING

Between 1988 and 1989 a total of 2 908 m of diamond core was drilled on the Kringelgruvan mining

concession, comprising 51 diamond drill holes (35 mm core size). The historic program resulted in

374 graphite analyses (carbon by Leco analyser).


In 2012 drilling was conducted Flinders. A total of 41 holes comprising 3 673 m of diamond core were

drilled. The program resulted in 1 345 graphite analyses (carbon by Leco analyser).

All remnant drill core, after sampling, is stored in boxes at the Kringelgruvan project.

1.11 SAMPLE PREPARATION, ANALYSES AND SECURITY

Historical sample preparation methods and quality control measures employed before dispatch of

samples to an analytical or testing laboratory have not been documented, nor the method or process

of sample splitting and reduction, nor the security measures taken to ensure the validity and integrity

of samples taken.

At Kringelgruvan, 374 valid carbon and 52 valid sulphur analyses are presented in both paper and

database (.dbf) format. The laboratory that completed analysis of the Kringelgruvan samples was the

Government owned SGAB ANALYS, (Box 801, Luleå, Sweden 95128).

The laboratories that carried out the sampling and analytical work are independent of the Woxna and

previous project vendors. No details of certification by any standards associations and the particulars

of any certification are known, however the laboratory was well regarded and applied best practice of

the day.

1.12 DATA VERIFICATION

The adequacy, archiving and standard of the data presented is of sufficient quality for the reporting,

subject to qualification, of the historical drilling and resources as presented by previous project

owners.

Where possible, Flinders Resources Ltd has surveyed all drill collars by DGPS. The exception is the

drill collars now located in the bounds of the pit which were removed during mining. Position for these

drill holes has been calculated by converting the historic local grid into coordinates of RT90 and are

assumed accurate. The RT90 coordinates were further converted into SWEREF 99 TM by Tyréns in

January 2013.

Coffey Mining had transcribed paper records for collar, assay, survey and geology data, as reported

by Claesson et al. (1991, 1992, 1993), for Kringelgruvan and compared these to digital data available

to the issuer. Coffey Mining concluded that the historical data is of sufficient quality and traceable

provenance that it is useable as exploration data. The then supervising geologist, who is a QP under

current NI 43-101 protocol, also verified the provenance of the data supplied.


1.13 MINERAL PROCESSING AND METALLURGICAL TESTING

After reviewing historical testwork and production reports including grades and size analyses of the

graphite concentrates sold, it was decided that further test work was required to assess whether it is

possible to improve the grade and increase the flake size of the graphite products produced thereby

yielding higher prices for those products.

Outotec was commissioned to complete concentrate dewatering test work in 2012 to support the

dewatering plant design. The results indicate that thickening is not viable and that pressure filtration

yields a filter cake moisture content of 22 %.

Amelunxen Mineral Processing Ltd (Aminpro) was commissioned to complete a comprehensive test

work investigation programme with the objectives of increasing the coarse flake recovery, overall

plant recovery and product grade (1).

The test work samples for the Aminpro test work were a combination of bulk hand samples and drill

core. Intervals of drill core were selected by Monte Carlo method to represent the mineralised body.

Rougher tests were carried out on these 20 intervals, and ten of these had mineralogy carried out on

them. In addition, the work index for each sample was estimated. The remaining flotation test work

was completed with a composite sample of high and low grade hand samples with a grade of 9 % C,

and another bulk hand sample with a grade of 12 % C.

The comminution results included:

Bond Work Index of the feed was 18.5 kWh/t.

Assessment of regrinding parameters for flotation cleaner concentrate resulted in Bond

Work Index greater than 40 kWh/t (grade dependent).

Rougher flotation was tested at varying grind sizes using MIBC as the sole collector. The following

was observed:

Graphite and gangue (silicates) are the coarsest components of the material with iron

sulphides (mainly pyrrhotite) being present as fines.

Graphite floated rapidly with good recoveries (>95 %) in all tests. Gangue and iron sulphides

had lower recoveries and flotation speeds.

Flash flotation in roughers should be considered to capture as coarse a flake product as

possible and that the finer graphite would be recovered in the scavengers.

The following conclusions were reached for the cleaner flotation:

The major gangue mineral in the concentrates is muscovite

Regrinding was necessary to achieve higher grades.

Flotation at low densities was important.

Use of dispersants was seen to help depression of gangue.


Concentrate grades in the above 90 % carbon were achieved in a number of tests aimed at

developing the cleaner circuit configuration, at the optimised conditions.

A locked cycle test was performed at optimised flotation conditions. The overall graphite recovery

obtained was above 96 % C. The assay results of the tests with size are shown in Table 1-1.

Table 1-1: Results of the Locked Cycle Flotation Tests

Size fraction µm

Rougher cleaners concentrate

Rougher-scavenger cleaners concentrate

Combine concentrate

% retained % C % retained % C % retained % C

+250 13.9 95 22.0 95 18.4 95

+180 -250 18.8 97 23.4 92 21.4 94

+100 -180 26.5 94 29.6 91 28.3 92

-100 40.8 89 25.0 87 31.9 88

The locked cycle test results were used to calculate the expected product tonnages, based on mill

throughput of 155 000 t/a, as shown in Table 1-2.

Table 1-2: Expected Product Tonnages and Recovery

Graphite Product Flake size Grade (% C)

Mass Pull (%)

Product Tonnage

(t/a)

Recovery (%)

Premium Concentrate > 250 μm 95 % 18 % 2 990 18.3 %

Coarse Concentrate 180 to 250 μm 94 % 22 % 3 650 21.1 %

Medium Concentrate 100 to 180 μm 92 % 28 % 4 650 27.3 %

Fine Concentrate < 100 μm 88 % 32 % 5 310 29.4 %

Total 100 % 16 600 96.0 %

1.14 MINERAL RESOURCE ESTIMATES

Reed Leyton Consulting reviewed the historic mineral resources available and conducted necessary

inspections. The Kringelgruvan Mineral Resource describes four main bodies of mineralization

separated by faulting drilled within an area approximately 1 200 m x 100 m to 200 m.

Following an audit of historical data, Flinders data, the compiled Flinders drilling database, and the

subsequent calculation of Mineral Resources, the quoted Mineral Resources at Kringelgruvan were


subdivided into CIM-compliant measured and indicated categories on the basis of the close density of

drilling, checked grades and inter-hole continuity.

Table 1-3: Resource Grade and Tonnage (at 7% Cg cut-off grade)

Classification Tonnes (Mt) Grade (Cg)

Measured 1.0 10.7

Indicated 1.8 10.7

Total 2.8 10.7

A relative density value of 2.7 t/m3 was run according to density test work by the Woxna previously

attributed to various assays within the geology database. Two block models were constructed and

wire framing of the geological boundaries was performed using the available drill hole data. A cut-off

grade of 7 % graphite has been applied as base case to the Mineral Resource estimation.

1.15 MINERAL RESERVE ESTIMATES

There are no NI 43-101 mineral reserve estimates available to, or commissioned by, Woxna.

1.16 MINING METHODS

Golder Associates AB (Golder) undertook the mining section of the Preliminary Economic Analysis

(PEA) Woxna Graphite Restart Project in Central Sweden. Pit optimisation, pit design and mine

production schedules were performed using industry standard Gemcom Whittle™ 4x optimisation

software, Maptek Vulcan™ and Gemcom Surpac™ mine planning software. Inputs into the design

and schedule included a block model, prices and costs along with mining and geotechnical factors.

The block model and economic parameters were supplied by Woxna Graphite AB (Woxna) and GBM

Minerals Engineering Consultants Limited (GBM) respectively to Golder.

Open pit optimisation was run on Kringelgruvan deposit and for the purpose of this study the

optimisation included only graphite under Measured and Indicated mineral resources categories

within the provided resource model. The model produced a volume of Potentially Economic

Mineralisation (PEM) and waste tonnes, distributed over the life of the mine. The following table is the

summary of the optimisation.


Table 1-4: Golder Selection of Whittle Pit Shell for Kringelgruvan Deposit Ultimate Open Pit

Pit Number Revenue

Factor

Discounted Net Cashflow

(USD M)

PEM Processed

(Mt includes dilution)

Waste

(Mt)

Cut-off C

(%)

11 1.00 19.6 1.9 10.8 7.0

The RF 1.0 shell was then used to generate operational, or smooth, pit designs. Two distinct open

pits were designed around the Whittle shells. The pits included haulage ramps located on the south

wall of the pits and tied into the existing haul road network on the site.


Table 1-5: Materials Contained in Pits

Source Tonnes PEM Grade Graphite Waste Tonnes Total Tonnes Waste: PEM Ratio

East Pit 889 706 10.54 3 757 246 4 826 380 4.42

West Pit 927 780 11.40 5 690 513 6 618 293 6.13

Total Inventory 1 817 486 10.98 9 627 187 11 444 673 5.30

Figure 1-1: Pit Designs

The pits were then scheduled over the Life-of-Mine (LOM) by months for the first year, quarters for

years 2 and 3 and then annually for the remainder of the life of the project. The schedules were

charted for every operating year and reported for years 1, 2, 3, 10 and at the conclusion of mining.


Figure 1-2: LOM Production Schedule

The cumulative LOM production totals for the reporting periods are summarised in the following table.

Production of 100 000 t/a was achieved over the life of the project. The waste to PEM ratio increases

over time due to the depth of the pits and designed strip ratios until the final years when the benches

are predominantly PEM. The Graphite total in the table below is the recovered mill product from the

processing facility with a recovery rate of 85 % of the Run-of-Mine (ROM) feed grade.

Table 1-6: Production Schedule Summary

Item Units Year 1-2 Total Year 3 Total Year 10 Total LOM TOTAL

Waste tonne Tonne 798 483 1 309 319 5 558 148 9 598 843

Overburden Tonne 17 879 26 432 85 316 826 173

Strip ratio (waste:ROM)

Tonne 4.0 4.4 5.6 5.1

ROM tonne Tonne 200 000 300 000 1 000 000 1 898 739

Graphite Tonne 17 630 26 183 84 879 166 229


Figure 1-3: Final Status Map of the Site

The waste dump was designed in 10 m lifts from 235 m to 315 m and rises at a nominal rate of

approximately 5 m p/a.

The waste rock was characterised into two types; A and B. Type A was characterised as neutralising

and suitable for tailings dam and construction materials. Type B waste was determined to be acid

generating and unsuitable for construction materials. Type B waste was to be stored in an engineered

waste storage area. The study planning did not segregate these materials and assumed that all

waste materials would report to the waste dump.

The tailings reported to the existing tailings storage facility north east of the process plant and mining

areas. The mine production plan produces a total of 1.8 M tonnes of tailings solids. Golder estimated

the total water required to be an additional 50 % to the PEM processed for a total of 2.6 M tonnes of

pumped tailings over the life of the mine.

The mine operating costs were derived from contractor quotations obtained by Golder. A base rate of

48 SEK/t was used for graphite mining.

For the purposes of economic modelling, the Plant throughput is assumed to be 0.155 Mt/a ROM

tonnes, and the mine production schedule was accelerated on a pro rata basis accordingly. Checks

were completed to verify the validity of this, and other differences in final optimisation parameters and

Golder are satisfied that the PEA level mine design would not materially change. The mining study

determined that the Kringelgruvan deposit could be technically feasibly exploited over a production

period of 19 years from 2 adjacent open pit mines. Golder are satisfied that this statement remains

315 m

180 m 170 m

170 m

200 m 190 m


true under the final recovery, throughput and price parameters used in the economics for a PEA level

report.

Through the execution of the project Golder developed an array of recommendations to progress the

Woxna Graphite Project.

1.17 RECOVERY METHODS

The Woxna project has an existing graphite processing plant which was last fully operated in 2001. A

comprehensive metallurgical test work programme by Aminpro led to a new process flow sheet

designed by Aminpro and GBM. This flow sheet predicts a significant improvement in quantity, grade

and graphite flake size distribution compared to historic production. Engineering under the PEA

proposes utilising as much of the existing facilities and infrastructure as possible in line with the new

design, to minimise the initial capital cost and enable production to commence in the shortest possible

time.

The process flow sheet incorporates conventional mineral processing technology starting with 2 stage

crushing of the ROM material followed by grinding in the existing rod mill. The ground product is

treated by flash and rougher flotation to maximise the recovery of large flake. The rougher tailings are

treated in a scavenger flotation circuit to maximise recovery. The flash and rougher concentrates

then undergo 2 stages of regrinding and cleaner flotation while scavenger concentrate has 3 stages of

regrinding and cleaner flotation. Most of the flotation section uses new equipment. Graphite

concentrate is dewatered in a new filter press then dried in the existing drier. Screening and packing

of graphite product largely utilises existing equipment. A new flake product exceeding 250 micron and

94 % purity is produced in addition to the graphite grades produced historically.

The nominal throughput of the plant is 155 000 t/a based on the maximum rate that can be milled in

the existing rod mill. There is potential to significantly increase throughput by using an existing re-

grinding mill as a ball mill in the primary milling circuit along with commensurate expansions of the

flotation, drying, screening and packing sections. An expansion of capacity would be considered once

markets permit.

1.18 PROJECT INFRASTRUCTURE

The mine site has a partially depleted existing open pit, tailings management facility (TMF), waste

rock dump areas, mine site roads, clarification ponds and processing facility under care-and-

maintenance within the Kringelgruvan mining concession. Infrastructure such as power, water, road

access, telephone, mobile network coverage and internet connection are connected and satisfactory


The mine shares approximately 9.5 km of gravel forestry road, from national Route 301 to the site

gate. The mine is in a cooperative with the local community that live along the road and contribute to

its upkeep. The cost of this upkeep is included in the operating cost.

There are currently offices, an accommodation block, a milling building with change rooms and

warehouse facility on the site. Where necessary additional facilities are planned to be installed, or

upgraded such as a laboratory.

The existing power supply to the project site is from the national grid. The capacity of this connection

is limited to 4 000 KVA. No allowance has been included for upgrades to the network other than

connection to the new transformer and relocation of a section of the existing network to allow for the

proposed mine pit advancement.

The mine has an existing permitted TMF, which was constructed during the previous operation of the

mine. Future production has been estimated to generate a total of 1.7 Mt of tailings, which will require

expansion of the current TMF. The expansion is based on the upstream dam method, which will be

conducted in stages.

A water balance has been completed for the site that indicates that the site will have a positive water

balance. It is however, desirable to have some raw water for the process for which Woxna is

permitted to draw from a nearby river, though some further hydrogeological tests are recommended.

1.19 MARKET STUDIES AND CONTRACTS

The European graphite market is estimated to consume approximately 85 000 t/a of the global

demand for natural flake graphite of approximately 540 000 t/a. Today more than 90 % of Europe’s

graphite demand is imported, mainly from China.

The Woxna graphite project will primarily target the European graphite market due to its short transit

times and low transport costs. At 16 600 t/a, the design production volume in the Woxna PEA is

deliberately sized so that graphite sales may be readily absorbed into the European market without

creating an oversupply situation. Further expansion of the Woxna graphite project is possible and will

be evaluated when European market conditions permit.

Graphite prices have risen considerably since the Woxna mine was last operated in 2001. In 2010,

following a long period of flat prices, graphite prices, particularly for flake grades, rose rapidly as a

result of a surge in demand due to restocking, growth from China and lithium batteries.

Simultaneously, the imposition of taxes and permits on Chinese graphite exports restricted supply.

Graphite prices peaked in 2012 and eased to today be approximately double the level seen in 2001.

Graphite market commentators predict that prices may be approaching the bottom and with signs of

economic recovery in the USA and the worst of the recession behind Europe, graphite prices may

again strengthen.


Graphite is not an openly traded mineral with prices negotiated privately between customers and

producers under spot or term contracts. The sale price used in the PEA was based on graphite prices

published by Industrial Minerals magazine (‘’IM’’).

It has been common practice to use a 24 month trailing average graphite price in graphite PEAs. The

Woxna PEA utilized a 12 month trailing average graphite price so as to exclude the peak price period

occurring in 2011/12. Applying Woxna’s planned product distribution to IM’s 12 month trailing average

graphite prices produces an average selling price of 1 199 USD/t for the Woxna PEA. This selling

price is considered to be conservative when compared to the 24 month trailing average graphite price

of 1 548 USD/t or prices used in other graphite project PEAs.

Table 1-7: Average Sale Price Estimation

Size (µm) Purity Proportion of Production

1 Year Trailing Average (USD/t)

Annual Quantity (t)

+ 250 95 % 18 % 1 824 2 990

+ 180, - 250 94 % 22 % 1 526 3 650

+ 100, - 180 92 % 28 % 1 009 4 650

- 100 88 % 32 % 787 5 310

Average / Total 1 199 16 600

1.20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

Swedish legislation requires the undertaking of an environmental (and social) assessment (in

Swedish “MKB”) at two different stages during the development of a mining project in order to obtain

the necessary permits. MKBs must be performed according to the requirements of the Swedish

Environmental Code (1998:808).

The first MKB is produced when applying for an exploitation (mining) concession in accordance with

the Minerals Act (SFS 1194:45) from the Mining Inspectorate of Sweden (“Bergsstaten”). The second

MKB is produced when applying for environmental permits.

Flinders already has existing valid exploitation and environmental permits for the Woxna project.

According to the Swedish legislation the mine is still in operation, even though the company has not

actively been mining since 2001. All the documentation (Environmental Impact Assessment (“EIA”)

etc.) is in Swedish and mostly originates from the permitting process in 1992.

In order to complete any expansions or significant variations in processing methodologies, Woxna

may need to apply for a new environmental permit. The operation is fully permitted for processing of

0.1 Mt/a of graphite mineralised rock, however the proposed design and economics of the project are


based on 0.155 Mt/a. Preliminary steps in permit extension are underway and it is anticipated that the

permit extension will be completed by mid to late 2015; before the annual cumulative throughput of

the plant is expected to exceed 0.1 Mt. Short term permit extensions may also be negotiated.

1.21 CAPITAL AND OPERATING COSTS

1.21.1 CAPITAL COSTS

The capital cost estimate has been prepared to the level of a preliminary study, with an accuracy of

-20 % to +30 %, and is presented in USD.

The total initial capital investment for the start-up is USD 16.7 M, which includes contingency and

working capital.

Table 1-8: Initial Capital Cost (M USD)

Cost Centre Total 000

General

100

Mining

200

Process

300

Waste

400

Product

500

Infrast

Total Capital Investment 16.72 5.69 0.15 6.80 3.37 - 0.70

Fixed Capital 14.34 3.75 0.15 6.36 3.37 - 0.70

100 - Direct 10.28 0.31 - 6.36 2.91 - 0.70

200 - Indirect 4.06 3.44 0.15 - 0.46 - -

Working Capital 2.38 1.94 - 0.44 - - -

1.21.2 OPERATING COSTS

An operating cost estimate has been developed for the normal operation of the mine, plant and

infrastructure. The annual cost is

This OPEX Breakdown for the mine and processing plant and infrastructure is shown in Table 1-9,

and are based on LOM average production figures.

Table 1-9: Operating Cost Summary

Area USD/t ROM USD/t graphite

Mining 25.8 240

Processing and Infrastructure 45.3 422

Cost of Sales 7.4 68

Total 78.5 730


1.22 ECONOMIC ANALYSIS

A financial analysis has been conducted to demonstrate the economic performance of the project.

The capital costs, operating costs and mine plan as discussed have been used for this analysis. The

following summarises the IRR and NPV yields of the base case analysis.

Table 1-10 - Summary Economic Analysis

Parameter Phase 1 Unit

LOM 13.0 years

Average Sale Price 1 199 USD/t graphite

Revenue 121.5 USD/t ROM

1 130 USD/t graphite

OPEX 71.1 USD/t ROM

662 USD/t graphite

Cost of Sales 7.4 USD/t ROM

68 USD/t graphite

Margin 50.4 USD/t ROM

468.6 USD/t graphite

Payback Period 3.86 years

IRR 34.0 %

NPV 26.6 M USD

1.23 ADJACENT PROPERTIES

There are no known operators of any relevant activities on directly adjacent properties or locally

adjacent properties.

1.24 INTERPRETATIONS AND CONCLUSIONS

A measured and indicated mineral resource of 2.8 Mt at an average grade of 10.7 % has been

estimated in accordance with NI 43-101 guidelines. Potential for increasing of the mineral resources

are good, with mineralization open down dip and along strike at the Kringel concession and with 3

other historic graphite resources available. This, which requires further drilling to investigate evaluate

its potential.

Golders have developed a mine plan which can be accelerated to deliver 155 000 t/a of potentially

mineralised material for processing.


New metallurgical test work has yielded enhanced flake graphite grades, recoveries and flake sizes

leading to a new process flow sheet to be developed. A combination of refurbished existing and new

processing equipment was specified.

From the process flow sheet and equipment list capital and operating costs have been developed.

A conservative graphite sales price of 1 199 USD/t was estimated based on 1 year trailing average

graphite prices and Woxna’s proposed product volumes.

Based on the production schedule, predicted metallurgical recoveries, estimated costs and base case

price, the Woxna project generates a positive operating margin and cash flows producing an NPV of

USD 26.6 M (post tax, 8% discount rate) and internal rate of return of 34.0 %.

There have been a number of risks and opportunities identified for the project.

Preliminary mining design positively or negatively predicting realistic sinking rate, dilution,

waste dump design, acid rock potential, stability and variations in pit optimisation parameters.

The conversion of metallurgical results to process design

Use of refurbished equipment relating to plant availability and potential for further expansion

Uncertainties relating to PEA level assumptions

1.25 RECOMMENDATIONS

The Woxna project PEA has produced a positive economic outcome and further development of the

project is recommended including:

Drilling of the Kringel deposit to increase the confidence of the resource and provide

geotechnical data for the mine plan

Development of the mine plan to generate mineral reserves

More detailed planning of the waste rock dumps including storage of potentially acid

generating rock

Detailed engineering design of the process plant and tailings facility


SECTION 2 INTRODUCTION

2.1 CLIENT

This report has been written for Flinders Resources Ltd (“Flinders”), a Toronto Venture Exchange

listed resource company, who is primarily focussed on advancing the Woxna Graphite mine in

Sweden.

2.2 TERMS OF REFERENCE AND PURPOSE

The Woxna Graphite Restart Project (the Project) consists of a partially depleted graphite mine,

graphite processing plant, related infrastructure, 4 mining concessions covering graphite deposits and

numerous adjoining exploration concessions located nearby the town of Edsbyn in the Ovanåker

Municipality, Gävleborg County, in the Kingdom of Sweden. The Project is owned by Woxna Graphite

AB (Woxna), which is 100 % owned by Flinders Resources Ltd (“Flinders”).

The Woxna mine operated from 1996 to 2001, when production was halted due to falling graphite

prices. Since then the site has been held on care and maintenance.

Historically the plant had a rated capacity exceeding 10 000 t/a of flake graphite limited by the

environmental permit condition restricting mining to 100 000 t/a of ore. Historical production achieved

graphite products ranging in purity between 85 % to 94 % Carbon (C). The Project aims to define and

evaluate the requirements to refurbish and restart the Woxna mine including implementing process

improvements.

GBM Minerals Engineering Consultants Ltd. (“GBM”) has been contracted by Flinders to undertake a

Preliminary Economic Assessment (PEA). The purpose of this PEA is to assemble contributions from

GBM, Woxna’s other appointed consultants and experts as outlined in Section 3 to define the

operational configuration, estimate the operating and capital costs of restarting the mine and produce

a financial evaluation of the Project. The PEA is to an accuracy of ± 30 %.

2.3 SOURCES OF INFORMATION

GBM has been provided with various existing project information by Flinders, refer to

Section 27 - References for a complete list. In addition to the published information GBM has visited

the site and conducted an audit of the plant and infrastructure. Other relevant consultants have also

visited the operation to verify information for their relevant scopes of work.

The information supplied has been sufficient to allow this PEA to be completed to the required level of

detail and accuracy.


SECTION 3 RELIANCE ON OTHER EXPERTS

3.1 QUALIFIED PERSONS

The Qualified Person(s) are:

Mr Christopher Stinton, BSc (Hons) (Minerals Engineering), CEng, MIMMM (QP), Senior Process

Engineer with GBM Minerals Engineering Consultants of the United Kingdom. Chris is an independent

consultant to the Company, and takes responsibility for Sections 1, 2, 3, 13, 17, 18, 19, 20, 21, 22, 23,

24, 25, 26 and 27 of this report.

Mr. Geoff Reed, B App Sc, MAusIMM (QP), Principal of Reed Leyton Consulting Ltd. Geoff is an

independent consultant to the Company, and takes responsibility for Sections 4, 5, 6, 7, 8, 9, 10, 11,

12, 14, 25.1 and 26.1 of this report.

Mr. Bryan Pullman, B.Sc. Eng., P.Eng. (Alberta) (QP), is a Senior Mining Engineer with Golder

Associates (UK) Ltd. Bryan is an independent consultant to the Company, and takes responsibility for

Sections 15, 16, 25.2 and 26.3 of this report.

Mr. Henning Holmström, M.Sc., Ph.D, MAusIMM, MAIG (QP), Director, Woxna Graphite AB Henning

is a non-independent to the Company, and has co-authored, together .with Christopher Stinton,

Sections 18.3 to 18.6, 25.5 and 26.6 of this report.

3.2 RELIANCE ON OTHER EXPERTS

In compiling the PEA, GBM reviewed and relied on expert opinions and information from Flinders and

Woxna Graphite on relevant legal, political environmental and tax matters. The extent of reliance,

source and effected sections are as follows.

All information on the mining legality and land ownership was provided by Mikael Ranggård,

Woxna Graphite via email exchange and conversation during late 2012. This information has

strongly influenced Section 4.

All information on environmental matters (including permitting) was provided by Henning

Holmström, Woxna Graphite via email exchange and conversation during 2012 and 2013.

This information has strongly influenced Section 20.

All information relating to Flinders tax assets (loss and depreciation carry forward) and

accounting of deprecation in Sweden was provided by Mats Lindén, accountant to Woxna

Graphite via email and conversation in August 2013. This information has been included in

Section 22.

Market trends and verification of off-market contract pricing ranges was provided by Mikael Ranggård,

lawyer of 20 years and Chairman of Woxna Graphite and Martin McFarlane, chemical engineer with


25 years of project development and former CEO of Flinders Resources, in support of the Industrial

Minerals database prices used (accessed under subscription) via email exchange and conversation

during 2012 and 2013. GBM have relied on the information from Mr McFarlane and Mr Ranggård

under knowledge that they have conducted discussions with potential off-takers and have monitored

the graphite market as part of their employment by Woxna and Flinders. This information has strongly

influenced Section 19. This information validated the database pricing obtained and as such no

further validation was deemed required.

3.3 VERIFICATION

GBM personnel have inspected the project site and verified its characteristics as stated in respective

sections of this report.

The authors have carried out due diligence reviews of the information provided to them by the Client

and others for the preparation of this report and are satisfied that the information was accurate at the

time of the report and that the interpretations and opinions expressed in them were reasonable and

based on current understanding of mining and processing techniques and costs, economics,

mineralisation processes and the host geologic setting. The authors have made reasonable efforts to

verify the accuracy of the data relied on in this report.

3.4 FINANCIAL INTEREST DISCLAIMER

Neither GBM nor any of the consultants employed in the preparation of this report have any beneficial

interest in the assets of Woxna Graphite AB or Flinders Resources Ltd.

As outlined, some data has been provided by Woxna Graphite and its employees. This work has been

reviewed and GBM has no reason to believe that pecuniary interest has influenced the validity of the

information presented.

Contributing consultants have been paid fees and will continue to be paid fees for this work in

accordance with normal professional consulting practices.


SECTION 4 PROPERTY DESCRIPTION AND LOCATION

4.1 LOCATION

The Woxna graphite project is located at 61º 26´ 08.94˝N and 15º 36´ 16.31˝E, some 8 km WNW of

the town of Edsbyn (Figure 4-1) in the Ovanåker Municipality, Gävleborg County, in the Kingdom of

Sweden.

Figure 4-1: Location of Flinders Resources Graphite Projects, Sweden


Table 4-1: Woxna Graphite Project Tenure

Property

Type of Area Valid Until Extraction permit

Conditions Mineral Tenure

(ha) (date) (Environmental

permit)

Kringelgruvan Mining Lease

30.77 31/12/2016

Permits received and still valid 1992-09-17

and 1992-10-27

Environmental permit valid for 100 kt annual ore processing.

nr 1 (Exploitation concession)

Mining has depleted 300kt from historical resource.

*

**

Gropabo Mining Lease

18.20 21/02/2025

Permit received 2005-03-21

*** (Exploitation concession)

Expired 2012-04-18

Mattsmyra Mining Lease

72.97 21/02/2025


*** (Exploitation concession)

Expired 2012-04-18

Månsberg

Mining Lease

24.77 27/12/2024 No application

filed (Exploitation

concession)

* Is automatically extended 10 years when a mine is in operation

** The central graphite deposit incl. mine, mill, tailings impoundment, clarification pond and the rest of the facilities.

*** Surrounding graphite deposit (no mine).

Woxna Graphite AB, Flinders’ 100 % owned Swedish subsidiary, owns 4 mining concessions over

graphite deposits (Kringel, Gropabo, Mattsmyra and Mansberg – the Woxna Project) located along a

40km trend in central Sweden. The PEA considers only one of these deposits, the Kringel deposit.

The partially mined Kringel deposit lies adjacent to the processing plant, tailings dam and related

infrastructure, and is fully permitted for an immediate restart. Woxna is currently reprocessing

graphite and last mined graphite in 2001.

The Gropabo, Mattsmyra and Mansberg concession contain historic flake graphite resources. These

will continue to be classified as historic resources until Flinders has the opportunity to upgrade them

to NI43-101 standards. These historic resources are not included in the economic analysis of the

PEA.

4.2 MINING LAWS AND REGULATIONS

Swedish mining laws changed profoundly in 1992 when the new Minerals Act of 1991 (effective 1 July

1992) for the first time allowed foreign ownership of mineral titles in Sweden. The right of the


Swedish state to acquire 50 % of a mine was repealed a year later. Exploration permits and mining

licences approved before 1 July 1992 are governed by the Minerals Act of 1974 that does not permit

foreign ownership of mineral title or surface rights.

Rules and regulations pertaining to mining exploration in Sweden are clearly outlined in the “Guide to

Mineral Legislation and Regulations in Sweden” (2000) available from the offices or the website of the

Geological Survey (www.sgu.se). The Mining Inspectorate of Sweden provides clear directives,

available from the Inspectorate website (www.bergsstaten.se), for conducting mining and exploration.

Flinders has, or will address, all requirements before undertaking any work on its concessions. The

company has the rights to access the properties, and no restrictions or limitations as defined for work

on the projects are evident. The company has the obligation to outline a work program and gain

permission from landholders prior to accessing the properties, and to provide compensation for any

ground-disturbing work conducted.

The information in the following subsections was provided from the website of the Mining Inspectorate

of Sweden (Bergsstaten), being the agency responsible for the administration of mineral resources in

Sweden.

4.2.1 MINING INSPECTORATE

The Bergsstaten is managed under the Ministry of Industry, Employment and Communications, and

reports to and receives administrative and other support from the Geological Survey of Sweden

(SGU). The director of the Inspectorate is the Chief Mining Inspector, appointed by the Government.

The functions of the Inspectorate are to issue permits under the Minerals Act (1991:45) for the

exploration and exploitation of mineral deposits and to ensure compliance with the Act.

The Mining Inspectorate became a single authority on 1 July 1998, when an earlier subdivision into

districts (mining inspectors' offices) was abolished by a parliamentary decision. The Inspectorate now

has offices in Luleå (the Head Office) and Falun. The Mining Inspectorate was established as a state

authority in 1637.

4.2.2 LEGISLATION ON MINERALS

The Minerals Act (1991:45) came into force on 1 July 1992. It has subsequently been amended as

follows:

1 July 1993, abolition of the rules giving the state a half share in mines (1993:690),

1 July 1998, introduction of protection zone rules for mines (1998:165),

1 January 1999, adapted to the new Environmental Code (1998:808), which entered into

force on the same date.

The other principal acts and ordinances governing the exploitation of minerals are:


Minerals Ordinance (1992:285),

The Act on the Continental Shelf (1966:314),

The Continental Shelf Ordinance (1966:315),

The Certain Peat Deposits Act (1985:620),

The Certain Peat Deposits Ordinance (1985:626).

Over the last century, Sweden has had the following laws relating to minerals:

The 1884 Mining Regulation (1884:24), which was replaced by

The 1938 Mining Act (1938:314), which was in turn superseded by

The 1974 Mining Act (1974:342).

The 1886 Coal Deposits Act (1886:46) and

The 1960 Graphite Act (1960:679). These were both replaced by

The Act concerning Certain Mineral Deposits (1974:890).

The Minerals Act currently in force replaced both the 1974 Mining Act and the Act concerning Certain

Mineral Deposits of the same year. Depending on the type of land affected and work to be carried

out, there are varying requirements for official approvals.

The following table reflects the normal steps to be followed and approvals gained from exploration

through to final approval of mining in Sweden.

Approval Required Authority

1. Exploration permit (undersökningstillstånd) (survey of the bedrock)

Mining Inspector

2. Exploration work (undersökningsarbete) (when the environment or land use is affected)

County Administrative Board etc; Landowner

3. Exploitation concession (bearbetningskoncession) (with environmental impact assessment and approval under chapters 3–4 of the Environmental Code)

Mining Inspector; County Administrative Board etc or Government in case of disagreement)

4. Permission under the Environmental Code (Chapter 9 of the Code)

Environmental Court

5. Designation of land (markanvisning) Landowner; Mining Inspector

6. Building permit etc. under the Planning and Building Act Local authority

4.3 PERMITS AND CONCESSIONS

4.3.1 EXPLORATION PERMITS

The Minerals Act relates to the exploration and exploitation of certain mineral deposits on land,

regardless of the ownership of the land. Applications for permits etc. are made to the Mining


Inspectorate (Bergsstaten). The Act defines to which mineral substances its provisions apply; these

are known as concession minerals. Concession minerals are divided into three categories, being

traditional ores, certain industrial minerals, and finally oil, gas and diamonds. Other minerals and

other kinds of rock, gravel and sand are excluded from the Act and are normally referred to as

landowner minerals.

An exploration permit (undersökningstillstånd) gives access to the land and an exclusive right to

explore within the permit area. It does not entitle the holder to undertake exploration work in

contravention of any environmental regulations that apply to the area. Applications for exemptions

are normally made to the County Administrative Board.

An exploration permit is granted for a specific area where a successful discovery is likely to be made.

It should be of a suitable shape and size and no larger than may be expected to be explored by the

permit holder in an appropriate manner. A permit is to be granted if there is reason to assume that

exploration in the area may lead to the discovery of a concession mineral.

An exploration permit is initially valid for a period of three years, after which it can be extended up to a

total of 15 years if special conditions are met.

Compensation must be paid by the permit holder for damage or encroachment caused by exploration

work.

When an exploration permit expires without an exploitation concession being granted, the results of

the exploration work undertaken must be reported to the Mining Inspector.

4.3.2 EXPLOITATION CONCESSIONS

An exploitation concession (bearbetningskoncession) gives the holder the right to exploit a proven,

extractable mineral deposit for a period of 25 years, which may be extended. Permits and

concessions under the Minerals Act may be transferred with the permission of the Mining Inspector.

An exploitation concession relates to a distinct area, designated on the basis of the location and

extent of a proven mineral deposit. A concession may be granted when a mineral deposit is

discovered which is probably technically and economically recoverable during the period of the

concession, and if the nature and position of the deposit does not make it inappropriate to grant a

concession. Special provisions apply to concessions relating to oil and gaseous hydrocarbons.

Under the provisions of the Environmental Code, an application for an exploitation concession is to be

accompanied by an environmental impact assessment. Applications are considered in consultation

with the County Administrative Board, taking into account whether the site is acceptable from an

environmental point of view.


4.3.3 PERMIT FROM THE ENVIRONMENTAL COURT

Under the rules of the Environmental Code, a special environmental impact assessment for the

mining operation must always be submitted to the Environmental Court, which examines the impact of

the operation on the environment in a broad sense. The Court also stipulates the conditions which the

operation is to meet.

4.4 ACQUISITION OF LAND

Land needed for exploitation is normally acquired by the mining company through contracts of sale or

leases. If there is a contract of sale, a property registration procedure must generally be undertaken

through the Land Survey authority in order for registration of title to be granted.

Before any land, inside or outside the concession area, may be used it has to be designated by the

Mining Inspector (markanvisning). This procedure usually regulates the compensation etc. to be paid

to affected landowners, normally on the basis of an agreement between the company and the

landowners, together with any other parties whose rights may be affected.

4.5 TAXES AND DUTIES

Mining companies (limited companies) pay corporation tax at a rate of 22 % under the same rules as

every other company. Accordingly, there are no special taxation rules for such companies.

For permits granted after 2005, a royalty is paid on the value of minerals produced at a rate of 0.2 %,

which is shared between the landholder and the State each receiving 0.15 % and 0.05 % respectively.

As Woxna’s exploitation concessions were granted prior to 2005, this royalty does not apply to

Kringelgruvan nr1 mining licence.

The application fee for an exploration permit is SEK 500 for each area of 2 000 ha or part thereof. The

exploration fee varies for different concession minerals and for different periods of validity.

The application fee for an exploitation concession is SEK 6 000 per area.

4.6 ENVIRONMENTAL CONSIDERATIONS

There are no current outstanding environmental liabilities on any of the licenses and, as required by

Swedish law, all landowners identified by Flinders have been informed by the Swedish Inspectorate of

Mines (Bergsstaten) that an exploration license has been applied for in accordance with

Chapters 1.1 and 2 of the Mineral Act.

No environmental or planning permitting is required for geological mapping and minor, scattered hand

till sampling. Permits are required however from the district authorities for systematic till sampling,


trenching and drilling programs. A small environmental bond has been paid for the exploration

license and for the new mining license about SEK 30 000 bond is required before any drilling may

commence.

4.7 FLINDERS’ PROPERTY LOCATIONS

As indicated above, Flinders holds 4 mining leases in central Sweden. Woxna mine site is located on

Kringelgruvan nr 1 mining lease which is the key project area and the subject of this report.

4.7.1 KRINGELGRUVAN CLAIM

A 500 000 SEK (Swedish kronor) security has been paid against the Mining Lease. The project co-

ordinates are defined in the Mining Inspector’s (Bergsstaten) decision DNR 320-71X-1991 with the

following cornice co-ordinates (Table 4-2) in Swedish Reference frame 1999 (SWEREF99TM). Note

that in Swedish convention x and y are interchanged from normal usage elsewhere.

Table 4-2: Woxna Graphite Project – Kringelgruvan nr 1 Mining Concession Co-ordinates

Point Northing (metres) Easting (metres)

1 6 808 741.31 533 122.19

2 6 808 436.43 533 123.96

3 6 808 433.90 532 191.28

4 6 808 477.36 531 986.81

5 6 808 493.44 531 912.64

6 6 808 601.71 531 937.29

7 6 808 628.45 531 997.94

8 6 808 632.78 532 024.88

9 6 808 724.75 532 590.56


SECTION 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESS

The project is accessible from the tarred east west Route 301, which is an all-weather road. Local

access to the Kringelgruvan Mining Lease is on unsealed all-weather forestry roads.

The operating season is all year round, with possible short and minor disruptions at the height of

winter with snowfall and very low temperatures.

5.2 PHYSIOGRAPHY

The landscape was sculpted by extensive glaciers to form shallow lakes and extensive boggy

lowlands during the most recent ice age, spanning a period between three and ten thousand years

ago. Broad valleys were scoured out in the direction of glacial transport flanking low-lying hills

underlain by resistant rocks. The landscape of Sweden is dominated by low rolling hills (70 %) and flat

lowlands (30 %) comprised of bogs and lakes. Hills are mostly covered by glacial moraine and sands

and forested, primarily with birch and pine.

The project elevation ranges from approximately 220 m to 280 m in relief and comprises NW-SE

orientated low hills with trellised local stream drainage and numerous fresh water lakes, of which the

Råttjärnasjön and Loftssjön are the largest. These, in the main, ultimately flow to the Woxnan River

which is an incised meandering river to the south of the Kringelgruvan mineral claim. The Woxnan

River is the source of hydro-power in the district.

The graphite mineralization of Kringelgruvan is located to the southern slope of Gräsberget.

5.3 CLIMATE

The climate is comparatively temperate, considering the Project’s northern latitude. The climate is

typical of Fennoscandia with cool summers and cold winters. The principal moderating influences are

the Gulf Stream and the prevailing westerly winds, which blow in from the relatively warm Atlantic

Ocean. In winter these influences are offset by cold air masses that periodically sweep in from the

east.

At Edsbyn, some 10 km to the south west of the project area, the monthly average minimum

temperature ranges from -8 °C to +11 °C and the range of average maximum monthly temperatures is

-1 °C to +23 °C. Edsbyn receives 30 mm to 70 mm of precipitation per month with fall and winter

typically drier and spring and summer typically wetter periods.


5.4 WATER, POWER AND TELECOMS

Connected grid power, water and telecoms including internet are available at the Kringelgruvan

processing plant and open pit. Mobile telephone services are widely available.

5.5 LOCAL RESOURCES

Local services, in terms of machine and engineering plant maintenance, are available in Edsbyn.

Road, rail and service infrastructure is well developed. Sweden has a long history of mining and local

and specialised labour is widely available.

The local economy has been focussed on forestry and plantation cropping since the 1890s. This is

now a largely mechanised enterprise and uses similar types of machinery used in mining operations.

There are some localised seasonal pasture and very minor cropping. The nearby Woxnan River is

used for tourism and recreation. The local population numbers approximately 11 000 people.

All social and industrial needs and services such as accommodation, provisions, supplies,

communications etc. are readily and commercially available. They are of high standard, typical of the

modern industrial economy that is Sweden. The national power grid extends throughout the region.

Water resources are plentiful.

Vegetation is a mixture of pine and fern forest, with some localised bogs in low-lying areas. The

Kringelgruvan lease, pit, processing plant and tailings facility are largely cleared of local vegetation.

The area has stands of commercial timber. The surface freehold owners are mostly Swedish and

international forestry companies. The extent and location of these surface holders in the vicinity of

the Kringelgruvan Mining Lease are known to the Company and verified by Reed Leyton.


Figure 5-1: Topography and Access of the Kringelgruvan Project Area


SECTION 6 HISTORY

The initial discovery of graphite at Woxna was made in 1983 by a prospector engaged by the Swedish

Geological Survey (SGU) as part of a regional mapping program tracking large boulders in

Quaternary age moraine. The original surveys were directed at uranium exploration using airborne

radiometric data. The SGU and its agencies followed the discovery with both regional airborne and

local ground-based surveys, including electromagnetic (Slingram) VLF, and magnetometer methods,

to delineate geophysical conductors under the thin cap of recent till.

EM methods have proven to be very efficient delineators of conductive zones containing schlieren

and blebs of coarse graphite, as well as associated zones with 1 % to 5 % pyrrhotite. In 1988-1989

sufficient drilling was undertaken at Kringelgruvan and the other 3 mining licences to delineate the

historic resources.

In 1993, the concessions passed to Woxna Graphite AB, a small Swedish-based company

specialising in the development of industrial mineral properties.

The historic field work, chemical analysis and historic resource estimates discussed in the Reed

Leyton Report (2012) were conducted by previous private organizations. Data generated in 2012 was

under the guidance of the staff and contractors of Flinders Resources Ltd.

6.1 HISTORICAL DEPOSIT OWNERSHIP AND EXPLORATION

Kringelgruvan mineralization was discovered in 1986 by Slingram measurements and subsequent

trenching. More trenching followed in 1987 and the first drilling commenced in 1988. A 2nd drill

phase was undertaken the following year. Additional ground geophysics was conducted in 1989 to

cover a larger area. All historic exploration work was completed by the precursors of today’s Swedish

Geological Survey; namely SGAB and NSG (Sveriges Geologiska AB and Nämnden för statens

gruvegendom respectively). In 1992 the concession passed to Mineral Resources AB (“MIRAB”) a

Swedish private company who later sold it to Tricorona AB.

Tricorona brought the project into production in 1996. The mine was in production until 2001 when it

closed down due to declining graphite prices.

Table 6-1: Drilling History of the Kringelgruvan Project

Hole Type Year Hole

Number Metres Tenement

DD 1988 28 1 595 Kringelgruvan



6.2 SAMPLING AND CHEMICAL ANALYSIS

Original assay results exist in a database compiled by the Swedish Geological Survey which was

taken over by Woxna Graphite AB in 1992. Analysis was made by the Leco furnace method by

SGABs laboratory in Luleå. No information is currently available as to specific quality assurance,

quality control (QA/QC) protocols used by the SGU in its drilling or analytical program, however work

completed by the SGU is routinely of a high standard.

6.3 DENSITY DETERMINATION

No bulk density determinations conducted on mineralized intervals have been viewed by Reed

Leyton.

6.4 PETROLOGY

In total 107 thin sections were prepared from core between 1988 and 1989 (as reported in

publications coded PRAP_88537 and PRAP 88532). The studies mainly focused on the graphite

flake size and associated mineralogical distribution of the mineralization.

6.5 HISTORICAL MINERAL RESOURCE ESTIMATES

The Kringelgruvan deposit has an historical estimate of Mineral Resources or Mineral Reserves,

which have previously been disclosed in accordance with paragraph 2.4(a) of the Instrument.

Table 6-2: Woxna Graphite Project Historical Resources (after Claesson, 2002)

Tenement Classification** Tonnes Grade C Cut-off

Date Notes (Mt) (%) (%)

Kringelgruvan Measured 1.11 11.3 7 1988 2

Kringelgruvan Indicated 0.22 11.4 7 1988 1

Total 1.33 11.3 7 1988 3

** Foreign resource as provided for in Part 2 of the Companion Policy to NI 43-101

The Company chose to disclose historical estimates for each of the Woxna tenements under Part 2,

2.4 Disclosure of Historical estimates. The estimates (Table 6-2) were performed by Dr L-A

Claesson, Eur Geol. and a Qualified Person as contemplated by the NI 43-101 instrument. The

estimate has been classified according to the principles of JORC and NI 43-101 by the Qualified

Person. The estimates are cross-sectional polygonal interpretations using a simple nearest-


neighbour, sectional approach and using a density of 2.7 g/cm³ to convert volume to tonnage.

Density assumptions are supported by test work.

Data used in calculating these Historic Mineral Resource Estimates is historical in nature and was

compiled prior to the implementation of NI 43-101 reporting standards. Reed Leyton has not

completed sufficient exploration to verify the estimates. Reed Leyton is not treating them as National

Instrument defined resources or reserves verified by a Qualified Person, and the historical estimate

should not be relied upon.


SECTION 7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 REGIONAL GEOLOGY

The geology of Sweden consists of three main components: Precambrian crystalline rocks, the

remnants of a younger sedimentary rock cover, and rocks of Caledonian Orogen (490 Ma – 390 Ma).

The Precambrian rocks are part of a stable area known as the Baltic or Fennoscandian Shield).

These consist of rocks formed during the Precambrian period, i.e., at some time between the

formation of the Earth about 4.6 billion years ago and the start of the Cambrian period about 545 M

years ago.

The oldest rocks preserved in Sweden are of Archaean age (>2.5 billion years old). Archaean rocks,

however, only occur to a limited extent in the northernmost part of the country. The rocks in the rest

of northern Sweden and in the eastern and southern parts of the country are generally between

2.0 billion years old and 1.65 billion years old. They formed, and were in many cases also

metamorphosed, during the Sveco-Karelian Orogeny, which also affected the older Archaean rocks.

Bedrock in southwestern Sweden is mainly between 1.7 billion years old and 1.55 billion years old. It

was metamorphosed during the Sveco-Norwegian Orogeny, which occurred about 1 100 M years ago

to 900 M years ago. In the south, bedrock was also metamorphosed at an intermediate stage,

between 1 450 M years and 1 400 M years.

Phanerozoic sedimentary rocks rest unconformably on the Precambrian shield area. They are less

than 545 M years old and cover large parts of Skåne, the islands of Öland and Gotland, the Östgöta

and Närke plains, the Västgöta mountains, the area around Lake Siljan in Dalarna and areas along

the Caledonian Orogen in northern Sweden.

The youngest rocks in Sweden are Tertiary age rocks, formed circa 55 M years ago. These occur in

the most southerly and southwestern parts of Skåne.

The Caledonian Orogeny is the youngest deformation event in Sweden, dated at

490 M years - 390 M years. The rocks of this remnant mountain chain vary in age from Precambrian

to early Devonian, i.e., >390 M years old.


Figure 7-1: Regional Geology of the Kringelgruvan Project Area


7.2 LOCAL GEOLOGY

The Kringelgruvan claim shows development of trace to massive graphite in metasedimentary and

metavolcanic host rocks which have been metamorphosed to sillimanite grade and intruded by felsic

units ranging from alkali pegmatite to granite. Kringelgruvan has variable cover of 2 m to 15 m of

Quaternary age moraine.

At Kringelgruvan, the geology is dominated by steeply-dipping, calcareous quartz-rich meta-tuff, with

interbedded metasedimentary units and cross-cutting pegmatite. Two discrete tabular zones of

graphite mineralisation are developed and trace pyrrhotite is associated with the mineralised zone, its

foot wall and hanging wall. The mineral assemblage includes accessory prehnite and zoisite and the

ubiquitous quartz-feldspar-chlorite-sericite assemblage indicating a lower grade of metamorphism.

The mineralisation is tabular in shape, and late in the structural history, postdating and cross-cutting

any remnant tectonised and metamorphosed lithologies.

Figure 7-2: Local Geology of the Kringelgruvan Project Area


SECTION 8 DEPOSIT TYPES

Graphite is developed as an accessory mineral as laminated aggregates dispersed through schistose

and siliceous metamorphic rocks. Graphite is an opaque mineral with six-sided form and crystallises

in the hexagonal system with rhombohedral symmetry. It has a perfect basal cleavage and thus

presents as flat flakes. These have a metallic lustre. Graphite is found as both flakes (>70 µm) and a

finer-grained amorphous, microcrystalline type. Graphite has a dark streak and is visually obvious in

core.

Graphite occurs mainly in five rock associations (Taylor, 2006) and these are:

Amorphous deposits formed by the thermal metamorphism of coal or carbon-rich

sedimentary rocks;

Disseminated in marble - metamorphosed dolomite or a calcareous protolith;

Veins filling fractures fissures and cavities in country rock;

Disseminated in metamorphosed silica-rich metasedimentary rocks such as quartzites;

Contact metasomatic or hydrothermal deposits in metamorphosed calcareous sedimentary

or volcaniclastic protoliths.

At Woxna, the lattermost is the dominant type, associated with prominent pegmatite intrusions that

are interpreted to be the heat source during contact metamorphism. The pegmatite intrusions

comprise quartz, orthoclase and phlogopite and intrude a metamorphosed, highly strained

stratigraphic succession dominated by sedimentary and volcaniclastic protolithologies, which have

undergone later brittle fracturing.

The graphite deposits occur beneath a thin blanket of Quaternary age moraine deposits. The

graphite and minor associated pyrrhotite are excellent conductors that allow for prospecting using

geophysical methods.


Figure 8-1: Graphite Mineralisation of the Kringelgruvan Project Area


Figure 8-2: Graphite Mineralisation of the Kringelgruvan Project Area


SECTION 9 EXPLORATION

Exploration in the early 1980s proceeded under the direction of the Swedish Geological Survey and

subsequently by MIRAB following their acquisition of the exploration and mining leases from the

Swedish State in 1992.

MIRAB defined and evaluated the Woxna graphite deposits and then sold the project to Tricorona AB

who constructed the Woxna graphite mine, where production started in 1996 by its subsidiary, Woxna

Graphite AB. Production ceased in 2001.

Initial discovery was of a single large mineralised boulder in Quaternary moraine float on the

Kringelgruvan claim. Systematic exploration took place from 1985 onwards using geophysics as the

primary anomaly definer. Magnetic, radiometric and electromagnetic methods were used; of these,

electromagnetic techniques have proved to be the definitive target generator. Electromagnetic

methods rely on an electrical or field response being induced on target rock types with responses

predicated on host rock mineralogy. VLF (very low frequency) Slingram methods at 3.6 kHz and 60 m

coil spacing proved to be the optimal settings. Kringelgruvan has been covered at 100 m x 80 m to

200 m x 80 m profile spacing. Follow-up diamond drilling took place from 1988-1989. All drilling is

diamond drilling at 35 mm core size diameter and all core is half-sectioned, with samples submitted

for analysis to the Sveriges Geologiska AB in Luleå (in the case of the Kringelgruvan tenement). All

sampling was analysed using the Leco thermal IR (infrared) methodology.

The Leco family of analyses are digitally controlled and designed to measure the carbon and sulphur

content in a wide variety of organic materials, as well as inorganic samples including soil, cement and

limestone. Analysis begins by weighing out a sample into a combustion crucible. On analysis, the

sample is typically combusted at >1 350 °C in a pure oxygen environment. All sample materials

contained in the crucible go through an oxidative reduction process which causes carbon-bearing

compounds to break down, producing elemental carbon, which oxidises to form CO2. From the

combustion chamber, the gases flow through two Anhydrone (MgClO4) tubes to remove moisture,

through a flow controller (3.5 l/min) then through to an infrared (IR) detection cell. The IR cell

measures the concentration of carbon dioxide gas present. The LECO analysers have an inherent

manufacturer specified accuracy of +/- 1 % carbon present.

During the spring and summer of 2012 Flinders Resource Ltd through its Swedish subsidiary Woxna

Graphite AB completed 3 673 m of drilling over 41 diamond drill holes at the Kringelgruvan mining

licence.

Sixteen north – south orientated profiles were drilled across the Kringelgruvan deposit at 50 m

spacing. Three of these profiles were infill type drilling on profiles existing from the historical drilling

programs. Of the 3 673 m of drilling, 270 m was overburden drilling, the remaining 3 403 m being


core. Drilling was completed by contractor Ludvika Borrteknik AB using a GM100 rig and BGM size

rods producing a core with diameter 42 mm.

Originally it was planned to drill 36 holes for a total of 3 000 m, but as initial results were favourable it

was decided to extend the program. Five additional drill holes were drilled resulting in the 41 holes

(Table 9-1).

The location of the 2012 drill holes are shown as red dots on Figure 10-1.

Table 9-1: Flinders Drilling of the Kringelgruvan Project

Hole Type YEAR Hole

Number Meters Tenement


The profile spacing is approximately 50 m and distance between holes on section is generally 50 m.

Most holes are dipping 50 °. Hole lengths are typically be around 100 m, resulting in a vertical depth

test of around 80 m. Shorter holes were drilled where the graphite was intersected close to surface.

Hole numbering starts with the abbreviation KRI followed by the year (12) and ends with a continuous

hole no from KRI12001 to KRI12041.

Twelve drill holes have been deviation surveyed to date. The start azimuth was measured using the

Reflex Azimuth Pointing System (APS), which is a GPS based compass that measures true north

azimuth and is not affected by magnetic disturbance. Any uncertainty in drill hole trend caused by the

lack of surveys is considered minor at the spacing of the drill holes and relatively short hole length in

relation to a scale of the resource.

Drill holes were laid out with the aid of a GPS, with hole spacing confirmed by tape and compass. At

the end of the drilling program, an independent Swedish company conducted a DGPS survey during

which the location of 40 holes where measured with an accuracy of between 0.01 m and 0.2 m (XYZ).

Additional DGPS surveying was conducted by Tyréns in January 2013. Where possible, Flinders

Resources Ltd has surveyed all drill collars by DGPS. The exception is the drill collars now located in

the bounds of the pit which were removed during mining. Position for these drill holes has been

calculated by converting the historic local grid into coordinates of SWEREF99 and are assumed

accurate.

The rock competence in holes viewed by Reed Leyton was in general very good. Fractured rock was

encountered only locally.


SECTION 10 DRILLING

Drilling activity reported is from exploration conducted by previous operators, primarily Sveriges

Geologiska AB and Flinders Resources Ltd. All remnant drill core, after sampling, is stored in wooden

core boxes at the Kringelgruvan project.

Sub-cropping graphite mineralisation was the first zone of mineralisation to be discovered by

geophysical methods and subsequently developed. At Kringelgruvan, graphite mineralisation is

associated with schlieren (sheared restite remnants) associated with pegmatite intrusion into

Paleoproterozoic age meta-argillites and meta-tuffites (Claesson et al., 1988; Claesson et al., 1989a;

Claesson et al., 1989b). Coarse-, medium- and fine-grained graphite is developed as coarse blebs in

monomineralic zones. Parts of the mineralised zone contain wispy pyrrhotite (Fe1-xS2). The

combination of both graphite and pyrrhotite is the cause of the strong geophysical response to ground

electromagnetic techniques applied during early exploration.

A total of 2 909 m of diamond core was drilled (Table 6-1) on the tenement in 1988 and 1989 and

comprises 51 diamond drill holes (35 mm core size). These holes occur on 6 cross-sections in local

grid coordinates extending over an approximate strike length of 600 m. Drill spacing is a nominal

20 m x 50 m. All drill core beneath Quaternary age moraine deposits (3 m to 20 m depth) was

sampled continuously. All 1988 drilling was at -60º dip and 1989 drilling at -50 ° to -55 ° dip.

Mineralisation dips to the SW at 70 ° to 80 °.

The historic program resulted in 374 graphite analyses (carbon by Leco analyser) and 52 sulphur

assays (by Leco). Selective samples were analysed by whole rock ICP for major and minor elements

and LOI (loss on ignition). Twenty-five samples were submitted to the petrophysical laboratory of

Sveriges Geologiska AB for density measurements using the Archimedes (water immersion) method.

Drilling conducted by Flinders Resources Ltd was carried out in 2012. A total of 41 holes comprising

3 673 metres of diamond core were drilled. Inside hole diameter was 42 mm. The holes were

designed to: 1) infill on historic drill sections or 2) as step-outs on a 50 m x 50 m nominal spacing.

The drill dip was -50 ° and the azimuth was either 16 ° or 340 °. Combined with the historic drilling, a

total of 6 581 m have now been drilled at Kringelgruvan.

All Flinders drill core was sampled continuously based on the logged geology. The program resulted

in 1 345 graphite analyses (carbon by Leco analyser). Every 3rd sample (33 %) was assayed for

Sulphur (by Leco analyser) up until drill hole KRI12015. Flinders was advised to assay every 2nd

sample for sulphur, starting with drill hole KRI12016. The average is every 3rd sample (33 %) was

assayed for Sulphur (by Leco analyser). In Total 441 samples were assayed for Sulphur. Every 12th

sample (8 %) was assayed by ICP-MS for major and minor elements. Density measurements were

conducted by Flinders staff using the Archimedes method. In total 1 424 measurements were made

covering each assay interval as well as the lithologies in the foot and hanging walls.


No historic data has been presented on sample recovery, but visual inspection of remnant core

indicates that sample recoveries were not good. Sample recovery from Flinders Resources drilling

was generally >95 %.

Figure 10-1: Drilling at the Kringelgruvan Project Area

Table 10-1: Kringelgruvan Drill Collar Co-ordinates (SWEREF99 TM Grid)

Hole_ID Easting

(m)

Northing

(m)

RL

(m) Total Depth

Dip

(°)

Azimuth

(°) Drill Type

Hole Size (mm)

KRIN88001 533 050 6 808 524 254.28 52.15 - 60 349 DD 35 mm

KRIN88002 533 058 6 808 507 252.52 48.80 - 60 349 DD 35 mm

KRIN88003 533 098 6 808 546 256.47 35.75 - 60 345 DD 35mm

KRIN88004 532 963 6 808 551 251.34 37.30 - 60 65 DD 35 mm


Hole_ID Easting

(m)

Northing

(m)

RL

(m) Total Depth

Dip

(°)

Azimuth

(°) Drill Type

Hole Size (mm)

KRIN88005 532 946 6 808 542 250.82 46.3 - 60 65 DD 35 mm

KRIN88006 532 980 6 808 583 255.10 25.05 - 60 345 DD 35 mm

KRIN88007 532 983 6 808 560 253.36 99.6 - 60 349 DD 35 mm

KRIN88008 533 007 6 808 607 257.50 34.15 - 60 349 DD 35 mm

KRIN88009 532 938 6 808 692 266.06 34.25 - 60 345 DD 35 mm

KRIN88010 532 955 6 808 655 260.42 48.85 - 60 345 DD 35 mm

KRIN88011 532 966 6 808 636 257.73 57.65 - 60 345 DD 35 mm

KRIN88012 532 985 6 808 691 265.32 30.35 - 60 349 DD 35 mm

KRIN88013 532 991 6 808 672 263.38 48.7 - 60 349 DD 35 mm

KRIN88014 532 992 6 808 642 259.04 43.1 - 60 349 DD 35 mm

KRIN88015 533 040 6 808 686 268.74 35.75 - 60 349 DD 35 mm

KRIN88016 533 060 6 808 654 263.05 56.15 - 60 345 DD 35 mm

KRIN88017 532 991 6 808 511 250.38 94.15 - 60 358 DD 35 mm

KRIN88018 532 899 6 808 673 261.82 39.8 - 60 349 DD 35 mm

KRIN88019 532 973 6 808 594 255.14 72.7 - 60 334 DD 35 mm

KRIN88020 533 001 6 808 620 256.09 66.8 - 60 349 DD 35 mm

KRIN88021 533 079 6 808 619 255.19 81.85 - 60 334 DD 35 mm

KRIN88022 533 095 6 808 596 258.22 100.85 - 60 334 DD 35 mm

KRIN88023 533 098 6 808 680 267.29 48.7 - 60 345 DD 35 mm

KRIN88024 533 120 6 808 643 264.08 76.8 - 55 345 DD 35 mm

KRIN88025 533 133 6 808 618 261.96 86.6 - 50 345 DD 35 mm

KRIN88026 533 019 6 808 573 255.83 98.8 - 50 345 DD 35 mm

KRIN88027 532 927 6 808 557 248.14 48.05 - 55 349 DD 35 mm

KRIN88028 532 869 6 808 508 252.12 45.75 - 50 341 DD 35 mm

KRIN89001 531 936 6 808 520 222.67 37.1 - 50 16 DD 35 mm

KRIN89002 531 931 6 808 500 222.50 59.25 - 51 12 DD 35 mm

KRIN89003 532 183 6 808 529 226.77 39.9 - 49 11 DD 35 mm

KRIN89004 532 179 6 808 510 226.41 46.9 - 51 21 DD 35 mm

KRIN89005 532 411 6 808 570 233.03 44.35 - 47 20 DD 35 mm

KRIN89006 532 405 6 808 547 233.02 69.2 - 42 14 DD 35 mm

KRIN89007 531 302 6 808 230 237.65 47.55 - 50 10 DD 35 mm

KRIN89008 532 562 6 808 559 245.95 45.5 - 60 22 DD 35 mm

KRIN89009 531 298 6 808 211 237.13 62.5 - 50 10 DD 35 mm

KRIN89010 532 557 6 808 536 244.02 59.65 - 55 18 DD 35 mm

KRIN89011 531 977 6 808 092 217.95 40.45 - 50 15 DD 35 mm

KRIN89012 532 744 6 808 636 248.02 55.25 - 53 18 DD 35 mm


Hole_ID Easting

(m)

Northing

(m)

RL

(m) Total Depth

Dip

(°)

Azimuth

(°) Drill Type

Hole Size (mm)

KRIN89013 531 971 6 808 073 218.50 61.8 - 50 15 DD 35 mm

KRIN89014 532 738 6 808 612 254.73 60.3 - 50 15 DD 35 mm

KRIN89015 532 757 6 808 532 250.70 41.1 - 49 14 DD 35 mm

KRIN89016 532 753 6 808 501 249.01 46 - 60 12 DD 35 mm

KRIN89017 532 819 6 808 426 246.01 44.85 - 60 11 DD 35 mm

KRIN89018 532 745 6 808 465 247.25 67.4 - 58 13 DD 35 mm

KRIN89019 532 881 6 808 465 250.84 67.95 - 60 348 DD 35 mm

KRIN89020 532 979 6 808 402 247.34 38.6 - 60 8 DD 35 mm

KRIN89021 532 927 6 808 616 256.23 63.2 - 60 342 DD 35 mm

KRIN89022 532 728 6 808 559 252.34 101.65 - 49 8 DD 35 mm

KRIN89023 532 553 6 808 483 240.70 113.1 - 60 13 DD 35 mm

KRI12DD001 532 849 6 808 620 258.00 48.9 - 50 358 DD 42mm

KRI12DD002 532 853 6 808 578 253.77 77.55 - 50 347 DD 42mm

KRI12DD003 532 921 6 808 497 251.11 119.65 - 50 340 DD 42mm

KRI12DD004 532 826 6 808 446 247.95 80.6 - 50 23 DD 42mm

KRI12DD005 532 783 6 808 447 246.24 49.3 - 50 18 DD 42mm

KRI12DD006 532 798 6 808 494 248.93 127.35 - 50 29 DD 42mm

KRI12DD007 532 807 6 808 542 252.09 124.2 - 50 349 DD 42mm

KRI12DD008 532 758 6 808 560 253.82 101.6 - 49 12 DD 42mm

KRI12DD009 532 761 6 808 594 257.38 68.7 - 55 19 DD 42mm

KRI12DD010 532 818 6 808 583 255.52 94 - 50 7 DD 42mm

KRI12DD011 532 671 6 808 634 240.34 71.9 - 50 20 DD 42mm

KRI12DD012 532 622 6 808 635 240.28 54.1 - 50 15 DD 42mm

KRI12DD013 532 859 6 808 506 251.99 137.45 - 50 13 DD 42mm

KRI12DD014 532 662 6 808 574 252.41 113.5 - 50 15 DD 42mm

KRI12DD015 532 598 6 808 558 246.94 113.35 - 50 15 DD 42mm

KRI12DD016 532 631 6 808 480 248.69 184.5 - 50 22 DD 42mm

KRI12DD017 532 594 6 808 499 247.53 87.5 - 50 17 DD 42mm

KRI12DD018 532 583 6 808 461 243.66 145.1 - 50 7 DD 42mm

KRI12DD019 532 490 6 808 480 235.36 115.1 - 50 8 DD 42 mm

KRI12DD020 532 489 6 808 519 238.74 92.6 - 49 19 DD 42 mm

KRI12DD021 532 444 6 808 537 233.11 77.7 - 50 13 DD 42 mm

KRI12DD022 532 434 6 808 491 232.18 122.15 - 50 37 DD 42 mm

KRI12DD023 532 328 6 808 463 229.97 84.45 - 50 22 DD 42 mm

KRI12DD024 532 338 6 808 515 230.76 80.4 - 50 15 DD 42 mm

KRI12DD025 532 347 6 808 556 233.41 48.45 - 50 16 DD 42 mm


Hole_ID Easting

(m)

Northing

(m)

RL

(m) Total Depth

Dip

(°)

Azimuth

(°) Drill Type

Hole Size (mm)

KRI12DD026 532 277 6 808 484 227.80 79.2 - 50 11 DD 42 mm

KRI12DD027 532 265 6 808 435 230.32 118.2 - 50 17 DD 42 mm

KRI12DD028 532 231 6 808 507 227.40 62.9 - 50 14 DD 42 mm

KRI12DD029 532 218 6 808 452 226.72 86.8 - 50 15 DD 42 mm

KRI12DD030 532 171 6 808 464 226.56 83.25 - 50 16 DD 42 mm

KRI12DD031 532 120 6 808 473 226.33 83.6 - 50 15 DD 42 mm

KRI12DD032 532 132 6 808 530 225.68 40.4 - 51 13 DD 42 mm

KRI12DD033 532 086 6 808 537 224.88 49.4 - 49 16 DD 42 mm

KRI12DD034 532 072 6 808 493 225.23 74.55 - 49 13 DD 42 mm

KRI12DD035 532 037 6 808 547 224.00 41.9 - 49 15 DD 42 mm

KRI12DD036 532 024 6 808 499 225.96 71.6 - 50 14 DD 42 mm

KRI12DD037 532 386 6 808 500 231.31 98.55 - 49 13 DD 42 mm

KRI12DD038 532 456 6 808 583 233.41 41.5 - 46 30 DD 42 mm

KRI12DD039 532 754 6 808 531 250.64 111.9 - 50 14 DD 42 mm


SECTION 11 SAMPLE PREPARATION, ANALYSES, AND SECURITY

Historical sample preparation methods and quality control measures employed before dispatch of

samples to an analytical or testing laboratory have not been documented, nor the method or process

of sample splitting and reduction, nor the security measures taken to ensure the validity and integrity

of samples taken.

All samples were half-sectioned and submitted in geologically meaningful lengths. Paper records of

these are available to Woxna. All lengths quoted are down hole and not “true” widths.

At Kringelgruvan, 374 valid carbon and 52 valid sulphur analyses are presented in both paper and

database (.dbf) format. The laboratory that completed analysis of the Kringelgruvan samples was the

Government owned SGAB ANALYS, (Box 801, Luleå, Sweden 95128).

The laboratories that carried out the sampling and analytical work are independent of Woxna and

previous project vendors. No details of certification by any standards associations and the particulars

of any certification are known, however the laboratory was well regarded and applied best practice of

the day.

No detail of the nature, extent, and results of quality control procedures employed and quality

assurance actions taken have been provided. This being so, the resource is quoted as a historic

estimate.

Drill core from the 2012 program was logged at the Kringelgruvan mine site. Once geologically

logged, RQD measurements were taken. Core was then photographed, and magnetically measured

prior to storage on pallets at the mine site. Regular batches of samples were then sent via

independent contractor.

Flinders geologists Lars Dahlenborg, Elin Ryösä and Janne Kinnunen supervised sampling of all

holes drilled in 2012.

Samples intervals were marked on the core and the core tray. Each interval was given a unique

sample number. The sample numbers were taken from unique sample ticket booklets made for

Flinders. One part of the sample ticket was placed in the bag together with the cut core. The sample

numbers range from 26 700 to 28 300. A total of 92 blank samples were inserted at a rate of

approximately 1 in 15, resulting in approximately 7 % of the submitted samples being blanks.

The core was then shipped by truck to ALS in Piteå for core cutting. The core was split by diamond

saw and put back in the core tray by ALS. Subsequently Flinders’ staff, supervised by Elin Ryösä or

Lars Dahlenborg bagged the samples. One half of the core was placed in a numbered plastic bag

together with the corresponding sample ticket and the other half was left in the core tray. The core

was cut taking in consideration the main foliation/banding of the rock. When it was possible to

reassemble the core, the same half of the core was submitted for assay. The residual half of drill core


was viewed by Reed Leyton in the mine site. The mine site is a key access only facility, and there is

no evidence that samples have been disturbed in any way since cutting.

The plastic bags containing samples were then handed over to ALS Chemex preparation laboratory in

Piteå.

The rocks on the property are fresh with little or no secondary minerals on the surfaces that would

enhance metal values.

Cutting of core and dispatch to the ALS Chemex laboratory in Sweden is in keeping with industry

practice, and security of the delivery chain is more than adequate. All drilling and subsequent

sampling and assaying during the 2012 drilling program was completed by independent persons and

at no time was an officer, director or associate of Flinders involved.


SECTION 12 DATA VERIFICATION

The adequacy, archiving and standard of the data presented is of sufficient quality for the reporting,

subject to qualification, of the historical drilling and resources as presented by previous project

owners.

Where possible, Flinders Resources Ltd has surveyed all drill collars by DGPS. The exception is the

drill collars now located in the bounds of the pit which were removed during mining. Position for these

drill holes has been calculated by converting the historic local grid into coordinates of RT90 and are

assumed accurate. The RT90 coordinates were further converted into SWEREF 99 TM by Tyréns in

January 2013.

Coffey Mining had transcribed paper records for collar, assay, survey and geology data, as reported

by Claesson et al. (1991, 1992, 1993), for Kringelgruvan and compared these to digital data available

to Woxna. Coffey Mining concluded that the historical data is of sufficient quality and traceable

provenance that it is useable as exploration data. The then supervising geologist, who is a QP under

current NI 43-101 protocol, also verified the provenance of the data supplied.

12.1 DRILL CORE

Twelve drill holes were examined in detail on 13 June 2012, at the Kringelgruvan mine site in Central

Sweden (Table 12-1).

Table 12-1: Drillholes and Core Examined by Reed Leyton Representative

Project Hole Number Approx Meters examined

Kringelgruvan KRIN88014 40






Kringelgruvan KRI12DD001 42







All core trays were laid out separately on the examination tables, washed down, and checked.

The core lithology and mineralogy was checked visually against the available copies of original core

logs for verification. No discrepancies were noted. Analytical result sheets were also checked against

the actual core and the core logs (to check sample locations).

Core sections of interest were photographed.

12.2 CHECK SAMPLING

The majority of the core examined had been previously cut by diamond saw into halves, and some

sections quartered. Many original sample intervals had been noted on the actual wooden core trays,

and occasionally on the core remnant itself.

It was decided to re-sample core lengths as close to the original lengths as possible for direct

comparison.

The re-sampling included 59 samples (Table 12-2). The core trays selected for re-sampling were

taken to the Flinders core saw facility by a Flinders field assistant, and quarter-core (as appropriate

and available) sections were re-sawn for the check samples.

Reed Leyton checked that the correct samples were taken, sawn, and the resulting sample bulks

were placed in individual sample bags with an identifying tag. The bags were sealed with a plastic tie.

The bags were retained under Reed Leyton’s supervision, and personally delivered to the ALS

Chemex laboratory manager (Tony Ökvist) at Öjebyn (Sweden) for further processing and transport.

Table 12-2: Check Sample Intervals by Reed Leyton Representative

Project Hole

Number

Check

Sample

From (m)

Check

Sample

To (m)

Check

Sample

Interval (m)

Check

Sample

Number

Kringelgruvan KRIN88014 4.75 6.75 2 27 796





Kringelgruvan KRIN88014 21.75 24.05 2.3 27 802




Kringelgruvan KRIN88015 21 22.4 1.4 27 806



Project Hole

Number

Check

Sample

From (m)

Check

Sample

To (m)

Check

Sample

Interval (m)

Check

Sample

Number









Kringelgruvan KRIN89022 10 11.2 1.2 27 817





Kringelgruvan KRIN89022 82 84 2 27 823





Kringelgruvan KRI12DD001 14 15 1 27 828

Kringelgruvan KRI12DD001 23.4 24.2 0.8 27 829

Kringelgruvan KRI12DD001 43.5 44.5 1 27 830

















Project Hole

Number

Check

Sample

From (m)

Check

Sample

To (m)

Check

Sample

Interval (m)

Check

Sample

Number






Kringelgruvan KRI12DD009 57.8 58.4 0.6 27 853







12.3 CHECK ANALYSES

No information exists on check analyses from historical operators of the project. Check sampling by

Flinders Resources has consisted of crush duplicates and pulp duplicates. Greater than 10 % of each

batch sent to ALS Chemex has been composed of these duplicates.

All QA/QC data for this Project has been deemed acceptable for the purposes of the Mineral

Resource estimation.

0

5

10

15

20

0 5 10 15 20

C_Duplicates


Figure 12-1: Coarse Duplicate Data for Cg

Figure 12-2: Pulp Duplicate Data for Cg

The paired plots in Figure 12-1 and Figure 12-2 demonstrate high degree of correlation between

Parent data (Vertical axis) and duplicate data (Horizontal axis). Blue lines provide 1 to 1 correlation

trend.

12.4 RESULTS AND DISCUSSION

Final results were received via direct email from ALS Chemex on 18 June 2012. Both the raw data

and the analysis certificate were received. Table 12-3 shows the analysis values for Cg only,

comparing the original sample interval and value versus the check sample interval and value. There is

extremely good agreement between the individual samples.

Table 12-3: Drill Core Re-sampled for Check Analysis, Matched with Original Assays

Project Hole

Number

From

(m)

To

(m)

Original Data

Interval

(m)

Original Data

Cg

(%)

Check

Sample

Number

Check

Data

Cg

(%)

Kringelgruvan KRIN88014 4.75 6.75 2 16.3 27 796 12.25




Kringelgruvan KRIN88014 19.75 21.75 2 6 27 801 5.31

0

5

10

15

20

0 5 10 15 20

P Duplicates


Project Hole

Number

From

(m)

To

(m)

Original Data

Interval

(m)

Original Data

Cg

(%)

Check

Sample

Number

Check

Data

Cg

(%)

Kringelgruvan KRIN88014 21.75 24.05 2.3 4.3 27 802 3.49




Kringelgruvan KRIN88015 21 22.4 1.4 5.1 27 806 4.29

Kringelgruvan KRIN88015 22.4 24.6 2.2 5 27 807 4.66









Kringelgruvan KRIN89022 10 11.2 1.2 3.4 27 817 3.06




Kringelgruvan KRIN89022 79.35 79.9 0.55 - 27 822 1.1

Kringelgruvan KRIN89022 82 84 2 8.6 27 823 7.86





Kringelgruvan KRI12DD001 14 15 1 8.77 27 828 8.51

Kringelgruvan KRI12DD001 23.4 24.2 0.8 2.68 27 829 2.52

Kringelgruvan KRI12DD001 43.5 44.5 1 1.77 27 830 1.41








Kringelgruvan KRI12DD007 74.6 75.6 1 0.05 27 838 <0.01


Project Hole

Number

From

(m)

To

(m)

Original Data

Interval

(m)

Original Data

Cg

(%)

Check

Sample

Number

Check

Data

Cg

(%)













Kringelgruvan KRI12DD009 57.8 58.4 0.6 3.97 27 853 4.03








Figure 12-3: Paired Historical and Modern Analytical Data for Cg

The paired plots in Figure 12-3 demonstrate high degree of correlation between historical data

(Horizontal axis) and modern data (Vertical axis). Blue lines provide 1 to 1 correlation trend.

12.5 DENSITY

A total of 1423 bulk density determinations have been completed with a range of values between 2.35

t/m3 and 3.67 t/m3. The majority of determinations range from 2.6 t/m3 to 2.8 t/m3 (Figure 12-4).

Reed Leyton has also divided the 1 423 bulk density determination by domain (Figure 12-5 and

Figure 12-6). The density determinations were calculated wet and dry weight volume determinations.

Figure 12 confirms that the majority of the determinations average 2.7 t/m3. The average for the

waste rock determinations was 2.7 t/m3.


Figure 12-4: All 1 424 Bulk Density Determinations

Figure 12-5: Domain ‘HG’ 376 Bulk Density Determinations


Figure 12-6: Domain ‘GRF’ 540 Bulk Density Determinations


SECTION 13 MINERAL TESTING AND PROCESSING

A significant amount of test work was carried out for the Woxna graphite deposit prior to construction,

including pilot plant work in 1992/3. However, the current plant is substantially different to the pilot

plant and the initial production plant. In addition, a variety of laboratory and plant trials were

performed during the life of the mine, however, GBM found only a few reports in the mine’s archives.

In addition to these reports there are process reports which relate to audits of the production and

recommendations on changing the process.

After reviewing the historical production reports including grades and size analyses of the graphite

concentrates sold, it was decided that further test work was required to assess whether it is possible

to improve the grade and increase the flake size of the products as these improvements would

significantly increase the price of the products.

13.1 RECENT TEST WORK

13.1.1 CONCENTRATE DEWATERING TESTS

Outotec was commissioned to complete concentrate dewatering test work in 2012 to support the

dewatering plant design. The test material was created using a fine graphite concentrate product

sample from the historical product which is stockpiled at Woxna. The results of this test work indicate

that thickening is not viable and that pressure filtration in horizontal or vertical filters yield filter cake

moisture contents of 22 % and 20 % respectively.

13.1.2 AMINPRO TESTS

In January to June 2013 a major test work programme was under taken by Amelunxen Mineral

Processing Ltd. (Aminpro) in Chile to reassess the process. Using the historical process as a basis,

GBM and Aminpro designed a test work programme to explore the various optimisations considered

pertinent to the plant design with the objectives of increasing the coarse flake recovery, overall plant

recovery and product grade. The key areas tests were:

Bond Work Index on feed and assessment of regrinding parameters

Rougher flotation tests to assess variables such as reagents, reagent dosage, % solids, pH

and grind

Rougher kinetic tests with screen analyses using optimum conditions

Contact cell tests

Cleaner flotation kinetic tests including grind size, pH, % solids, redox, reagent addition

points, column parameters

Rougher flotation and mineralogy variability


Locked cycle flotation tests

Microscopy in the form of mineral liberation analysis

Settling tests on tailings

Magnetic separation test work on graphite concentrate.

The test description and results are reported in the Flinders Resources Woxna-Graphite Metallurgical

Test Work and Front End Engineering Program by Roger Amelunxen.

13.1.3 SAMPLES USED IN THE AMINPRO TESTS

Two types of samples were taken from the Woxna deposit; bulk hand samples from pits in the

mineralised zone and selected intervals from core samples. Initially two bulk hand samples were

taken from a pit where the drilling had indicated that the rock was above cut-off grade. One was to

represent marginal grade ROM material and the second higher grade material. However, the ROM

sample (ROM#1) assayed only 3.3 % C (ROM is expected to be approximately 10 % C) so this was

only used for the initial tests, such as comminution and rougher flotation. The high grade sample was

handpicked in the laboratory to produce high and low grade samples. The high grade sample was

combined with some of the ROM#1 sample to form a composite, COMPO#1, with a grade, 9 % C,

approximating to the expected ROM grade. A third sample, ROM#2, was taken from another area to

provide a single sample which had a grade, 12.2 % C, approximating to the expected ROM grade.

ROM#2 was used for detailed tests including the locked cycle flotation tests.

Intervals of drill core were selected by Monte Carlo method to represent the mineralised body.

Rougher tests were carried out on these 20 intervals, and ten of these had mineralogy carried out on

them. In addition, the work index for each sample was estimated.

13.1.4 SUMMARY OF AMINPRO TEST RESULTS

The Bond Work Index of the sample was 18.5 kWh/t and for cleaner feed material greater than 40

kWh/t, dependent on the graphite grade. This confirms historical reports.

Rougher flotation tested at varying levels of grind, using MIBC as the sole collector, indicated that:

Graphite and gangue (silicates) are the coarsest components of the sample with iron

sulphides (mainly pyrrhotite) being present as fines.

Graphite floated very fast with good recoveries (>95 %) in all tests. Gangue and iron

sulphides had lower recoveries and flotation speeds.

Flash flotation in roughers should be considered to capture as coarse a flake product as

possible and that the finer graphite would be recovered in the scavengers.

In testing other variables in the roughers, the following was observed

No effect was observed on graphite recovery and grade when varying the feed %

solids


No significant effect was seen in varying the pH

No positive effect occurred in testing other reagents (diesel, Lila Flot GS13 and

Aerofroth 88 were compared to MIBC).

Contact cell tests on rougher feed gave good results with regard to recovery, but failed to produce

high (above 90 % C) concentrate grades.

Microscopy (Mineral Liberation Analyser) and liberation studies on the variability samples found that

in the majority of the samples muscovite minerals were locked with the graphite flakes, possibly

preventing upgrading concentrates in the cleaner stage without further regrind. It showed that

between 14 % and 20% of the graphite was locked. This finding was confirmed with additional MLA

tests on a concentrate sample. This MLA also confirmed the results of previous MLA completed on

ten geological samples in 2012 which nine samples had muscovite values of between 10.3 % and

25.6 %.

The following conclusions were reached for the cleaner flotation:

Regrinding was necessary to achieve higher grades.

Flotation at low densities was found very important.

Use of dispersants was seen to help depression of gangue.

Concentrate grades with above 90 % C were achieved in a number of tests aimed at developing the

cleaner circuit configuration, at the optimised conditions.

A locked cycle test was performed at optimised flotation conditions requiring seven cycles to achieve

stability and equilibrium. The overall graphite recovery obtained for the ROM#2 sample, with a head

grade of 12.2 % C, was above 96 % while producing concentrate grades above 93 % C.

Approximately 12 % of the feed mass was recovered as concentrate. Of the 12 %, 5.3 % reported into

the rougher cleaners and the rest into the rougher-scavenger cleaners. The products of the locked

cycle test were screened showing that over 18% of the concentrate reported to the +250 µm size

assaying 95 % graphite and over 39 % of the concentrate reported in the +180 µm size assaying over

94 % graphite.

The assay results of the tests with size are shown in Table 14-7.

Table 13-1: Results of the Locked Cycle Flotation Tests

Size fraction µm



Combine concentrate


+250 13.9 95 22.0 95 18.4 95

+180 -250 18.8 97 23.4 92 21.4 94

+100 -180 26.5 94 29.6 91 28.3 92


Size fraction µm



Combine concentrate


-100 40.8 89 25.0 87 31.9 88

Figure 13-1: Locked Cycle Flotation Test Diagram

Settling and rheology tests were carried out on a tailings sample to assess the potential for thickening

the final tailings before pumping to the TSF. The following conclusions were reached from these tests:

16 g/t of SNF 2070 flocculant is required to achieve a settling rate of 0.15 m².d/t

Thickened tailings will be between 65 % and 68 % solids to allow pumping by horizontal type

pumps.


SECTION 14 MINERAL RESOURCE ESTIMATES

The Mineral Resources estimated and disclosed herein supersede the Historic Mineral Resources as

quoted in Section 6 above and apply current practices and assumptions.

14.1 RESOURCE DATA

14.1.1 DRILLING DATA

Ninety two (92) diamond drill holes totalling 6 581 m, drilled into the Kringelgruvan area in 1988, 1989

and 2012.

Data relating to the collar locations, drill collar orientations and drill hole surveys were sighted by the

Geoff Reed in sections and plans of the day. Individual hard copy data of down hole surveys or

assays were sighted.

Geoff Reed inspected the area with Woxna’s personnel and was able to locate many 2012 drill hole

collars, and selected 1988 and1989 collars.

The profile spacing is approximately 50m and distance between holes on section is generally 50m.

Most holes are dipping 50 degrees. Hole lengths are typically be around 100m, resulting in a vertical

depth test of around 80m. Shorter holes were drilled where the graphite was intersected close to

surface. Hole numbering starts with the abbreviation KRI followed by the year (12) and ends with a

continuous hole number from KRI12001 to KRI12041.

Half of the drill holes have been deviation surveyed to date. The start azimuth was measured with a

hand held compass. Any uncertainty in drill hole trend cause by the lack of surveys is considered

minor at the spacing of the drill holes and relatively short hole length in relation to a scale of the

resource.

DGPS surveying was conducted by Tyréns in January 2013. Where possible, Flinders has surveyed

all drill collars by DGPS. The exception is the drill collars now located in the bounds of the pit which

were removed during mining. Position for these drill holes has been calculated by converting the

historic local grid into coordinates of SWEREF99 and are assumed accurate.

All drill core for Ninety two (92) diamond drill holes has been located by Woxna’s staff in Woxna,

Sweden. Core from 12 holes has been inspected by the Geoff Reed.

Ninety two (92) diamond drill holes were included in the current mineral resource estimation.


Table 14-1: Kringelgruvan Drilling Database Summary

Hole Type Drill Series Drill Number Drill Meters Resource Intersection

Meters

DD 88 28 1 595 512.2

DD 89 23 1 313 284.9

DD 12 41 3 673 960.5

Total 92 6 581 1757.6

14.1.2 DATABASE INTEGRITY

Capture of digital data was completed by the Issuer’s staff. Hard copy data has been verified and all

data is stored in a database and managed by the Issuer.

Drilling data from drill programs were transferred in digital format by the Issuer’s staff.

Digital data has been both randomly and systematically checked by Reed Leyton and shown to be

correct using a number of checks listed below. Assay data in original laboratory sheets has not been

sighted from the 1988,1989 drilling program.

The digital data was compiled directly into Microsoft Excel by the Issuer, validated in Microsoft Access

and exported into a csv format. The database was then imported into Maptek Vulcan software in the

csv format.

The database for Kringelgruvan was attributed to Ninety two (92) drill holes, which provided the

verified information for compositing (specifically the collar, survey, lithology and assay tables). The

database included drill holes with recorded collar elevation. This database was named viersh.vih.isis.

14.1.3 DRILL SPACING

Ninety two (92) drill holes for 6 581 m of diamond drilling were drilled at Kringelgruvan, 90 drill holes

intersected mineralisation and were subsequently assayed.

Hole spacing was completed on a 50 m x 50 m drill pattern.

For wire framing purposes 90 degrees strike was considered the optimal orientation. Strike of

mineralisation varied from 80 degrees to 100 degrees

Polygons were created every 50 m through the ninety 90 resource drill holes at the project.

14.1.4 DRILLING ORIENTATION

Holes have been drilled mostly at two orientations 16 degrees and 348 degrees at Kringelgruvan.


Due to the amount of drilling and orientation, the true thickness is generally considered to be

70 % of drilled thickness.

The likelihood that mineralisation is developed in an orientation other than that interpreted is

considered to be low.

14.1.5 CHEMICAL ANALYSIS

Core drilled were sampled and analysed by ICP method at the laboratory of SGAB ANALYS,

(Box 801, Luleå, Sweden 95128).

A total of 374 samples from 50 drill holes were analysed in total at Kringelgruvan for diamond drill

holes with the current resource estimation.

The method applied by was the standard for the industry of the day, and although no quality

assurance data is available, it is considered to be of a very high quality.

A total of 1433 samples from 41 drill holes were analysed in total at Kringelgruvan for diamond drill

holes with the current resource estimation.

Core drilled were sampled and analysed by ICP method at the laboratory of ALS Chemex, Pitea. Sweden.

The method applied is to current industry standards, and although no standard data is available, it is

considered to be of a very high quality.

14.1.6 SAMPLE LENGTH

All holes drilled at Kringelgruvan were sampled with an average of one (1) m intervals. Check

sampling by Woxna at the request of Reed Leyton used identical sample intervals.

Composites of the drill hole assays are generated using Maptek Vulcan software with run lengths of

1 m.

These composites honour the geological wireframes. Checking was undertaken by generating an Isis

file and visually inspecting the result of the composite.

Specific components of the compositing include:

Run Lengths of 1 m;

Data Field C_pct was composited;

The composite file was then applied a tag for each composite with the character

(a2, a3, a6, a9, a13, a14, c01-c03, c05-c06, c08-c19) in the ‘bound’ column. This new

composite isis file was called viersa.cmp.isis and used in the estimation process.


Figure 14-1: Histogram of Raw Sample Lengths for Kringelgruvan

14.1.7 RELATIVE DENSITY

Using the bulk density (“BD”) density default function of Vulcan, the variable BD was populated.

The value 2.7 was run according to density test work by Woxna previously attributed to various

assays within the geology database. The author of the Reed Leyton report has created a file with an

average BD taken between various C % grades within the resource and waste blocks outside the

resource.

14.1.8 GEOLOGICAL MODEL

Mineral Resources has been calculated by Reed Leyton on the bearing of 90 degree strike.

The project was drilled within an area approximately 1 200 m x 100 m to 200 m.

The mineralisation was intersected on all the drilling sections and is so far known to at least a depth of

150 m below the surface.

Mineralisation is present as a six main mineralised bodies and eleven smaller mineralised bodies. The

six main grade domains were defined using a lower cut-off grade of 7 % Cg while a broader outer

domain used the lithology code FGRF. The other eleven smaller mineralized bodies were defined

using the lithology code FGRF. The thickness in the section of the plane was usually more than 10m,

but varied between 5 m and a little more than 15 m.


One block model were constructed, named vie_woxna_apr2013_75.bmf. The parameters used in the

setup file vieuncutivdcpct1_10sep.bef for Kringelgruvan.

The block model was created using the one bdf file, vie_woxna_apr2013_75.bdf. This original block

model contained only default values except for the variable domain, which was populated in relation

to the wireframes in which the blocks resided in.

A rotation of 90 Bearing, 0 Plunge and 0 Dip was applied.

Parent block size was 5 m x 25 m x 5 m with sub blocks at 1.25 m x 5 m x 1.25 m.

An offset of 1 500 m x 600 m x 400 m was applied.

The variables include the type and their default values before estimation.

Figure 14-2: Mineral Resource Cross Section, Kringelgruvan

14.1.9 WIRE FRAMING

Using the above drill hole data, wire framing of the geological boundaries were performed by joining

digitised section outlines at 50 m spacing.


The digitised sections are snapped to drill holes within +/-25 m influence using above 7 % C for the 10

high grade wireframe domains at Kringelgruvan.

The digitised sections are snapped to drill holes within +/-25 m influence using the mineralized

boundary as the low grade wireframe domains boundary limit at Kringelgruvan.

Vertical plane sections were digitised at 12 degree and 348 degree orientation at 50 m spacing. There

is sufficient evidence for continuity of the mineralised envelope between sections.

All modelled wireframes were checked in plan, cross section, long section and 3D rotated views

All geological wireframes were checked for crossing, inconsistencies and closure.

All wireframes were updated too match the new drillhole collar coordinates and the adjusted

mineralization data points.

Table 14-2: Kringelgruvan ‘hg’ Domain (above 7 % Cg) and ‘grf’ Domain (lith code FGRF)

‘hg’ Domain ‘grf’ Domain

a2, a3, a6, a9, a13, a14 c01-c03, c05-c06, c08-c19

Table 14-3: Kringelgruvan Domain Volume Validation

Domain Wireframes Volume Model Volume Domains

a02 293 653 291 865 99 %

a03 237 686 232 627 98 %

a06 591 542 559 453 95 %

a09 181 821 177 109 97 %

a13 53 299 52 139 98 %

a14 6 789 6 787 100 %

c01 44 915 44 834 100 %

c02 476 361 475 010 100 %

c03 598 359 589 014 98 %

c05 58 427 56 729 97 %

c06 1 397 402 1 401 015 100 %

c08 174 655 174 629 100 %

c09 517 768 514 478 99 %

c10 26 177 26 406 101 %

c11 27 350 27 207 99 %

c12 2 457 2 432 99 %

c13 72 825 72 157 99 %

c14 14 587 14 661 101 %


c16 3 938 3 818 97 %

c17 7 282 7 363 101 %

c18 10 177 10 264 101 %

c19 5 417 5 195 96 %

Total 4 802 886 4 745 191 99 %

Figure 14-3: Mineral Resource Cross Section, Kringelgruvan


14.1.10 GRADE INTERPOLATION

Grade interpolation was undertaken using inverse distance defined by the domain wireframes. The

allocations of composites were calculated using a hard boundary at the domain wireframes.

Using Maptek Vulcan’s Estimation Editor the grade estimation was run for Kringelgruvan. Variables

were populated using one single search ellipses with no cutoff to the mineralised domains.

Constant parameters used in this block estimation file, vieuncutivdcpct_1to10sep.bef include:

The grade variable populated was C_uncut. The default given was 0

The number of samples used was populated in the variable numsam. The default given

to this variable was 0

The number of drillholes used was stored in nodrill. The default given was 0

The sample distance used was stored in the variable samdis

The weighted average anisotropic distance to the samples used was populated in the

variable samdis

The inverse distance method was applied.

Table 14-4: Block Model Parameters for Kringelgruvan

Variables Description

c_uncut Carbon grade - reportable

s_uncut Sulphur grade – not reportable

nnp_uncut Nnp grade – not reportable

bd Bulk Density

category Resource category by script

mintype Mineralisation Domain

nodrill Number of Drill holes

samdis Average sample distance

numsam Number of samples

pass Estimation flag

type Air or fresh rock

mined Mined or insitu

lithtype Graphite Pegmatite Metasediment Overburden

rsc_cat Final Resource Category meas = 1, ind = 2, inf = 3, additional min = 4.


Table 14-5: Search Parameters for Kringelgruvan

Pass Min Sample Max Sample Distance

1 2 12 70

2 1 20 140

3 1 30 400

Table 14-6: Estimation Parameters for Kringelgruvan

Domain Strike Plunge Dip Major Minor Semi Major Discretisation

A2, A3, A6, A9, A13, A14 82 0 0 4 1 2 2x:4y:2z

C01-3, C05-6, C08-19 82 0 0 4 1 2 2x:4y:2z

14.1.11 MINIMUM WIDTH

No minimum width has been applied in the estimation of the Kringelgruvan Mineral Resources.

14.1.12 CUT-OFF GRADE

A cut-off grade of 7 % Cg has been applied to the Mineral Resource estimation.

14.1.13 ADDITIONAL VARIABLES

Once the estimations had run, a number of additional variables were added or calculated. These

variables included:

The category variable, category. A script, resourcecatflagged.bcf was run on the block

model. This script looked at the nearest neighbour distance variable (“samdis”). If samdis

was >0, then the category variable was set to inf (inferred). This variable was used to

classify the resource.

The category variable, called rsc_cat. A calculation, rev_rsccat1_2_3_4.bcf was run on the

block model. This calculation looked at the previous script run. This variable was used to

classify the resource based on drilling density, continuity and general confidence in each

modeled wireframe.

Using the BD density calculation function of Vulcan the variable bd was populated. The

script was run according to density test work by Woxna previously attributed to various

assays within the geology database. Reed Leyton has created a script file with an average

BD taken between various C grades.

14.1.14 MINING ASSUMPTIONS

No assumptions have been made as to future mining methods, dimensions or dilution.


14.1.15 METALLURGICAL ASSUMPTIONS

No assumptions have been made as to the metallurgical behaviour of mineralization.

14.2 MINERAL RESOURCE ESTIMATE

This Mineral Resource estimate has been prepared in accordance with the CIM Definition Standards

of June 2011(became law) or 27 November 2010 (published). The classification of the resource at the

appropriate levels of confidence are considered appropriate on the basis of drill hole spacing, sample

interval, geological interpretation and all currently available assay data.

The Kringelgruvan Mineral Resource, quoted to the appropriate level of confidence, is provided in

Table 14-7.

Table 14-7: Kringelgruvan Mineral Resource Estimate @ 7 % Cut-Off

Classification TONNES

(Mt)

Grade

Cg

%

Measured 0.99 10.68

Indicated 1.86 10.63

Total 2.85 10.65

The above numbers are literal, whereas the accuracy of the techniques requires that the estimates’

parameters should actually result in rounded figures to better reflect the order of accuracy. Hence

Reed Leyton has rounded the mineralisation tonnage to the nearest ten thousand tonnes. The

resource estimates then become as shown on Table 14-8.

Table 14-8: Rounded Kringelgruvan Mineral Resource Estimate @ 7 % Cut-Off

Classification TONNES

(Mt)

Grade

Cg

%

Measured 1.0 10.7

Indicated 1.8 10.7

Total 2.8 10.7


14.3 DISCUSSION

The Kringelgruvan Mineral Resource describes four main bodies of mineralization separated by

faulting drilled within an area approximately 1 200 m x 100 m to 200 m. The lithology logging of

graphite has been selected to best represent the margin of the mineralised body.

The sample spacing is approximately 50 m x 20 m x 1.0 m. No mining parameters are attached.

The mineralization remains open laterally and at depth.

14.4 NI 43-101 COMPLIANCE

Following the enclosed audit of historical data, Flinders data, the compiled Flinders drilling database,

and the subsequent calculation of Mineral Resources, the quoted Mineral Resources at

Kringelgruvan are subdivided into CIM-compliant measured and indicated categories on the basis of

the close density of drilling, checked grades and inter-hole continuity.

It is the opinion of Reed Leyton that this Mineral Resource estimate for Kringelgruvan satisfies the

definitions of Measured and Indicated Mineral Resources as per the CIM Definition Standards of

November 2010.

Table 14-9: Kringelgruvan Graphite Combined Resource Grade and Cumulative Tonnage at Various Cut-Off Grades

Cut-off Grade (%) Measured and Indicated

Resource (Mt) Grade Cg (%)

Contained Graphite (tonnes)

2 5.7 7.4 424 490

3 5.1 8.0 408 160

4 4.3 8.8 381 973

5 3.8 9.5 355 852

6 3.1 10.4 318 217

7.0 (Base Case) 2.8 10.7 303 509

8 2.6 10.9 286 613

9 2.3 11.3 255 508

10 1.8 11.7 214 418

Table 14-10: Kringelgruvan – Drill holes and Intervals Used in Resource Calculation

Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)

Kringelgruvan KRI12DD026 60.4 71 10.6 10.72

Kringelgruvan KRI12DD026 71 71.05 0.05 7.24


Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)



Kringelgruvan KRI12DD028 36.5 40.1 3.6 5.21









Kringelgruvan KRIN89003 11.824 22.8 10.976 9.68






Kringelgruvan KRI12DD020 67.4 71.4 4 11.95







Kringelgruvan KRIN89005 16.1 21.1 5 9.27



Kringelgruvan KRIN89010 38.1 50 11.9 12.81











Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)














Kringelgruvan KRIN88007 75 77.7 2.7 9.56

























Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)




Kringelgruvan KRIN88001 12 20 8 9.83




















Kringelgruvan KRI12DD027 103 113 10 3.30















Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)







































Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)









Kringelgruvan KRIN88019 51 53 2 4.31






























Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)







































Project Hole

Number

From

(m)

To

(m)

Interval

(m)

Cg

(%)















SECTION 15 MINERAL RESERVE ESTIMATES

There are no NI 43-101 mineral reserve estimates available to, or commissioned by, Woxna.


SECTION 16 MINING METHODS

16.1 INTRODUCTION

16.1.1 PREAMBLE

Woxna Graphite AB (Woxna) requested that Golder Associates AB (Golder) undertake pit

optimisation, design and scheduling as part of the Kringelgruvan Preliminary Economic Analysis

(PEA) study of the Woxna Project situated in Central Sweden.

Pit optimisation, design and sequencing were conducted using industry standard mining software.

The optimised open pit was established first using financial parameters. The selected shell generated

from the optimisation results formed the basis for the subsequent final pit design and mine production

schedule. The mine schedule was the key driver to estimate the contractor equipment fleet and

workforce required to manage and operate the mine. The Kringelgruvan geological model and

financial parameters have been supplied by Woxna and GBM to Golder.

16.1.2 TECHNICAL ACTIVITIES

The following technical tasks have been completed as part of this mining study:

1) Import Vulcan geological model provided into Datamine software and set up required

parameters for pit optimisation;

2) Export model outputs to Whittle software;

3) Set up Whittle Model (value, pit revenue factor ranges, processing and selling costs,

processing and mining rates) complete open pit shell optimisation;

4) Perform basic sensitivity analysis of revenue, graphite price and slope angle;

5) Prepare and present a preliminary PowerPoint presentation of results for review and

discussion;

6) Provide pit limits for subsequent geology and mine planning tasks, i.e. grade control, pit

design, waste dump design and mine scheduling;

7) Perform preliminary drill and blast design for production drilling;

8) Select equipment fleet size and determine manpower plan;

9) Provide an overall site layout drawing;

10) Present mining contractor operating expenses (OPEX) based on contractors’ estimates for

stripping, drilling, blasting and mining activities; and

16.1.3 SUPPLIED INPUT INFORMATION

Woxna and GBM supplied the following items used as a basis of design for the mining study:


Vulcan geological block model including current topography and geological boundaries;

Mineral types and grades;

Cut-off grade;

Royalties;

Modifying factors (processing recoveries);

Direct and indirect processing costs;

Graphite sale price; and

Discount rate.

Golder determined the following inputs for the mining study:

Modifying factors (mining losses and mining dilution);

Direct and indirect mining cost based on mining contractors quotes; and

Pit design parameters.

A table of the provided input parameters for the Kringelgruvan deposit that formed the basis of the

optimisation and open-pit design are included in below and summarized in the following sections.

Table 16-1: Pit Optimisation Parameters

Kringelgruvan deposit Unit Value

Graphite price USD/t 1 500

Discount factor % 10

Exchange rate SEK:USD 6.57:1

Mining royalties % 0.2

Processing plant recovery rate % 85

Waste mining cost USD/t 4.50

Graphite mining cost USD/t 7.20

Plant and infrastructure operating expenses (inclusive of general, administration and selling costs)

USD/t 70.20

Mining dilution % 2.5

Mining losses % 2.5

Inter-ramp slope angle (ramp not included) ° 55

The final processing costs, throughput and sale prices are different to those shown above, however

Golder has verified the new values do not materially change the mine design at PEA level.


16.2 GEOLOGY AND MINERAL RESOURCES

The Mineral Resources for the Kringelgruvan graphite deposit were reported in “Mineral Resource

Estimates for Kringelgruvan Woxna Deposit” Memorandum authored by Geoffrey Charles Reed from

Reed Leyton Consulting and dated April 2013.

The mineral resource estimate at a cut-off grade of 7.0 % graphite presented in the aforementioned

memorandum is summarised below.

Table 16-2: Kringelgruvan Deposit 2013 Mineral Resource Estimate

Classification Tonnes

(Mt) C

(%)

Measured 1.0 10.7

Indicated 1.8 10.7

Total 2.8 10.7

The Kringelgruvan mineral resource is reported only in the Measured and Indicated categories. The

Inferred classification category has not been reported in this mining study.

The “Technical Report for Kringelgruvan Graphite Deposit, Part of the Woxna Graphite Project,

Central Sweden, and prepared for Flinders Resources Limited,” (2012 Technical Report) authored by

Geoffrey Charles Reed from Reed Leyton Consulting and dated November 2, 2012 indicated that:

Kringelgruvan has variable cover of 2 m to 15 m of Quaternary age moraine;

Graphite is found as both flakes (>70 µm) and a finer-grained amorphous, microcrystalline

type;

Graphite has a dark streak and is visually obvious in core;

Mineral resources are classified as Indicated and Measured based on the density of drilling,

checked composited grades, inter-holes continuity;

A cut-off grade of 7 % Carbon graphite (Cg) was used as the base case to calculate the

resource; and

Grade interpolation was undertaken using inverse distance within defined mineralisation

domain wireframes.

The property has a thin overburden layer and near surface mineral resource. Consequently the

conventional open pit mining method previously used at Kringelgruvan continues to be the most

suitable and economic method to restart mining.

The Kringelgruvan open pit mining method, which includes drilling, blasting, loading, and hauling, is

designed for multiple open pit phases throughout the LOM. Drilling and blasting will be required on all


of the mineral resource and waste rock. Drilling is planned to be performed using conventional down-

the-hole hammer rigs capable of single pass drilling. Blasting will use commercial bulk explosive

(emulsion) with down-hole delay initiation. A hydraulic shovel will load 40 tonne off-highway haul

trucks.

16.3 GEOLOGICAL BLOCK MODEL

16.3.1 INTRODUCTION

Golder has undertaken open pit optimisation using the mineral resource block model prepared by

Reed Leyton Consulting.

Whittle 4x was used to develop optimal pit geometries using the Lerchs - Grossmann (LG) algorithm

with the mineral resource block model using preliminary parameters and assumptions. The optimised

open pit(s) geometries delimit a potentially mineable mineral resource and are used to generate a

mine design and LOM production plan.

Woxna supplied the following geological block in two different file formats as follow:

vie_voxna_2013apr_75.bmf (Vulcan file format); and

vie_voxna_2013apr_75.csv (CSV file format).

16.3.2 KRINGELGRUVAN BLOCK MODEL

16.3.2.1 GEOLOGICAL BLOCK MODEL VALIDATION

Validated geological models were provided by Woxna, and the information has been assumed to be

accurate by Golder. Golder has not updated or validated this previous work. Golder conducted an

interrogation of the geological block model “vie_voxna_2013apr_75.bmf” solely to understand the

mineralization domains and the variables used to determine the mineral resource estimate. Golder

did not audit Kringelgruvan mineral resource estimate.

Golder was supplied with a 3D block model in Maptek Vulcan™ format for the Kringelgruvan deposit.

The geological model provided and employed in this report is in the SWEREF99 survey reference

system. The geological model provided did not contain air blocks which had been removed.

Manipulation of the geological model was required to prepare the model for input into Whittle software

used for pit optimisation. Golder utilized Datamine™ software for the preparation and export to a

Whittle block model format. The model manipulation steps included:

Import of the supplied geology block model in comma-separated values (CSV) file format into

Datamine and comparison of tonnes and grade to verify conversion;

Creation and population of additional model variables used in the optimisation process; and


Export to Whittle 4 x format (.mod file) and cross check of model tonnes and total graphite

units in Whittle.

Woxna supplied the following densities presented in Table 16-3 to use for waste, graphite and

overburden material.

Table 16-3: Density by Material Type in Geological Block Model “vie_voxna_2013apr_75.csv”

Model Lithtype Attribute Density

t/m3 Material

Bmtu 2.7 Banded meta-stuff

Grf (B type graphite) 2.7 Graphite mineralisation

Hg (A type graphite) 2.7 Higher grade graphite mineralisation

Msed 2.7 Meta-sediment

Ovb 1.8 Overburden

Peg 2.7 Pegmatite

The following sections summarise modelling steps and checks applied to the Kringelgruvan deposit.

16.3.2.2 BLOCK MODEL PARAMETERS

The block model parameters are shown in Table 16-4. The block model is in SWEREF99 survey

coordinate reference system.

Table 16-4: Kringelgruvan Geological Block Model Parameters

X Coordinate

(Easting)

Y Coordinate

(Northing)

Z Coordinate

(RL)

Model origin 531 800 6 808 250 0

Block size (m) 5 25 5

Sub Block size (m) (minimum) 1.25 6.25 1.25

Rotation 90 (Bearing) 0 (Plunge) 0

16.3.2.3 BLOCK MODEL VARIABLES

The key variables required to describe the geology (PEM type) and grade are show in Table 16-5.

The Kringelgruvan geological model does not contain air blocks. The model contains in-situ blocks

(not mined) and mined out blocks.


Table 16-5: Geological Block Model Key Variables

Variable Name Description

Lithtype Banded meta-stuff, graphite mineralization, high grade graphite mineralization, meta-sediment, overburden or pegmatite

C_uncut Carbon

Bd Bulk density

Category Resource category (inf, ind or meas)

Rsc_cat 0 non classified, 1 measured , 2 indicated, 3 inferred, 4 additional mineralisation

Mined In-situ or mined

Volume Block volume

Graphite tonnes Tonnes

16.3.2.4 IN-SITU RESOURCES

The in-situ mineral resource tonnage and grade from the supplied Kringelgruvan Block Model CSV file

are presented in Table 16-6, Table 16-7 and Table 16-8. The Whittle open pit optimisation is based

exclusively on the in-situ Measured and Indicated mineral resources in this resource block model and

excluded blocks mined previously. Note that the confidence in Inferred mineral resources is

insufficient to allow for the meaningful application of technical and economic parameters or to enable

an evaluation of economic viability worthy of public disclosure. The cut-off grade used in the Whittle

optimisation was 7 % graphite.

Table 16-6: Kringelgruvan Geological Model In-situ Tonnes and Grade

Resource category

Tonnage

Mt

Average Graphite

%

Total Graphite

Mt

Waste 568.564 0 0

Measured 1.936 7.2 0.140

Indicated 3.99 6.8 0.270

Inferred 0.954 5.7 0.54

Additional mineralization 1.949 3.3 0.64

Table 16-7: Kringelgruvan Geological Model In-situ Tonnes and Grade Using a 7 % cut-off Grade

Resource category Tonnage

Mt

Average Graphite

%

Total Graphite

Mt

Measured 0.934 10.7 0.099

Indicated 1.784 10.7 0.190


Resource category Tonnage

Mt

Average Graphite

%

Total Graphite

Mt

Inferred 0.379 11.0 0.042

Total Measured and Indicated 2.718 10.7 0.289

Table 16-8: Kringelgruvan Geological Model In-situ Tonnes and Grade Per Mineral Type

Mineral Type Tonnage

Mt

Average Graphite

%

Total Graphite

Mt

Higher grade graphite mineralization (A type) 3.411 10.2 0.346

Graphite mineralization (B type) 5.382 3.4 0.182

Total 8.793 6.0 0.528

16.3.2.5 EXPORT TO DATAMINE BLOCK MODEL

The geological model CSV file format was imported in Datamine software.

As a check the Datamine model tonnes and grades were compared to the original CSV file format.

There is a very minor difference due to re-blocking of the model after the regularization process. The

Vulcan block model reported tonnes and percentage of graphite were presented in Table 16-6.

16.3.2.6 TOPOGRAPHY AND GEOLOGICAL BOUNDARIES

The topographic surface and geological boundaries supplied are included in Table 16-9.

Table 16-9: Topography and Geological Boundaries

File Description

zone1_2m_rt90_25gv_clipped_solid.00t Topography solid

open_pit_aug_solid.00t Existing Kringelgruvan open pit solid

ovb_mar13.00t Overburden

mxatall.00t Higher grade graphite mineralization (A type)

mxbtall.00t Graphite mineralization (B type)

mx_msed.00t Meta-sediment

mx_peg.00t Pegmatite

mx_peg1.00t Pegmatite1


16.3.2.7 VERIFICATION OF DATAMINE BLOCK MODEL

As a check the Datamine model tonnes and grades were compared to the original CSV file format.

There is a very minor difference due to re-blocking of the model after the regularization process. The

Datamine block model reported tonnes and percentage of graphite are presented in Table 16-10.

Table 16-10: Kringelgruvan Datamine Geological Model Tonnes and Grade (no cut-off applied)

Resource Category Volume

(Mm3)

Tonnes

(Mt)

Density

(t/m3)

Graphite

Grade

(%)

Total Tonnes Graphite

(t)

Waste 214.1 568.6 2.66 0.00 -

Measured 0.717 1.94 2.70 7.23 139,946

Indicated 1.48 3.99 2.70 6.78 270,640

Inferred 0.353 0.954 2.70 5.68 54,248

Additional Material 0.722 1.95 2.70 3.27 63,660

Total/Average 217.4 577.4 2.67 0.09 528,494

The Datamine block model reconciled to within 0.46 % of the Vulcan block model tonnes and volume.

16.4 WHITTLE OPTIMISATION

16.4.1 INPUT BLOCK MODEL

The Whittle Four-X optimisation for the Kringelgruvan deposit is based on Measured and Indicated

mineral resources only with no cut-off applied.

The Datamine model vie_voxna_2013apr_75.dm was imported into Whittle and reblocked to (10 m x

12.5 m x 10 m) and (10 m x 25 m x 10 m). Re-blocking in Whittle is used to complete the pit

optimisation with a practical selective mining unit (SMU), which is manifested in the block size used in

the optimisation.

Variables defining the waste, mineral resources tonnes for each block were added to the block model.

In order to run the scenario in Whittle it was necessary to add a new calculated variable into the

model for the total tonnage of the block.

TOTN: Total tonnage of the block = density x volume of the block.

All blocks in the supplied Vulcan block model had a 2.7 t/m3 value as bulk density. The bulk density

value of the blocks coded as “ovb” was therefore changed to 1.8 t/m3 to more accurately represent the

overburden tonnage.


16.4.2 DATAMINE CONVERSION AND DATA

The surface and geological boundaries were used as received, as per the supplied Vulcan model.

16.4.2.1 VERIFICATION OF MODEL ESTIMATES

A manual check was made comparing block values to the CSV block model. Slices were made

through the block model and tonnes and grade were interrogated for several blocks. The final

optimisation block model, vie_voxna_2013apr_75.dm, tonnes reported by material type is included in

Table 16-11.

Table 16-11: Optimisation Block Model Tonnes

Rsc_cat MTonnes

Waste 568.560

Measured 1.936

Indicated 3.990

Inferred 0.954

Additional Mineralisation 1.950

The total Measured and Indicated mineral resources tonnage at no cut-off grade in the model is

5.926 M tonnes.

16.4.3 EXPORT TO WHITTLE

The optimisation block model (vie_voxna_2013apr_75_x_final.dm) was exported from Datamine to a

Whittle block file (vie_voxna_2013apr_75_x_rev1.mod) and imported into Whittle software. A check

of the model quantities is shown in the screenshot in Figure 16-1.

Due to the small and irregular block sizes used to define the geological boundaries, the model was re-

blocked to 10 m x 25 m x 10 m in Whittle to reduce the processing run time. Total tonnes and units

were checked and matched those in Figure 16-1 below.


Figure 16-1: Screenshot of Imported Whittle Block Models Summary for Kringelgruvan

16.4.4 WHITTLE OPTIMISATION PARAMETERS

16.4.4.1 GEOTECHNICAL OVERALL SLOPE ANGLE

An overall slope angle of 55 ° was applied based on the existing Kringelgruvan open pit for a high wall

design incorporating 5 m high benches, 2.5 m wide berms and a 65 ° face angle for all material types.

These angles were applied on both the hanging wall and footwall of the resource for the optimisation.

Ramps were not separately accounted for in the Whittle open pit economics determination and

optimisation.

A sensitivity to design wall slope angles was performed as part of a basic sensitivity analysis in

Whittle. The following two different slope angles cases were run in Whittle:

55 ° overall slope angle; and

65 ° inter-ramp slope angle for first two bench and 55 ° inter-ramp slope angle for the

remaining bench.

The Whittle results indicate that pit is relatively insensitive to slope angle.

The pit slope design is discussed in Section 16.5 below.


16.4.4.2 MINE AND PLANT PRODUCTION SCHEDULE

For the purposes of optimisation, the annual mine production capacity was limited to 100 000 t ROM

for Kringelgruvan.

Subsequent process design has increased production requirements to 155 000 t/a. Golder reviewed

the impact of this and other changed optimisation parameters and determined that mine planning

would not be significantly affected by this change and the mine design was therefore not re-iterated.

16.4.4.3 DILUTION AND MINING RECOVERY

Based on Golder’s experience, the mining modifying factors were applied in Whittle for graphite

dilution and losses. An overall dilution factor of 2.5 % in volume was assumed. The graphite grade of

the diluting material surrounding the mineralised domains was assumed to be zero. The overall

mining recovery was estimated at approximately 97.5 % based on a 2.5 % mining losses assumption.

Golder is assuming the high mining recovery will only be achievable if selective mining practices

combined with sampling and grade control are followed.

16.4.4.4 COSTS

Cost assumptions were supplied by Woxna, GBM and Golder. In Whittle optimisation any

expenditure that would stop if mining stopped must be included in the cost of mining, processing or

selling graphite.

The costs for this study were developed considering; direct mining costs, processing costs, selling

costs, and overhead or time dependent costs.

16.4.4.4.1 DIRECT MINING COSTS

Direct mining costs relate to the cost per unit of production such as load and haul, and blasting. A

combined mining cost per tonne was provided by SMI and listed in Table 16-12 for each material type

mined.

Table 16-12: Direct Mining Costs Assumptions

Material type SEK /t USD /t

Overburden 35 4.50

Waste 35 4.50

Graphite 48 7.20

Vertical Mining Cost Component N/A N/A


16.4.4.4.2 PROCESSING COSTS AND RECOVERY

The following assumed values for graphite plant recovery and operating cost at the processing plant

were provided by GBM:

Processing plant graphite recovery 85 %; and

Plant and infrastructure operating expenses (inclusive of general & administration, and selling

costs) 70.2 USD/ROM tonne.

16.4.4.4.3 INDIRECT OR TIME RELATED COSTS

Extra costs which result from extending the life of the mine were not accounted for. Time costs

include site overhead costs such as inductions, training, supervision, workshop, offices, technical

services, light vehicles, and other similar associated costs.

Woxna did not supply an annual indirect cost estimate therefore no cost was included in Whittle for

Kringelgruvan deposit.

16.4.4.4.4 SELLING COSTS

Selling costs are those costs that are incurred when the product is sold. In this particular case it

applies payment of royalties. A 0.2 % royalty was provided by GBM for use in the optimisation and

was converted to a selling cost per unit in Whittle.

16.4.4.5 SELLING PRICE

An average graphite sale price of 1 500 USD/t was provided by GBM based on the mineralogy,

expected graphite recoveries and current market prices.

16.4.4.6 DISCOUNT RATE

A 10 % discounting rate supplied by Flinders Resources was applied in Whittle to calculate the net

present value of the optimisations. Results are reported as discounted net value for the deposit

assessed.

16.4.5 WHITTLE .PAR FILE

A Whittle optimisation parameter file (.par) was developed to run the Whittle optimisation. The PAR

file defines the various whittle inputs developed and determines how the Whittle optimisation uses

them in the optimisation process.

The PAR file includes information regarding the number of optimisation iterations and how they are

calculated. The number of optimisations is controlled by the revenue factor (RF), which was

calculated over the range of 0.5 to 1.2 (at an increment of 0.01). The RF is used to determine pit limit


sensitivity to graphite price. The RF range approach is a simple way to scale values to produce a

range of optimal pit shells for different graphite prices. This is a powerful tool for mine planning as

commodity prices generally are subject to a high degree of uncertainty over time.

The revenue factor range is relative to the defined base graphite price, i.e. RF 1.0 = 1 500 USD/t.

For example RF 0.5 uses a graphite price of 750 USD/t and RF 1.2 is using a graphite price of

1 800 USD/t.

16.4.6 WHITTLE OPTIMISATION RESULTS

Whittle optimisation software generates a series of nested pit shells by varying the Revenue Factor

parameter based on a set of financial and other parameters such as costs and graphite price. This

section gives a summary of Kringelgruvan optimisation results.

A summary of the pertinent Whittle pit shells generated by revenue factor are presented in

Table 16-13 and illustrated in Figure 16-2. In the absence of a determination of feasibility, the term

Potentially Economic Material (PEM) is used to describe material that may in future be deemed to

have an economically viable grade of mineralisation. This material is also referred to as ROM.

Table 16-13: Kringelgruvan Whittle Pit by Pit Summary of Results

Pit Number Revenue

Factor

Discounted Net Cashflow

(USD M)

PEM Processed

(Mt includes dilution)

Waste

(Mt)

Cut-off C

(%)

8 0.85 21.8 1.3 4.9 7.0

11 1.00 19.6 1.9 10.8 7.0


Figure 16-2: Whittle Resultant Pits (RF 0.85 – Magenta, RF 1.0 – Blue)

The optimisation results are presented graphically in Figure 16-3 pit tonnage waste and mineral

versus revenue factor and discounted net value. Figure 16-4 for pit value and mineral tonnage versus

the revenue factor and discounted net value.

Figure 16-3: Graph of Kringelgruvan Graphite and Waste Tonnage Versus Revenue Factor and Discounted Net Value


Figure 16-4: Kringelgruvan Graph of Results Pit Discounted Net Value and Tonnes of Graphite by Revenue Factor

16.4.6.1 ANALYSIS OF RESULTS

The highest value pit generated is the RF 0.85 pit, Pit 8. Pit 8 generates 1.3 Mt mineral resource

including dilution and 4.9 Mt of waste with a discounted net value of USD 21.8 M.

The RF 1.0 results in Pit 11 which generated 1.9 Mt of mineral resource including dilution and 10.8 Mt

of waste with a discounted net value of USD 19.6 M.

16.4.6.2 SENSITIVITY

The sensitivity of the discounted net value generated for the Pit 11 shell to slope angle, cut-off grade,

processing recovery rate and the graphite price are shown in Figure 16-5.

The graph shows that Kringelgruvan pit value is relatively insensitive to slope angle and cut-off grade.

The pits were relatively sensitive to the graphite selling price and process recovery rate. A 10 %

decrease in the graphite price results in a change in graphite inventory of approximately 40 Kt.


Figure 16-5: Kringelgruvan Spider Graph Sensitivity of Pit Value to Variance of Costs and Graphite Price

16.4.7 WHITTLE PIT SELECTION

The highest value pit shell may not necessarily be the shell chosen for the pit design as a number of

other criteria are generally considered when choosing pit shells as a basis for stages and the final pit

designs.

The RF 1.0 pit detailed in Table 16-14 was selected by Golder for the purposes of the PEA study and

forms the basis of the smooth pit designs and the Kringelgruvan production schedules.

Table 16-14: Selected Pit Shell and Whittle Results for Kringelgruvan Deposit

Pit Number Revenue

Factor Discounted Net Cashflow

(USD M) PEM Processed

(Mt includes dilution) Waste (Mt)

Cut-off C(%)

11 1.00 19.6 1.9 10.8 7.0

The RF 1.0 Kringelgruvan Whittle pit contours are illustrated in Figure 16-6. The RF 1.0 optimisation

resulted in the generation of two distinct pits, the East Pit and the West Pit.


Figure 16-6: Kringelgruvan RF 1.0 Whittle Pits

The smooth pit design employed the Whittle outlines as guidance for the bench outlines and ramp

locations. The open pit design is described in Section 16.6.1 below. The geotechnical criterion used

in the mine design is detailed in Section 16.5 below.

16.5 PIT SLOPE DESIGN CONCEPT

16.5.1 GENERAL

A key objective for hard rock open pit mine slope design is to determine the most stable maximum pit

slope angles that provide an acceptable level of safety whilst minimising waste stripping and

maximising financial return. It is good practice to continuously monitor pit wall stability as the pit

development progresses and to implement preventive or remediation measures to detect and manage

slope instability through design and operating standards and procedures.

This section gives an overview of the terms typically used in the design of pit slope geometries as well

as key geotechnical and mining pit slope design factors.

16.5.2 PIT SLOPE CONFIGURATIONS

Figure 16-7 illustrates the pit slope terminology that is used throughout this report to describe

benches and haul road ramps and the geometric arrangement of the pit walls.

West Pit

East Pit


Figure 16-7: Inter-relationships Between Bench Geometry, Inter-ramp Slope Angle, and the Overall Slope Angle (Wyllie & Mah, 2004)

The geometry illustrated in Figure 16-7 of the open pit wall is influenced by slope angle, inter-ramp

angle, and the bench geometry. Note that the bench face angles are defined between the toe and

crest of each bench, whereas the inter-ramp slope angles between the haul roads/ramps are defined

by the line of the bench toes. The overall slope angle is always measured from the toe of the slope to

the topmost, or daylight, crest.

16.5.2.1 BENCH GEOMETRY

The bench height is determined by the reach of the excavator chosen for the mining operation and to

minimise mining dilution and maximise recovery. The bench face angle is usually determined to

reduce, to an acceptable level, the amount of material that will likely fall from the face or crest and

remain stable over the required life of the project. The bench width is sized to prevent small wedges

and blocks from the bench faces falling down the slope and posing a hazard to men and equipment.

Figure 16-8 illustrates the bench geometry that results from the bench face angle and bench width

which will ultimately dictate the inter-ramp slope angle. In order to steepen the inter-ramp slopes, and

minimise the excavation of waste rock, the pit slopes at Kringelgruvan will incorporate double

benches of 10 m in height.


Figure 16-8: Catch Bench Geometry (originally from Call, 1986) (Kuchta & Hustrulid, 2006)

16.5.2.2 INTER-RAMP SLOPE

The maximum inter-ramp slope angle is dictated by the bench geometry. However, it is also

necessary to evaluate the potential for multiple bench scale instabilities due to large-scale structural

features such as faults, shear zones, bedding planes, foliation etc. In some cases, these persistent

features may completely control the achievable inter-ramp angles and the slope may have to be

flattened to account for their presence.

Other factors that may reduce the overall slope angles are things such as rock mass strength,

groundwater pressures, blasting vibration, drill and blast precision, stress conditions and mine

equipment requirements.

16.5.3 PIT SLOPE DESIGN FACTORS

Golder has not performed any stability analysis on the open pit design. A geotechnical analysis

should be conducted to determine safe design parameters for the ultimate open pit slopes wall

geometries. Development of slope design criteria involves analysis of predicted failure modes that

could affect the slope at bench, inter ramp and overall scales. The level of stability can then be

assessed against the acceptance criteria, which is usually specified by the owner-operator and/or

regulator in terms of safety and economic risk.

It should be recognised that in stronger rocks failure is likely to be structurally controlled and that in

weaker rocks the rockmass strength is a controlling factor down to the bench scale.

The process of slope design conventionally starts with the development of a geotechnical model for

the proposed pit which is subdivided into geotechnical domains. The domains are rock types with


similar characteristics in terms of geology, structure and material properties. For each domain the

potential failure modes can be assessed and designs at the bench, inter-ramp and overall scale are

based on the acceptance criterion be it ‘Factor of safety’ or ‘Probability of failure’.

16.5.3.1 SITE HYDROGEOLOGICAL CONDITIONS

High groundwater pressures and water pressure in tension cracks will reduce rock mass shear

strength and can adversely impact slope stability. Depressurization programs can reduce water

pressure behind the pit walls and allow steeper pit slopes to be developed in the ultimate pit slope

walls.

To the south east of the proposed mine site is the Östermyrorna, which is classified as national

interest nature conservation area (kommun, 2012) and slightly encroaches on the Kringelgruvan

mining area. The Östermyrorna area is ecologically sensitive and described as follows: “multiform

wetland area with the value of wetland complexes, swamp forest, slightly domed bog, made at

waterways, soli genic and top geological marshes. Much of the area is limestone and has a rich

flora.” (Blonde, 1998). Figure 16-9 shows an overview map of the Ovanåkers kommun is presented

which illustrates the relative location between Kringelgruvan area (R22) and Östermyrorna area (R2).

The presence of free water contiguous to the open pits presents some operating risk to the pit wall

stability. The site hydrogeologic conditions need to be investigated in greater detail and used as an

input into the pit design.


Figure 16-9: Ovanakers Kommun Overview Map

16.5.3.2 STRUCTURAL GEOLOGY

The orientation and cohesive strength of major, continuous geological features such as faults, shear

planes, weak bedding planes, structural fabric, and/or persistent planar joints will strongly influence

the overall stability of the pit walls.


Drill core and geotechnical logging will help in assessing and determining rock fractures and intact

mechanical rock properties. Additional data relating to the geometrical properties of the

discontinuities should also be collected as part of a geological site investigation.

16.5.3.3 ROCK MASS STRUCTURE

The ultimate inter-ramp slope angle is a function of individual benches geometries stability as well as

the orientation, strength, and persistence of smaller scale structural features such as joints. Rock

mass structure is also controlled by:

Intact properties of both mineral resources and waste material (modulus and uniaxial

compressive strength); and

Mine geometrical properties (ramp specification, bench and berm).

16.5.3.4 LITHOLOGY AND ALTERATION

Graphite lithology has been logged and geological grade domains have been modelled in Vulcan to

best represent the margin of the graphite mineralised bodies during mineral resource estimation.

Grouping rock masses within different geological domains according to geotechnical characteristics

and level of alteration of their rock types is a key tool in the evaluation of the stability of the ultimate pit

walls.

16.5.3.5 ROCK MASS STRENGTH

Rock mass strengths are typically estimated using intact rock strength and rock mass classification

schemes such as the Rock Mass Rating (RMR) system. Shallower overall slope angles may have to

be considered if the graphite or waste materials are of a lower rock mass quality.

16.5.3.6 BLASTING PRACTICES

Production blasting around the walls can considerably affect interim and final pit walls. The blast-

induced disturbance damages rock walls and reduces the effective strength of the rock mass.

Controlled blasting programs near the final wall can be implemented to reduce blasting induced

disturbances and allow steeper slopes. Good safety practice at the mine site should include scaling

of blast induced fracturing to minimise rockfalls and local failures.

16.5.3.7 ROCKMASS PROPERTIES

The host rock is in metasediments (Reed, 2012). Laboratory tests of pegmatite granite at Forsmark

(Jacobsson, 2007) in Sweden show a uniaxial compressive strength of 167 MPa. The estimated

uniaxial compressive strength is estimated to be between 100 MPa and 200 MPa. In order to

determine more accurate value for uniaxial compressive strength, tests on rock samples are required.


The pegmatite granite has a density of 2.62 kg/m3, Young modulus of 70 GPa and poisson ratio of

0.29 (compressive).

16.5.3.8 ROCK STRESS CONDITIONS

Mining induces stress changes due to lateral unloading around the walls of the open pit. Stress

release can negatively affect the quality of the rock mass and increases in slope displacements.

Localised stress decrease can reduce confinement and result in an increased incidence of ravelling

type failures in the walls. Modifying the mining arrangement and sequence can sometimes manage

these stress changes to enhance the integrity of the final pit walls.

No stress measurements have been performed at the Kringelgruvan mine site. The World Stress

Map (WSM) has an online predefined stress field map of the Earth’s crust based on the WSM

database 2008 available for the Nordic region (Heidbach, Tingay, Barth, & Reinecker, 2008). The

WSM displays the orientations of the maximum horizontal compressive stress SH. The length of the

stress symbols represents the data quality, with A being the best quality. Quality A data are assumed

to record the orientation of SH to within 10 ° to 15 °, quality B data to within 15 ° to 20 °, and quality C

data to within 25 °. Quality D data are considered to give questionable tectonic stress orientations

(Zoback, 1992; Sperner et al., 2003; Heidbach et al., 2007).

The WSM database has some stress field measurement nearby the Kringelgruvan mine site, which

shows the maximum horizontal stress has a direction of 352 ° in Azimuth and 38 ° for plunge, the

medium stress has an azimuth of 156 ° and 51 ° for plunge. Besides the WSM, there are some

equations show the variation of horizontal stresses for Fennoscandia (Stephansson, 1993) for

example the following equations:

10.8 0.037

5.1 0.029

0.8 0.02

In which , and are major, intermediate and minor stresses in MPa, and represents the depth

in meters. An overall slope angle of 55 degrees was therefore preliminarily assumed based on the

foregoing data.

16.6 OPEN PIT MINING

16.6.1 OPEN PIT MINE DESIGN

The ultimate open pit mine design layout for Kringelgruvan was generated based on the Whittle

optimisation RF 1.0 pit shell, the geotechnical parameters discussed in Section 16.5 above and the


operating ranges of the proposed mining mobile plant. The key inputs to the open pit mine design are

summarised in Table 16-15.

Table 16-15: Open Pit Design Parameters

Parameter Units Value

Bench Height m 10

Batter Angle Degrees 78 °

Berm Width m 3.0

Design Slope Angle Degrees 55 ° – 65 °

Ramp Width m 10

Ramp Gradient 1:10 (10 %)

Final Slope Angle Degrees Variable

Minimum Mining Width m 15 - 20

The final pit designs were completed using the parameters listed in Table 16-15. The pits are

accessed by ramps that connect to the current haulage routes to the waste tip areas and the

processing plant. The two pits are illustrated in Figure 16-10.

Figure 16-10: Final Kringelgruvan Pit Designs

West Pit

East Pit


The 10 m wide haulage ramps were designed in the geological hanging-wall of the deposit because

the previous operators on the site had experienced problems maintaining the ramp in the footwall of

the mineralisation. The haul ramps were decreased to 6 m wide to access lower benches. All of the

ramps were designed to be within the Whittle RF 1.0 shell to minimise the amount of additional waste

stripping and associated mining costs. Whilst not included in the mining schedule, some PEM in the

ramp can be recovered by retreat mining up the ramp.

The final pit designs were quite close to the outlines of the selected Whittle shells. The two designs

as shown in Figure 16-10 are illustrated with the RF 1.0 shells in Figure 16-11 below. A bench by

bench comparison is shown in Figure 16-12.

Figure 16-11: Final Pit Designs Compared with the RF 1.0 Whittle Shells

West Pit

East Pit


Figure 16-12: Comparison of the Final Pit Design with the RF1.0 Whittle Shell (207.5 m Elevation) in Plan

Typical sections through the East and West pits comparing the final design to the RF 1.0 shells are in

Figure 16-13. Due to the method used to generate the slopes in Whittle, the final pit designs are

steeper than the Whittle shells. In the Whittle runs the slopes were set at 55 ° for the first two

benches and then at 65 ° to the pit bottoms.

532950 East (East Pit)

West Pit

East Pit


532300 E (West Pit)

Figure 16-13: Comparison of the Final Pit Design on Section (as labelled)

The total mined resource included in the pit designs is tabulated by pit in Table 16-16.

Table 16-16: Summary of Total In-pit Resources

Source PEM

kTonnes Grade Graphite

Waste kTonnes

Total kTonnes

Waste: PEM Ratio

East Pit 889 706 10.54 3 757 246 4 646 952 4.2

West Pit 1 009 033 11.4 5 841 597 6 850 631 5.8

Total Inventory 1 898 739 10.98 9 598 843 11 497 583 5.1

The pit design tonnages and grades compare favourably with the RF 1.0 Whittle shell results,

recovering 85.4 % of the PEM tonnes and just 81 % of the waste included in the Whittle shells.

The distributions of the mineable tonnage by bench in two pits are illustrated in Figure 16-14. The low

grade tonnages indicated in Figure 16-14 is material with a grade between 5 % and 7 % graphite.


Figure 16-14: Distribution of Tonnage by Bench in the West (left) and East (right) Pit Designs

16.6.2 MINE PRODUCTION SCHEDULING

The mine production schedule was developed in a set of steps looking to determine an efficient start-

up of mining operations. A key objective of the schedule was to limit the amount of waste stripping in

the early phases of mine development to reduce the capital expense of re-starting the mine.

The mine was scheduled initially by quarters for 2 years and then in years for remainder of the LOM.

The target production was 100 000 t of feed to the processing plant per annum, with the feed coming

from both the east and west pits simultaneously.

The mine production schedule was developed in four phases. Years 1 and 2 were scheduled using

selective mining units (SMUs) 25 m x 15 m by 5 m high (equivalent to 5 062.5 t). The blocks were

scheduled in monthly periods producing and average of 8 333 t of ROM material per month. The

second phase of scheduling was year three, when the SMUs were scaled up to 100 m by 30 m by 5

m (20 250 t) and scheduled in quarters producing approximately 25 000 t of PEM per quarter. After

Year 3 through to Year 10, the mine was scheduled by years producing 100 000 t of PEM per annum.

The remainder of the LOM was scheduled in 5 year increments to a LOM that was just over 18 years.

The Mine waste was hauled to waste tips above the West Pit and in the vicinity of the existing waste

tips above the East Pit.

16.6.2.1 MINE OPENING DEVELOPMENT

The mine production schedule includes the opening cuts and re-start of mining operations in the East

Pit as well as the start of mining in the West Pit. The start of mining operations was driven by

maintaining a relatively low strip ratio. This was achieved by using a low revenue factor Whittle shell

as a starting point for the mining. Golder used the RF 0.7 pit for the opening cut design to minimise


the strip ratio and to start mining in relatively higher value ground. The RF 0.7 pits are illustrated in

Figure 16-15.

Figure 16-15: RF 0.7 Whittle Shells Used as the Starting Point for Mining in the East and West Pits

16.6.2.2 YEAR 1 – 2 MINE ADVANCE

The mine stripping and waste development starts against the north highwall in both of the pits and

progresses west and down, bench-by-bench to expose the graphite mineralisation. The East Pit is

mined along the north wall to expose the final pit wall that is coincident with both the RF 1.0 and

RF 0.7 pits. The West Pit is developed in a similar manner starting in early Year 2. The end of year

status maps are shown in Figure 16-16.

Year 1

Year 2

Figure 16-16: Kringelgruvan Status Map – Up to End of Year 2

West Pit

East Pit Starter Pits

N th


The mine produces the target PEM tonnage of 100 000 t in both years 1 and 2. The low production

rate and minimised waste stripping enable the fast plant feed ramp up and maximised process plant

utilisation. The monthly, for Year 1 and Quarterly production schedule with the rolling waste to PEM

ratio is charted in Figure 16-17.

Figure 16-17: Kringelgruvan 2 Year Mining Schedule

The pits are relatively high strip ratio pits at 4.4 and 6.1 for the East and West pits respectively.

Whilst the overall strategy to minimise the waste transport in the early years is achieved there are still

requirements for exposing the PEM on the lower benches. This has resulted in a variable strip ratio

over short periods but the overall stripping, on a larger timescale is consistent with the objective of

minimising waste generation and the stripping costs. The total production in years one and two are

summarised in Table 16-17. The Graphite production is the processed recovered tonnage of product

at a recovery of 85 % graphite from the ROM feed.

Table 16-17: Summary of the Year 1 - 2 Production Schedule

Item Units TOTAL

Waste (inclusive of ob) Tonne 798 483

Waste (exclusive of ob) Tonne 753 332

Graphite Tonne 17 879

A-type graphite Tonne 17 630

C-type graphite Tonne 249

Strip ratio (waste:ROM) Waste Tonne/ROM Tonne 4.0


Item Units TOTAL

ROM tonne Tonne 200 000

Total tonne Tonne 960 363

The mine advance is illustrated in Figure 16-18 for Year 1 and Figure 16-19 for Year 2. In Year 1 the

East Pit is developed from the existing topography down to the 220 m elevation. The waste dump

rises from the existing topography nominally at the 230 m elevation to a level surface at the 245 m

elevation.

Figure 16-18: End of Year 1 Status Map

In Year 2 (Figure 16-19) the pit floor in the East pit is at approximately the 215 m bench and in the

West pit, the development has reached the 215 m elevation. The main waste dump has risen to the

250 m elevation during this period.

240 m 220 m

245 m



16.6.2.3 YEAR 3 MINE ADVANCE

Golder continued the scheduling through to Year 3 with the results charted in Figure 16-20 and

tabulated in Table 16-18. The total PEM production is 300 000 t over the three years with 1.3 Mt of

waste for a project to date strip ratio of 4.4. There are 26 500 t of graphite produced in the first three

years of production.

220 m 215 m 215 m 225 m

250 m


Figure 16-20: Kringelgruvan Production Schedule to Year 3

Table 16-18: Kringelgruvan Schedule to the End of Year 3

Item Units TOTAL

Waste tonne (inclusive of ob) Tonne 1 309 319

Waste tonne (exclusive of ob) Tonne 1 230 832





ROM tonne Tonne 300 000

Total tonne Tonne 1 585 648

The strip ratio in Year 3 is marginally higher and continues to increase somewhat to account for

deferred waste mining in Year 1 and 2 of the plan.

Mine development in Year 3 is mapped in Figure 16-21. The East pit is well established by this time

and the opening cut in the West Pit progressed on the 215 m bench further to the west. The waste

dump crest is at 255 m.



16.6.2.4 YEAR 10 MINE ADVANCE

The 10 year schedule continues mining in both the East and West pits. At the end of Year 10,

1 M tonnes of PEM have been mined and 85.3 kt of graphite had been produced. The waste moved

is 5.6 M tonnes. The PEM to waste ratio increases in this period from 4.4 to 5.6 due to the increased

depth of mining. The 10 year schedule is charted in Figure 16-22 and summarised in Table 16-19.

255 m

215 m 230 m

215 m225 m


Figure 16-22: Kringelgruvan 10 Year Production Schedule

Table 16-19: Kringelgruvan 10 Year Mine Production Schedule

Item Units TOTAL







ROM tonne Tonne 1 000 000


Mine development (Figure 16-23) between years 3 and 10 has almost completed the East Pit, which

has three benches remaining. The West Pit continued to expand to the west along the final north

wall. The waste dump had raised a total of 55 m from the toe to 285 m elevation.



16.6.2.5 LOM SCHEDULE

The LOM extends to approximately 19 years at the proposed production rate of 100 000 t/a. Both of

the pits are mined out and the waste tip is fully stacked. The waste to PEM ratio starts to decrease in

the later years because the benches are increasingly PEM-only or contain a high proportion of

graphite per bench. The LOM schedule is charted in Figure 16-24.

205 m

210 m 220 m

285 m


Figure 16-24: Kringelgruvan LOM Schedule

The total PEM and waste mined over the 19 years equate to just over 9.6 M tonnes of waste and 1.9

M tonnes of PEM producing 166.2 kt of graphite. The final totals of mined material are summarised in

Table 16-20.

Table 16-20: Kringelgruvan LOM Production Summary

Item Units TOTAL



Overburden Tonne 826 173



C-type graphite Waste Tonne/ROM Tonne 448

Strip ratio (waste:ROM) Tonne 5.1

ROM tonne Tonne 1 898 739


The final status map of the site is shown in Figure 16-25.


Figure 16-25: End of Mine Status Map

Figure 16-26: Site Overview

170 m 180 m

170 m

200 m 190 m

315 m


16.7 OPEN PIT MINING OPERATIONS

16.7.1 DRILLING AND BLASTING

Explosives will be used to fragment both the waste and mineralised rock. The blasted material will

then be sized and crushed to feed the process plant. The typical geometry of a drill and blast pattern

is illustrated in Figure 16-27.

Figure 16-27: Open Pit Drill and Blast Geometry (courtesy of DynoNobel.com)

The commercial explosive used for blasting will be an emulsion due to its relatively low cost, ease of

loading and safety. Typical emulsion product is a mixture of ammonium nitrate acting as an oxidiser

and diesel fuel oil in a stabilised liquid state with a gassing agent. The product density ranges

between 1.15 g/cm3 to 1.25 g/cm3. The drill and blast parameters are listed in

Table 16-21.

Table 16-21: Drill and Blast Parameters


Rock Density (wet) g/cc 2.80

Bench Height, BH m 5

Explosive Diameter, D mm 64



Explosive Density g/cc 1.20

Explosive Energy, AWS j/g 3 290

Burden, B (25-40D) m 2.1

Burden Stiffness (2.0<BS<3.5) BH/B 2.4

Spacing (=1.15B), S m 2.4

Stemming Length (0.7B), SL m 1.5

Energy Distribution % 70 %

Sub-Drill (BH x 3-15D), SD m 0.4

Blasthole Length, L m 5.4

Explosive Length, C m 3.9

Explosive Loading Density kg/m 3.86

Explosive Weight kg/hole 15.1

Explosive Energy kJ/hole 49 679

Volume Shot, V bcm/hole 25

Mass Shot, T t/hole 70

Powder Factor, PF kg/bcm 0.60

Powder Factor, PF kg/t 0.22

Energy Factor kJ/t 710

Control blasting of the final walls should be conducted as a matter of good mining practice. The

drilling and blasting pattern for pre-split blasts are based on the geometry of the production blasts

summarised in Table 16-21. The pre-split blasts are drilled using a 1.00 m burden and 0.77 m

spacing from the last production row on the bench. A typical blast plan and design is shown in

Figure 16-28.


Figure 16-28: Typical Blast Pattern Design

Loading of explosives will be undertaken using a custom built bulk pumping truck vehicle. Blasts will

be fired using a non-electric down-hole delay initiation system. Production blasts patterns will be

blasted at fixed times during the day and no later than 6:00 pm. Notification of blasts will be

performed as per Swedish regulations and an alarm will always sound prior to production blasts

several minutes in advance.

16.7.2 GRADE CONTROL

Poor grade control of PEM during mine operations can contribute to dilution and decreased mineral

recovery and consequently will reduce the head grade of the mill feed and displace economic material

in the processing stream.

PEM grade control guidelines are provided below:


16.7.2.1 PRIOR TO BENCH PRODUCTION

Geological mapping and sample collection on production benches. Chip or channel samples

can be taken along benches and walls, and blasthole cuttings can be sampled and analysed

in a lab. Additional check samples can be taken from the blasted muck piles;

Monitor and map blast holes during drilling to identify the waste-PEM contact locations; and,

The block model bench should be reconciled with the field observations listed above.

16.7.2.2 DURING PRODUCTION

Visit production benches daily to monitor for dilution and estimate the ROM grade. All oversize

material should be stored in a re-handle area until a geologist can assess if it is PEM or waste, and

operations personnel should periodically collect grab samples, either from an excavator bucket or

from a storage/re-handle area. The excavator operator can be supplied with sample bags and tags,

and assign a prescribed percentage of the production to be sampled (i.e. one sample for every ten

buckets).

16.7.2.3 STOCKPILING STRATEGY

All of the ROM feed reports to a ROM pad situated adjacent to the processing plant. The stockpile is

designed to hold approximately one week of feed material for the processing plant. In the winter

months the pad live stockpile should be reduced to minimize the risk of parts of the pile freezing.

Low grade PEM could be stockpiled to be processed at a later date in the event of open pit production

issues or to be blended with very high grade PEM. Further analysis of the economics of storing and

re-handling low grade material should be conducted as part of the operating strategy of the mine and

processing facilities.

16.7.3 MINING EQUIPMENT SELECTION

Mining the Kringelgruvan deposit is envisioned as an open-pit mine using haul trucks, a hydraulic

shovel, and a front-end loader operated by a mining contractor. A wheel loader will be employed at

the ROM pad area where the machine will make frequent moves between the stockpiles and the

crusher, feeding to the processing plant.

Golder has estimated the productivity of the proposed excavator fleet for both PEM and waste

handling. The production estimate inputs and results are found in Table 16-22.


Table 16-22: Kringelgruvan Prime Loading Unit Productivity Estimate

Parameter Units Cat 385 Exc Cat 385 Exc Cat 938 FEL Cat 938 FEL

Bucket Size m3 3.90 3.90 2.30 2.30

Fill Factor 0.90 0.90 0.90 0.90

Material PEM Waste PEM Waste

Bulk Density t/m3 2.70 2.70 2.70 2.70

Swell 1.30 1.30 1.30 1.30

Bucket Load BCM 2.70 2.70 1.59 1.59

Bucket Load Tonnes 7.29 7.29 4.30 4.30

Nominal Truck Payload Tonnes 40 40 40 40

Calc. Passes to Fill 5.5 5.5 9.3 9.3

Use Passes to Fill 5.0 5.0 9.0 9.0

Calc. Truck Payload 36.5 36.5 38.7 38.7

Load Factor 91.1 % 91.1 % 96.7 % 96.7 %

Time Per Pass Minutes 0.50 0.50 0.75 0.75

Load Time Minutes 2.50 2.50 6.75 6.75

Spot Minutes 0.75 0.75 0.75 0.75

Load + Spot Minutes 3.25 3.25 7.50 7.50

Efficiency Minutes/hr 50 50 50 50

Propel Factor 0.95 0.95 0.95 0.95

Presentation Factor 1.00 1.00 1.00 1.00

Productivity t/hr 533 533 245 245

Productivity BCM/hr 197 197 91 91

Scheduled h/a Hours 5,280 5,280 5,280 5,280

Mechanical Availability 85.0 % 85.0 % 85.0 % 85.0 %

Use of Availability 75.0 % 80.0 % 75.0 % 80.0 %

Utilisation 63.8 % 68.0 % 63.8 % 68.0 %

Operating h/a Hours 3 366 3 590 3 366 3 590

Production/Year Tonnes 1 793 172 1 912 717 824 859 879 850

Production/Year BCM 664 138 708 414 305 503 325 870

The selected equipment is capable of excavating the required tonnages to maintain steady operation

and production from the mine. The mobile plant estimates are based on expected productivities for

similar mining operations. Because the mine is contractor operated, the mining contractor is expected

to select, use and dispatch an appropriate fleet to complete the mining tasks. The contractor fleet

may vary from what Golder has proposed in the following tables.

Primary mine production is achieved using the equipment listed in Table 16-23. A swell factor of

30 % was assumed for both graphite and waste material when it was blasted and excavated.


Table 16-23: Preliminary Mine Equipment List

Equipment Description Quantity

Hydraulic shovels CAT 385 – 5.54 LCM 1

Front-end Loaders CAT 938 – 2.3 m3 1

Rear-dump Trucks CAT 771 – 40.8-Tonne Payload 3

Rotary drills Sandvik D25KS – 172-mm,124-kN Pull-down 1

Bulldozers CAT D6 – 112 kW 1

Graders CAT 12M – 129 kW 1

Water tankers 9 500-liter – 283-kW Truck 1

Service/tire trucks 4 000-kg Crane – 15 000-kg GVWR Chassis 1

Bulk trucks 272 kg/min (26.8 m3) hopper 1

Light plants 9-m Trailer Mounted, 4 x 1000 Watt Flood 3

Pumps 200-gpm, Centrifugal Trash Pumps, 4 kW 2

Pickup trucks 1-ton 4x4 2

Material will be drill and blasted using a bulk emulsion product. The loading of bulk explosives into

blast holes will be conducted using a bulk mixing and pumping truck. One bulk powder truck will be

required for the mining operations at Kringelgruvan. A single drill rig will be required for drilling

operations.

16.7.4 MINE WORKFORCE AND ORGANISATION

16.7.4.1 MINING WORKFORCE

The mining will be conducted by a mining contractor retained to operate and maintain the production

and service equipment. Golder has estimated the number of mobile plant units (Table 16-23) and

used the fleet size to estimate the number of operators and maintenance staff on each shift required

to sustain the operation of the mine. The estimated workforce each shift is summarised in

Table 16-24.

The mining contractor will be tasked with managing both the equipment fleet and workforce in order to

achieve the planned production targets at the budgeted unit costs. The mining contractor planned

workforce may vary from Golder’s estimates of the required labour needed to execute the mine plan.

Table 16-24: Mining Workforce Estimate

Equipment Quantity Operators per Shift

Hydraulic excavators 1 1

Front-end Loaders/Bulldozer 1 1


Equipment Quantity Operators per Shift

Rear-dump Trucks 2 2

Rotary drills 1 1

Bulldozers 1 N/A

Graders 1 1

Water tankers 1 N/A

Service/tire trucks 1 N/A

Bulk trucks 1 N/A

Light plants 3 N/A

Pumps 2 N/A

Pickup trucks 2 N/A

Total 17 6

To determine the size of the contractor crews required to operate the mine Golder assumed that two

crews working eight hour shifts seven days a week would operate the mine. To allow two working

crews and two crews on a rest cycle will require a total workforce of 24 employees, not including

contractor maintenance crews, supervision or management.

16.7.4.2 MINE STAFFING

Golder has estimated the number of full time staff employees required to manage and operate the

mine. The mine Technical Services department is outlined in Table 16-25. The total mine technical

staff is estimated to be 2 in total. Note that additional positions are required for the mine accounting,

health safety and environment and other supporting groups.

Table 16-25: Estimated Mine Staffing Requirement

Position Per shift

Mine Superintendent 1

Chief Geologist 1

Total Staffing 2

The staff on-site as Woxna Graphite employees would be responsible for the daily oversight of the

mine. Other roles such as mine planning and pit slope monitoring would be out-sourced on an as

required basis over the life of the mine.


16.7.4.3 MINE STAFF ORGANISATION

Golder has developed a mine technical services department organogram for Woxna Graphite see

Figure 16-29. The organisation is based on a typical small mining and geology group. The Technical

Services Superintendent (TSS) is responsible for all of the on-site company, consultant and contractor

activities. The TSS liaises directly with their equivalent in the mine contractor organisation and

ensures that planning, design and regular surveying for the site is conducted.

Figure 16-29: Woxna Graphite Mine Technical Services Department Organogram

16.8 WASTE ROCK STORAGE AND MANAGEMENT FACILITY

16.8.1 WASTE CHARACTERISATION

16.8.1.1 NEUTRALISATION POTENTIAL

The Acid Rock Drainage (ARD) generation potential is determined by a static test called Acid Base

Accounting (ABA). The ABA procedure is based on the calculation of the Acid Potential (AP) and the

Neutralization Potential (MPA). According to the current legislation for defining inert mining waste

(SFS2008:722, based on directive 2006/21/EC and 2009/359/EC) the following apply.


Waste shall be considered as being inert waste, within the meaning of Article 3 (3) of

Directive 2006/21/EC, where all of the following criteria, are fulfilled in both the short and the

long term:

Waste will not undergo any significant disintegration or dissolution or other significant

change likely to cause any adverse environmental effect or harm human health;

Waste has a maximum content of sulphide sulphur of 0.1 %, or waste has a maximum

content of sulphide sulphur of 1 % and the neutralising potential ratio, defined as the ratio

between the neutralising potential and the acid potential, and determined on the basis of

a static test EN 15875 is greater than 3;

Waste presents no risk of self-combustion and will not burn;

Content of substances potentially harmful to the environment or human health in the

waste, and in particular As, Cd, Co, Cr, Cu, Hg, Mo, Ni, Pb, V and Zn, including in any

fine particles alone of waste, is sufficiently low to be of insignificant human and ecological

risk, in both the short and the long term. In order to be considered as sufficiently low to

be of insignificant human and ecological risk, the content of these substances shall not

exceed national threshold values for sites identified as not contaminated or relevant

national natural background levels; and

Waste is substantially free of products used in extraction or processing that could harm

the environment or human health.

⁄

The following three steps are performed in order to obtain NP and AP values:

Acid Production (MPA) determination;

Acid Consumption (NP) determination; and

Neutralization Potential Ratio calculation (NPR).

Theoretically, waste rock material possessing NPR values above 1 have no potential for acidification

whereas waste rock with NPR values below 1 demonstrate a risk potential. In practice, a safety factor

is applied and waste material with a NPR value greater than or equal to 3 is generally regarded as

having no acidification potential.

The ABA test does not distinguish between mineralogical differences and the NP values are often

overestimated (Lawrence & Scheske, 1997). The material mineralogy is a very important parameter

influencing ARD generating potential due to different minerals ability or inability to neutralize acid

drainage in different pH range.

The LOM production schedules tracks mine waste NPR generated by bench per phase and is based

on NPR values estimated from the geological block model provided by Woxna.


16.8.1.2 WASTE ROCK CHARACTERISATION

NPR and Sulphur values were interpolated into the geological block model provided by Woxna.

These values could be used to predict the ARD potential generation of the waste material.

The description for the different waste material categories used to classify different ARD generating

levels are:

A: if NP/MPA ratio ≥ 3: unlikely to generate ARD – low acidification potential; and

B: if NP/MPA ratio < 3: likely to generate ARD – high acidification potential.

16.8.1.3 WASTE ROCK MANAGEMENT

For the purpose of this PEA, no waste rock management plan was developed but the LOM production

schedule could enable the development of such a plan as the NPR and sulphur values have been

tracked in the schedule. Such a plan could allow for the potential separation of low potential

generating waste material from the high potential generating waste material in different waste storage

facilities based on the anticipated sulphur and NPR values within the waste rock. As such, waste

material could be categorized and managed selectively based on their ARD generating potential as

follows:

A category: could be stored in a waste dump or used as construction material for the tailings

embankment; and

B category: need to be deposited in a specific waste storage management facility suitable for

preventing and handling ARD generation.

16.8.2 WASTE STORAGE FACILITIES DESIGN AND STORAGE CONSTRAINTS

The following design constraints were assumed for design and development of the rock stockpiles:

Overall face angle of ex-pit rock stockpiles of 33 °;

Natural angle of repose for the waste rock is approximately 38 °;

Nominal height of rock stockpile lift: 10 m;

Catch bench between rock stockpile toe and crest not to be less than 5 m giving 3:1 overall

slope;

In areas where lift height exceeds 10 m, a protocol will need to be developed to ensure safe

working practices. This protocol will need to include monitoring stockpile movement,

development of criteria for cessation of placement, and techniques to safely place over high

banks such as opening and closing sectors of the crest to allow time for the material to

stabilize;

Material swell factor 30 % in the truck;


Material volume assumed to be the same as the bank volume once placed and stored in the

waste dump;

Rock stockpile footprints to be inspected for natural springs, permafrost and should be auger

drilled to evaluate geotechnical conditions prior to placement; and

Offset from ultimate pit high walls a distance equivalent to half of the depth of the pit that

borders the rock stockpiles.

The preliminary dump designs for the project were developed to minimize haulage and associated

costs and store up to 4.9 M loose m3 (LCM) or 10.6 M tonnes of waste. This was accomplished by

taking advantage of topography to the north of the deposit. It should be noted that the extents of the

preliminary dump locations remained inside of the Freehold lease boundaries. The waste dump

configuration and capacity can be changed to reflect any future changes to the project boundary.

The dumps were not scheduled to manage ARD generation. Consideration should be given to the

proper separation and development of potential ARD dumps in a feasibility study. Table 16-26 shows

the waste dump capacity and Figure 16-30 shows the preliminary dump location. Overburden

quantities were estimated and tracked in the LOM schedule.

Table 16-26: Dump Storage Capacity by Elevation

Elevation Area Volume Tonnage

235 1 900 19 004 51 311

245 32 475 324 750 876 825

255 48 629 486 286 1 312 972

265 51 014 510 143 1 377 386

275 52 828 528 275 1 426 343

285 75 588 755 884 2 040 887

295 54 967 549 667 1 484 101

305 30 630 306 295 826 997

315 10 446 104 461 282 045

Total 358 477 3 584 765 9 678 866


Figure 16-30: Location of the Waste Storage Area

16.9 MINING INFRASTRUCTURES

16.9.1 PERSONNEL INFRASTRUCTURE

A special fuel depot will be constructed in the northern part of the industrial area to supply the mining

fleet. The contractors involved with mining activities at the site have the option to construct an

additional fuel depot within the industrial area.

The mining contractor will be self-contained with regard to sanitary waste system since there is

currently no existing facility able to accommodate an increase in the site work force.

16.9.2 ELECTRICAL POWER

The pit shell suggests the existing power line narrowly cuts across the mineralisation on the south

west side of the mining concession and will need to be moved before mining production year 2 or

after 1.6 M tonnes of material have been mined. This is due to the presence of the West Pit ramp

being located on the south wall of the pit.

Waste Dump

East Pit

West Pit


Figure 16-31: Power Line and Pit Interaction

It is believed that additional geotechnical information and mine optimisation and design iterations will

enable an improved ramp location and pit wall design in the vicinity of the power line, thus deferring

its movement until nominally year 4 of production.

Electric power for the open pits is anticipated to be limited to supplying energy to the in-pit dewatering

pumps. There is a power connection to the existing pit.

16.9.3 EXPLOSIVES STORAGE

The commercial explosive used for blasting will be emulsion due to its relatively low cost, ease of

loading and safety.

Detonators, primers, and stick powder will be stored in a magazine complying with the new Swedish

statute for the storage of explosives, MSBFS 2010:5. The explosive storage facility will provide

storage for approximately one week of explosive consumption. The magazine storage for explosives

will be located in a secure location at a distance from other buildings consistent with Swedish

regulations MSBFS 2010:5. All of the blasting practices will comply with Swedish regulations in force

during mine operation.


SECTION 17 RECOVERY METHODS

There is an existing plant at Woxna for processing the graphite PEM. This plant was last used in

2001. It is the intention of the Woxna Graphite Restart Project to utilise existing facilities and

infrastructure where possible.

17.1 HISTORY AND BACKGROUND

The original plant was built in 1996 with trial product delivery in December, 1996. Officially,

production started in March 1997. The plant extension took place in 1998 with an objective to

increase the throughput capacity and the product recovery. A rod mill was installed and an additional

flotation circuit was established. Three new column cells were installed at this time to improve the

efficiency of the fines flotation process.

Additional refurbishment/expansion of the plant took place in 1999, with installation of two additional

regrinding mills in order to increase the number of regrinding/flotation steps to three. To obtain

cleaner material from the rougher flotation, two stage cleaning of the rougher concentrate was

included, where the first two cells were used for cleaning and the four remaining for

rougher/scavenger flotation. An additional three flotation cells were used to float coarse particles

before it went to tailings in order to improve the total recovery. Blending equipment was installed to

make it possible to produce and sell more customised products with regard to grain size and carbon

contents.

In 2000, the feed arrangement to the drier was improved and the dry screening was improved

resulting in both better quantity and quality in the screening process. Automatic sampling was

installed at crucial points in the plant.

Most of the equipment in the comminution and flotation circuits was second hand equipment taken

from other mines.

17.2 BASIS FOR DESIGN

The basis for proposed process is the estimated maximum throughput of the existing rod mill and the

flotation process modelled by Aminpro in June 2013. The crushing plant will be new since the

previous one was the mining contractor’s. The flotation, regrinding and dewatering equipment will be

new. The existing dryer and bagging plant will be used for graphite concentrate together with

additional new bagging plant for the coarse high grade graphite.

The process is shown in the block diagram in Figure 17-1.


Figure 17-1: Process Block Diagram


17.2.1 OPERATING PHILOSOPHY

The crushing plant will be operated 4 hours per 12 hours. The process plant with battery limits from

the silo to the discharge of the dryer will be operated three eight hour shifts per day, seven days per

week.

Major maintenance will be carried out on day shift by contractors and the plant operators will perform

minor maintenance during the rest of the time.

17.2.2 CRUSHING

The crushing was carried out by a contactor and the equipment is no longer available. Options for the

crusher include purchasing and operating or contracting out the crushing operation. GBM proposes

that the crushing plant is owned and operated by Woxna.

The design is based on crushing test work by Metso in 2012. Metso reported that the sample is

abrasive but the crushability is easy. The slippery low-friction nature of the graphite in the material

has been taken into account in selecting the liners. The product from this plant will be 80 % passing

6.35 mm to maximise the throughput of the rod mill.

The ROM material is fed at 70 t/h directly into the ROM bin by a mine haulage truck or from the ROM

stockpile by a front end loader. The ROM has a top size of 600 mm. The ROM is screened by a

vibrating grizzly to remove minus 60 mm particles. The oversize is fed into a jaw crusher where it is

crushed to minus 88 mm. The grizzly undersize and crushed products are transferred by conveyor to

a triple deck vibrating screen where it is screened at 8 mm. The oversize is conveyed to a bin and

then fed by a vibrating feeder to a cone crusher. The crushed product is conveyed back to the triple

deck screen. The crushed product screen undersize, which has 80 % passing 6.35 mm, is transferred

either to the fine ore bin, or using a stacking conveyor stored in an emergency stockpile adjacent to

the crushing plant. Finely crushed product can be loaded by a front end loader from the emergency

stockpile onto the silo feed conveyor via a chute when the crushing plant is down for maintenance.

There is no longer intermediate storage and conveying to the rod mill and therefore, this equipment

requires replacing with either an undercover stockpile or silo to prevent freezing of the crushed

material. GBM proposes a new fine ore bin is installed as part of the new feed delivery equipment with

sufficient capacity that the crushing plant need only be operated for approximately 4 hours every 12

hours.

A new crushing plant with silo is used for the PEA.

17.2.3 MILLING

The rod milling is used to ensure that there is no overgrinding of the graphite. The milling circuit

consists of the existing rod mill and new cyclones. The mill is in closed circuit with cyclones to


produce a product with 80 % passing 275 µm. The throughput of the mill is calculated to be 21 t/h.

The existing mill is 2.7 m diameter and 3.75 m long. It requires a new 360 kW motor to maximise the

throughput.

The crushed material from the bin is fed onto the rod mill feed conveyor by one vibrating feeder at a

controlled rate of 21 t/h. The mass flow is measured by a weightometer on the conveyor belt which

controls the feed rate to the mill. The mill discharge is screened by a trommel with the oversize being

collected in a bin for recycling to the mill feed on a batch basis. The undersize is pumped to cyclones

for classification at 80 % passing 275 µm. The cyclone underflow flows by gravity to rod mill feed box

for regrinding. Water is added to the mill feed to ensure that the density of the slurry in the mill is 80 %

solids. Water is added to the mill discharge to produce the correct slurry density for the cyclones.

The cyclone overflow is pumped to the flash flotation cell.

For the PEA the existing rod mill with new motor will be used. The feed conveyor, cyclone and

cyclone feed pumps and associated equipment will be new.

17.2.4 FLOTATION

The proposed beneficiation process is simpler than used previously. The old process included

multiple flotation stages with regrinding, sizing and gravity separation. The new circuit will use only

flotation with regrinding, comprising of flash/rougher/scavenger flotation with two stage rougher

concentrate cleaning with regrinding and three stage scavenger concentrate with regrinding.

The cyclone overflow is pumped to the flash flotation where high grade coarse graphite is floated.

This concentrate is reground in a vertical type mill to liberate gangue minerals before being upgraded

twice in column flotation cells with the concentrates being reground before being floated. The tailings

from the final column is recycled to the preceding column and the tailings from the first column is

treated in the scavenger cleaner circuit. The concentrate from this is then pumped to the concentrate

holding tank. The flash flotation tailings is treated in mechanical rougher flotation cells. The

concentrate from this is recycled back to the flash cell. The rougher tailings is treated in a scavenger

flotation circuit comprising of mechanical flotation cells. The initial concentrate is reground and

pumped to the mechanical scavenger flotation cleaner cells. The second concentrate is recycled back

to the feed to the scavenger cells. The scavenger tailings is pumped to a holding tank. The

scavenger cleaner concentrate is cleaned a further two times in flotation columns. The concentrates

are reground before being refloated. The tailings are recycled to the preceding flotation cell. The final

scavenger concentrate is pumped to the concentrate holding tank.

A new flash flotation cell is proposed to recover coarse graphite flakes as soon as possible. Since the

existing Sala/Denver rougher/scavenger and scavenger cleaner flotation cells are in very poor

condition, GBM is proposing that they are replaced by new mechanical and column cells. New cells


will last longer than reconditioned ones, have better control, and produce better flotation

characteristics, resulting in improved recovery and grade. For the PEA, GBM is using new cells.

Three flotation columns were historically used for cleaning the fines, of which two are suitable for

reuse with some minor modifications. It is there proposed that two columns existing are reused, and

two new columns are installed as the cleaner cells. The wash and air require improving. Therefore

existing and columns are used in the PEA.

It is proposed that the proper air supply is made available by installing a new blower for the

mechanical cells and new compressor for the columns. This is incorporated in the PEA.

The original reagent handling system will be replaced with a more accurate dosing system using

peristaltic pumps. New pumps are included in the PEA.

17.2.5 REGRIND MILLS

The concentrates require regrinding before cleaning to separate the gangue minerals from the

graphite. This will be done by using vertical type mills. The pebble and Svedala agitated mills are old

and require reconditioning. GBM is proposing using 2 refurbished and 5 new SMD mills. The new

units will be marginally easier to operate and maintain, however both can use ceramic grinding media

and will produce a more accurate product than historical steel balls.

The regrind mills operate best at about 50 % solids however the test work showed that the feed

density in the regrind/flotation circuits has to be very low. It is not possible to increase the feed density

hence the regrind will be operated at less than optimal conditions.

17.2.6 DEWATERING

The concentrate from the cleaner flotation is transferred to an agitated holding tank. From here the

concentrate slurry is pumped into a horizontal pressure filter on a batch basis where it will be

dewatered. The concentrate cake will be dropped onto a conveyer and transferred to the existing

thermal dryer for removal of the remaining moisture. The dry product is then transferred to a storage

silo. The filtrate from the filter will be pumped a small holding tank and from there transferred to the

process water tank.

The existing plant used a vacuum drum filter to dewater the concentrate before the drying. This was

apparently not efficient and could only reduce the moisture to 28 %. Recent test work showed that a

pressure filter can handle unthickened feed and filter it to approximately 22 % moisture. The

installation of a pressure filter should enable the dryer to operate at the expanded duty.

The existing dryer was specified to treat 2.0 t/h with a feed of 28 % moisture, with a feed moisture of

22 % the throughput can be increased to approximately 2.8 t/h.


A feed tank, new horizontal pressure filter, and associate equipment is included in the PEA. The

existing dryer and associated equipment will be used.

17.2.7 DRY SCREENING AND BAGGING

The dry graphite concentrate is transferred from the storage silo to a splitter box. This box splits the

concentrate between three vibrating triple deck screens. These screens size the concentrate into four

products; minus 80 µm, 80 µm to 160 µm, 160 µm to 250 µm and greater than 250 µm. The plus

250 µm fraction is conveyed to a bagging plant where the premium jumbo concentrate is bagged in

one tonne bulk bags for dispatch. The minus 250 µm fractions are transferred to separate storage

silos before being conveyed to a bagging plant where the concentrates are bagged in one tonne bulk

bags for dispatch.

The existing dry screens have three decks although they only produced three sizes as the top deck

was used to relieve the second deck. It is planned to produce a +250 µm product by replacing the top

deck with the correct screen aperture to screen out the +250 µm fraction.

Two existing dry screens will be used and a third will be new. The transfer of the +250 µm fraction will

be via new tubular drag conveyor from the screens to a new dedicated bagging plant. The rest of the

equipment is existing refurbished equipment. Therefore the PEA includes a new screen, conveyors

for the +250 µm fraction and bagging plant. The PEA provides for the replacement of the external

bucket elevators, which transfer the graphite fractions to the silos, and renovation of the silos.

17.2.8 TAILINGS

The final tailings from the flotation circuit is pumped to the tailings management facility (TMF).

The existing tailings tank with the existing two stage pumping system, with new two stage stand-by

pumps will be used. The first pump is fixed speed and the second variable speed.

17.2.9 WATER

The process water is pumped through the plant from the water tank using the existing ringmain.

Tailings decant water is pumped back from the TMF to the new acid rock drainage (ARD) treatment

plant. Acidic water will be contacted with lime, and neutralised in an ARD agitated tank. The

neutralised water is pumped to a clarifier where the precipitate is settled out using flocculant. A

portion of the clarifier underflow is recycled to the ARD tank to improve the formation of precipitates.

The remaining underflow is pumped to the tailings sump. The clarifier overflow is pumped to the

process water tank for recycling in the plant.

Filtrate from the concentrate filter will be pumped to the clarifier.


There are certain operations which require clean water, for example, gland service, and the existing

plant does not have a facility for this, so GBM proposes installing a small fresh water tank and running

the gland seal pump off it. A borehole pump supplies the fresh water.

The existing water circuits for process water and hosing will be used. A new fresh water tank and

piping will be installed and the existing gland seal pumps will be used. The ARD tank, pumps, and

clarifier will be new. The lime addition equipment will use existing equipment.

17.2.10 SAMPLING AND ANALYSIS

Automatic samples are taken from the cyclone overflow, scavenger and scavenger cleaner tailings,

rougher cleaner and scavenger cleaner concentrates.

Manual samples are taken from the fresh feed to the rod mill, various concentrate and tailings flows in

the flotation circuits and from the bagged product.

An on-site laboratory is used to assay the mine and process samples for sulphur and carbon.

Moisture and size analyses are also done in this laboratory.

17.2.11 ANCILLARY FACILITIES

The air for the mechanical flotation cells is provided by a new dedicated blower. Air for the flotation

columns is provided by a new dedicated air compressor. Instrument and general plant air is supplied

by the existing air compressor.

New and existing air equipment is located in the decommissioned milling room.

Flotation reagents are dosed using new peristaltic pumps to the flotation cells and columns from the

storage containers. Reagents will be purchased and delivered in ready to dose form and will be

pumped directly from their delivery vessel.

17.3 IMPROVEMENTS TO THE PLANT

17.3.1 CONDITION OF THE EXISTING EQUIPMENT

The existing plant is in poor condition, and was operated in 1996 using mostly used equipment. The

plant was subsequently modified to increase capacity and improve the product quality and recovery.

From GBM's site visit in 2012 it is assumed that the plant was not properly shut down in 2001, and

has not been kept under a proper care and maintenance programme.

Some of the equipment has been removed, for example, the drum filter, and some has been used for

retreating old concentrate, for example, the dryer and bagging plant. Pumps and piping have been

relocated and used for other purposes within the plant.


GBM is proposing to renovate the existing equipment which is proposed to be used in the new

process, that is, rod mill, two regrind mills, flotation columns, dryer, screens, and bagging plant. The

remaining the equipment will be new, for example, cyclone, flotation cells, three regrind mills,

dewatering filter, air compressors, pumps, and acid neutralising circuit.

An assessment of the work required to refurbish the required existing equipment, and the costs

included in the capital estimate are representative of the best estimate for these repairs.

17.4 BASIS FOR THE ESTIMATE OF GRAPHITE PRODUCTION

17.4.1 CRITERIA

The following criteria have been used to size the equipment.

Table 17-1: Crushing Criteria

Item Description Unit Phase 1

Operating schedule

Days per year d 365

Hours per day h 8

Utilisation % 75

Operating h/a h 2190

Annual tonnage crushed kt 155 000

Primary crushing rate – design t/h 70

Moisture content % 5.0

Density kg/m³ 2720

Bulk density t/m3 1.6

Crusher feed size - nominal mm 600

Crusher product d80 – nominal mm 6.35

Table 17-2: Fine Ore Bin Parameters

Item Description Unit

Bin type - Silo

Live capacity (mill feed) h 6

Live capacity t 100

Number of draw points - 2

Feeder type - Vibrating

Quantity installed - 2

Quantity operating – normal - 1



Design tonnage (each) - maximum t/h 36

Table 17-3: Primary Milling Criteria


Operating schedule

Days per annum d/a 365

Hours per day h/d 24

Utilisation % 85.0

Operating h/a h/a 7 446

Fresh mill feed t/h 20.8

Mill feed moisture content % 5.0

Mill type Rod

Mill dimensions

Diameter m 2.7

Length m 3.6

Speed

Normal % of critical 67

Power draw at normal conditions kW 250

Rod load

Normal % of mill volume 35

Maximum % of mill volume 50

Make-up rod diameter mm 75

Rod material - Steel

Liner material - Rubber

Mill discharge density % solids 80

Trommel undersize launder transport density % solids 80

Mill cyclone

Cyclone underflow % solids 77

Cyclone overflow % solids 33

Circulating load ratio % 250

Overflow size p80 µm 275

Table 17-4: Flash Flotation Criteria


Feed grade

Carbon % 10.3

Type of flotation cell - Column



Quantity of flotation cells - 1

Residence time min 3.1

Volume per cell m3 1

Air flow per cell Nm3/h 56.5

Table 17-5: Rougher Flotation Criteria


Type of flotation cell - Mechanical





Table 17-6: Scavenger 1 Flotation Criteria







Table 17-7: Scavenger 2 Flotation Criteria







Table 17-8: Scavenger Cleaner 1 Flotation Criteria





















Volume per cell m3 3.1


Table 17-11: Rougher Cleaner 1 Flotation Criteria







Table 17-12: Rougher Cleaner 2 Flotation Criteria








Table 17-13: Regrind Mill Criteria (all SMD’s)

Feed from Flash Flotation

Concentrate

Rougher Cleaner 1

Concentrate

Scavenger 1 Concentrate

Scavenger Cleaner 1

Concentrate

Scavenger Cleaner 2

Concentrate

Solids t/h 1.5 1.2 1.3 2.1 1.8

Feed % solids 26 5.8 2.6 6.5 2.3

Feed F80 µm 317 263 302 218 217

Prod P80 µm 282 226 271 202 197

Wi Overall 42.4 29.2 45.1 37.8 42.9

Power kW 2.4 1.2 2.4 4.8 4.8

Table 17-14: Concentrate Dewatering Criteria


Feed solids t/h 2.2

Holding tank capacity h 6

Feed Solids moisture % 89

Cake Solids moisture % 20

Solids flux t/(m² h) 0.2

Table 17-15: Bagging Plant

Item Description Unit Premium

Concentrate Coarse

Concentrate Medium

Concentrate Fine

concentrate

Operating time h/d 24 24 24 24

Graphite production t/d 9.9 11.5 15.2 17.2

Bulk density kg/m³ 550 550 550 420

Bag capacity t 1 1 1 1

Average bagging rates bags/d 10 12 15 17

bags/h 0.4 0.5 0.6 0.7

Table 17-16: Flotation Reagent Requirements

Unit (mill feed) MIBC Diesel Sodium Silicate

Flash

Rougher 1

Rougher 2

Scavenger 1

Rougher Cleaner 1

Rougher Cleaner 2

Scavenger cleaner 1

g/t

g/t

g/t

g/t

g/t

g/t

g/t

40

-

-

20

5

2

10

10

-

-

5

1

-

1

5

-

-

15

25

20

25


Unit (mill feed) MIBC Diesel Sodium Silicate

Scavenger cleaner 2

Scavenger cleaner 3

g/t

g/t

10

10

1

1

20

15

17.4.2 THROUGHPUT

The initial throughout is 100 000 t/a based on the existing Permit. It is expected that the Permit can be

extended to allow a greater throughput after approximately one year. Therefore the nominal

throughput of the plant is 155 000 t/a. This is based on the maximum that can be milled in the

existing rod mill and an estimated running time efficiency of 85 %. This efficiency is relatively low but

takes into account the small amount of buffering between circuits and the proposed operating

procedure where the operators take responsibility of maintenance during afternoon and night shifts.

The crushed throughput, 70 t/h, is based on the smallest crushing plant capable of crushing to

6.35 mm.

The mill throughput of 20.8 t/h is based on the maximum tonnage that can be milled to 80 % passing

270 µm in the existing rod mill.

The flotation section is based on the 20.8 t/h

The proposed flotation flowsheet and equipment sizes are based on the test work carried out by

Aminpro in 2013. Aminpro modelled the results of the tests to produce a flowsheet with mass balance

and flotation equipment sizes. The predicted graphite concentrate grade and recovery are based on

the locked cycle flotation test.

17.4.3 POTENTIAL FOR INCREASING THROUGHPUT

The throughput of the plant has been limited by the capacity of the existing rod mill for primary milling,

however, there is the option to use one of the existing mills which was used for regrinding for

additional primary grinding. This mill could be converted to a ball mill. The milling circuit would then

be open circuit rod milling feeding the cyclone feed sump. The cyclone would be in closed circuit with

the ball mill. This would result in a significant increase in the throughput of the milling circuit as well

as improving the milling efficiency as the ratio of reduction in the rod mill would be reduced and the

sizing by the cyclone would be done on a coarser product and hence reduce overgrinding.

The throughput of the flotation circuit could be increased by adding additional mechanical flotation

cells and increasing volume of the cleaner cells.

The new filter is capable of handling an increased throughput, however the dryer is limited to

approximately 2.8 t/h. The throughput of the dryer could be increased only if the moisture content of


the feed was reduced which test work indicates is not possible by filtering or using a centrifuge.

Therefore an additional dryer would be required.

The maximum capacity of the screens is not known, however, a fourth screen could easily be

installed.

The maximum capacity of the bagging plant is not known, however, it should be possible to improve

the throughout by rationalising the conveying system or installing an additional bagging system.

GBM has estimated the running time be a conservative 85 %. There is potential to improve on this by

improving maintenance and replacing older equipment with more reliable new units.

17.4.4 GRAPHITE RECOVERY AND GRADE

The graphite recovery is based on the results of the test work by Aminpro and its flotation model.

Table 17-17: Screen Analysis of Concentrates of the Locked Cycle Test

# µm Rougher cleaner

Rougher scavenger concentrate

Combined

% Ret % C % Ret % C % Ret % C

60 +250 13.9 95 22.0 95 18.4 95

80 +180 -250 18.8 97 23.4 92 21.4 94

140 +100 -180 26.5 94 29.6 91 28.3 92

-140 -100 40.8 89 25.0 87 31.9 88

100.0 92 100.0 91

Wt distribution 5.29 6.86 12.15

By comparing the above metallurgical results to a selection of production reports from historical

operations, further validation can be made of the achievable recoveries and grades of the various

products.


Table 17-18: Summary of Available Production Reports Year 2000

Month 2000

Head grade % of total graphite production Grade % C

% C +160 µm +80 µm

-80 um +160 µm +80 µm

-80 um -160 um -160 um

January 11.1 26.2 34.9 38.9 91.7 92 90.1

February 9.9 23.8 35.3 41.3 92.1 92.7 91.2

March 8.9 24.1 35.0 40.8 91.6 91.8 89.1

April 6.5 22.6 34.4 42.9 92.0 92.3 88.3

August 12.1 32.2 31.6 36.2 93.9 93.3 88.8

September 12.1 36.9 30.7 32.4 94.1 93.9 91.3

October 9.4 34.9 31.0 34.0 94.3 94.1 89.5

November 8.9 32.2 28.6 39.2 94.3 93.8 84.8

December 12.1 41.1 29.6 29.3 93.6 92.2 84.5

The following tables describe some product specifications used by the plant. These have been

considered by GBM and Woxna in support of the pricing assumptions.

Table 17-19: Coarse Flake Production Specification

Coarse Flake Production Specification

Grade No 160/92 Coarse

Type Natural Flake

Size Specification Min 80 % >150 micron (+100 mesh)

Carbon Specification 92 % Minimum

Moisture Max. 0.5 % H2O

Bulk Density 550 kg/m³

Typical Product Analysis Carbon 94 % C

Ash 6 %

Sulphur 0.14 % S

Moisture 0.15 % H2O

Typical Ash Analysis SiO2 46.60 %

Al2O3 25.20 %

CaO 0.29 %

Fe2O3 14.10 %

K2O 6.40 %

MgO 5.20 %

MnO2 0.08 %


Coarse Flake Production Specification

Na2O 0.25 %

P2O5 0.04 %

TiO2 0.61 %

LOI 0.20 %

S 0.02 %

Table 17-20: Medium Flake Production Specification

Medium Flake Production Specification

Grade No 80-160/92 Medium

Type Natural Flake

Size Specification Max 20 % >150 micron (100 mesh)

Max 20 % < 75 micron (200 mesh)




Typical Product Analysis Carbon 93 % C

Ash 7%

Sulphur 0.23 % S

Moisture 0.15 % H2O


Al2O3 25.80 %

CaO 0.29 %

Fe2O3 13.80 %

K2O 6.50 %

MgO 5.10 %

MnO2 0.06 %

Na2O 0.31 %

P2O5 0.03 %

TiO2 0.62 %

LOI 0.20 %

S 0.02 %

Table 17-21: Fine Flake Production Specification

Fine Flake Production Specification

Grade No -80 -/85 Powder

Type Natural Flake

Size Specification Max 20 % > 75 micron (200 mesh)


Fine Flake Production Specification




Typical Product Analysis Carbon 85 to 90 % C

Ash 15%

Sulphur 1.1 % S

Moisture 0.35 % H2O


Al2O3 21.50 %

CaO 2.20 %

Fe2O3 20.70 %

K2O 4.30 %

MgO 3.50 %

MnO2 0.09 %

Na2O 0.43 %

P2O5 0.10 %

TiO2 0.93 %

LOI 0.30 %

S 0.03 %

17.5 FURTHER BENEFICIATION

As part of the product development work, the former owners of the operation invested in a study (Lu

and Forsberg, 2001) to develop methods for the production of high-purity graphite utilising the fine

flake graphite. Production tests had shown that such a plant can increase the carbon content in the

end-product to over 99.5 % as compared to current carbon contents of 95 % maximum. Based upon

further technical studies (Boliden, 2002), a high purity plant concept was consequently developed to

fulfil the following criteria for the end products:

Reach a carbon content of 98 % - 99 % after chemical purification in one stage

Reach a carbon content of >99.5 % after chemical purification in two stages

Reduce particle size to 5 µm - 10 µm using extremely fine grinding and micronisation.

A feasibility study for a leaching plant for high-purity graphite was carried out (Boliden Contech, 2001)

and an environmental permit for such a process was granted by the local county in 2005, but has

subsequently expired.


Additional metallurgical test work and marketing studies are required before applying for a new

permit.


SECTION 18 PROJECT INFRASTRUCTURE

18.1 GENERAL

The mine site has a partially depleted existing open pit, tailings management facility, waste rock dump

areas, mine site roads, clarification ponds and a processing facility. Parts of the plant have been

abandoned and parts have been mothballed. The mobile network coverage is good and internet

connection is available.

Figure 18-1: Processing Plant the Woxna Project Area

Figure 18-2: Open Pit at the Woxna Project Area


18.1.1 OFFICE AND PERSONNEL FACILITIES

There is an office that is currently in use by the mine staff of approximately 200 m2. Several semi-

portable office sections have been installed expanding the office space to approximately 300 m2. The

condition of the office is fair, however with a longer term view a new office will be required, although

no allowance has been made for this in the PEA.

Potable water is currently brought in by water tanker to a small on site storage facility. This will

continue during operations due to the low cost and adequate service.

There is a small accommodation block of approximately 120 m2 adjacent to the administration

building. It has 4 rooms and a shared kitchen facility. It is currently utilised for short stays by visitors. It

is not envisaged that rotation staff will require this accommodation for shift purposes and it will remain

in service for intermittent use only.

The milling building has a recently refurbished bath-house to support the intended crew size.

18.1.2 ROADS

The mine shares the access, of approximately 9.5 km, from national Route 301 to the site gate. The

road is of unsealed laterite construction and during the periods of precipitation/snow and freeze/thaw,

requires consistent maintenance. The mine is in a cooperative with the local community that live along

the road and contribute to its upkeep. The financial liability is apportioned to the contribution to traffic,

and a suitable allowance has been made in the operating cost estimate to cover this.

Site roads for access and maintenance experience low levels of traffic and are of basic unsealed

laterite construction. Maintenance is performed on an as needs basis using mine and plant equipment

and contributes a negligible cost to operations.

Haul roads will be constructed and maintained by the mining contractor and included in their operating

cost.

18.1.3 WAREHOUSE, STORES AND MAINTENANCE

A warehouse/stores facility of approximately 950 m2 is located adjacent to the plant. This area will be

sufficient for process plant spares and graphite product storage. This building is enclosed and

lockable, though not heated.

Due to the manning structure, most maintenance will be performed either in-situ or in a contractors’

facility in the nearby town, and therefore, a large dedicated maintenance area with cranage is not

required on site. There is some space adjacent to the gravity equipment that has been used for small

maintenance projects. This area should remain the location of these works, as well as tool storage, as

this area is part of the heated building.


The mining contractor will provide and operate their own maintenance facilities.

18.1.4 LABORATORY

A prefabricated laboratory building will be installed adjacent to the process plant building so as to be

free from contamination of dust and vibration from the plant. The lab will be prefabricated and capable

of processing approximately 10 mining samples, 16 plant samples and 50 product samples per day

for size analysis, carbon content, sulphur content and moisture content.

As part of the laboratory supply, safety equipment, training, operations manuals and procedures will

be included.

18.1.5 FUEL

Current diesel storage on site is within a containerised facility. This will be positioned adjacent to the

building within the bunded area that was used for fuel and oil storage previously.

The mining contactor will provide and operate its own diesel storage facility.

18.1.6 VEHICLES

The plant operations require only simple mobile equipment to sustain steady operations. A loader will

be leased to feed ore from the stockpile to the crushing plant at an approximate utilisation of 30 %.

During other times, it will be performing road maintenance and with the use of a jib arm, act as

cranage for the plant.

A forklift already owned by the operation will be in use in the bagging plant and stores area.

The operation also owns a small ‘JLG’ elevated work platform for maintenance around the plant.

As the plant is compact and crew numbers small, only one light vehicle is required to move about the

site for periodic visits to the tailings management facility and clarification pond.

Any further equipment required by the plant will be hired for specific tasks on an hourly basis. Access

to this equipment is good from the nearby towns.

18.1.7 FENCING AND SECURITY

Allowance has been made to install a 2 m chain wire fence around the entire site. It is estimated that

this fence will be approximately 5.5 km in length.

18.2 ELECTRICAL

The existing power supply to the project site is from the national grid. The capacity of this connection

is limited to 4 000 KVA. No allowance has been included for upgrades to the network other than


connection to the new transformer and relocation of a section of the existing network to allow for the

proposed mine pit advancement.

The power supply to the plant shall supply power to the existing process loads as well as the new

loads required as part of the plant restart project.

A summary of the process plant power demands are listed in Table 21-12. The maximum demand

includes an allowance to ensure that there is sufficient capacity to start the largest motor on site while

all other equipment is under normal operating conditions. However, sequence starting could be used

to ensure the maximum demand does not exceed the average operating load.

New electrical equipment, inclusive of plant PLC/SCADA system, transformer, mill motors and drives

and cabling is required in order to re-start plant operations.

Existing electrical equipment must be tested, refurbished if required and re-commissioned in order to

re-start plant operations.

A limited number of new control loops, as required by GBM process control philosophy shall be

implemented.

18.2.1 POWER DISTRIBUTION

There are currently two 1 000 kVA transformers installed in a Transformer / MV switchgear / LV

switchgear building located close to the plant.

The transformers of different vector groups (T1 = Dy 0; T2 = Dy 11) and can therefore not be

connected to the bus in parallel. Attempts to do this in the past have resulted in the bus feeder

breakers tripping.

The transformer rooms are undersized and have inadequate ventilation or air filtration. Part of the

transformer building is a common MV (12 kV) and LV (400 V) room.

A new MV 1 900 kVA substation to be installed and supplied in a kiosk type enclosure, including MV

switchgear with spare feeder to allow for future expansion. This would remove the need for the MV

switchgear in the existing location thus separating the LV and MV switchgear.

Power will be distributed throughout the process plant on existing and new circuits from the plant main

LV switchgear assembly, which will remain in the existing substation building. Reticulation throughout

the plant will be via radial circuits at 400 V.

There is an additional electrical room inside the plant building and opposite the MV/LV building which

houses two PFC (Power Factor Correction) panels. This is assumed to be sufficient for the new and

existing loads and no allowance has been made for additional PFC. Allowance is made for tested and

refurbishment.


18.2.2 MCCS AND MOTORS

From breakers in the existing LV distribution room a number of existing starter panels (approx. 20) are

fed throughout the plant building. These panels shall be tested and refurbished where required prior

to re-commissioning.

A new MCC room will be installed to house the motor starters for the new process loads.

Variable speed drives (VSDs) where required will be supplied as part of the MCC package.

The mill motors and drives are to be replaced. A variable speed drive has also been included for the

Rod Mill.

Basic motor tests on all other existing plant motors will be required prior to re-commissioning.

At present, the motors can only be started from the PLC/SCADA. There are no field start or field

emergency stops installed at the motors. However, every motor has a local isolator. Some local

isolators are located in positions which are not easily or safely accessible. The only way a motor can

be stopped in the field is by activating the local isolator. The motor starter circuit must be modified for

emergency stops and start-up warnings (where required). Installation of an emergency stop near each

motor and associated cabling is required as a minimum however, local control stations are allowed for

to facilitate plant operation by a small crew.

18.2.3 CABLE SIZING AND SELECTION

All new cables will be sized and selected according to Swedish or equivalent international standards.

All new power and control cables will be steel wire armour cables for mechanical protection.

18.2.4 EMERGENCY POWER

No allowance has been made for emergency power.

18.2.5 CONTROL AND INSTRUMENTATION

The plant motors are all started and stopped from the PLC/SCADA system. Currently there are no

field start or stop push-buttons in the plant. Therefore, the PLC is an important requirement in order to

operate the plant. The PLC/SCADA was apparently damaged by lightning some years ago and is not

functioning at present.

The model of Mitsubishi PLC which is currently installed in the plant is regarded as outdated ‘legacy’

equipment. It will be replaced with an up to date Mitsubishi PLC. This would have the advantage of

being able to re-use the original software (program).

The flotation columns level control loops are using outdated instrumentation (Fisher Bailey Porter).

These are to be replaced with ABB Instrumentation


The existing instruments have no tag numbers or service descriptions attached. These shall be

replaced during testing re-commissioning.

The mills, or other rotating or moving equipment, do not seem to have start-up warnings

(audio/visual). These shall be added to improve the safety of the plant.

18.2.6 VENDOR PACKAGES

The proposed crushing plant package will be vendor supplied with on-board electrical switchgear,

motor starters and controls. A spare LV feeder breaker in the LV distribution panel will be utilised to

supply this package

The pressure filter will have its own control supplied as a vendor package. This will be for the control

of all functions of the filter as well as quoted accessories and process control functions.

There is a drying plant with a Triplex dryer which is supplied by ‘Gebrüder Pfeiffer AG Kaiserslautern’

as well as a packing plant. Both plants are ‘vendor packages’. The local control panels of these

‘vendor packages’ are supplied (fed) from the LV room.

The packing plant is controlled from a local control room which contains an electrical control desk as

well as a relay logic panel and mimic. The graphite packing plant local control panel is to be upgraded

to PLC control.

18.3 TAILINGS MANAGEMENT FACILITY (TMF)

Tailings Management Facility evaluation and design has been undertaken by Tailings Consultants

Scandinavia AB (TCS) who were contracted directly by Woxna.

Woxna mine has an existing TMF, which was constructed during the previous operation of the mine

(operation shut down in 2001). The TMF consists of a tailings beach and a downstream area where

the water collects. Tailings were deposited from the south abutment of the east dam, causing the

tailings to slope towards the west dam.

The future production has been estimated to generate a total of 1.7 Mt of tailings. With an estimated

dry density of 1.5 t/m3 – based on samples from the existing tailings – a total of about 1.0 M m3 is

expected to be stored in the TMF. A conceptual tailings design has been completed to contain

1.5 M m3

Tailings will be pumped from the process plant and deposited in the TMF using spigots. Deposition

will occur from the East and West dams, as well as along the south boundary. Deposition will occur

from the east dam, forcing the water pool to gather in the west part of the tailings impoundment. A

spillway is constructed at the north-west boundary for discharge of water to the clarification pond.


Figure 18-3: Tailings and Water Management at Woxna

The current dam embankments are in poor condition, and will be restored according to Swedish

guidelines (GruvRIDAS) prior to the reopening of the mine.

18.3.1 OPTIONS

In addition to an upstream dam, which is the selected option and is described below, there are a

range of other designs that may be suitable – most notably to a downstream dam. At present, a

downstream dam this is not considered possible due to the higher cost. However, if it is possible to

use waste-rock as construction material (non acid generating) as well as material from the stripping of

the ore, this type of design would be viable and also a way of using waste-rock instead of putting it on

a waste-rock dump. There are also possibilities of using thickened tailings disposal which could mean

that more tailings can be stored in the area. However this needs to be further investigated since this

could affect the plant water balance and the discharge.


A third option is to use dry-stacked tailings in the facility to reduce costs further, as the use of dry-

stacked tailings would reduce the need for dam walls This is considered the most risky operation and

not normally suited to the local climate. The operation is however comparatively small and under

certain conditions; variations of the above options might present a possible alternative.

18.3.2 PROPOSED DAM STRUCTURE

Initially the dams will be reinforced and tailings deposited to fill up the existing free volume. The dams

will then be raised as upstream dams with a 1V:6H slope downstream, which provides long term

stability and requires less filter materials (a low hydraulic gradient minimizes the risk for internal

erosion). Horizontal filters will be placed on top of the existing dam crests and will function as drains in

the upstream construction. The dams will be raised at regular intervals during the life of the mine to

accommodate the deposited tailings. The highest crest level of the dams will be +283.5 m.

Figure 18-4 shows the conceptual design of the upstream dams. The freeboard to the tailings is 0.5

m, and a beach of at least 50 m is recommended.

Figure 18-4: Upstream Tailings Dam

18.3.3 SPILLWAY

The choice of spillway for emergency discharge will require more detailed studies. Three options are

available:

a) spillway in the northern part of the TMF, with a channel leading the water around the higher

ground north of the western dam. This design will not be suitable in an early stage but may be

a good option after raising the dams. (In the initial phase this solution will require a very deep

channel)

b) Sloping the TMF to accumulate the free water in the area of the northern part of the western

dam, where a spillway may be installed. To allow for deposition of tailings close to the area of


free water a diversion berm is constructed from the western dam eastwards into the

impoundment. This design may be suitable at all stages of the TMF. Regarding costs, this

option will require a diversion berm and an impervious dam in the northern part of the TMF.

These will both become more and more costly for each raise of the TMF.

c) A central intake tower with a culvert leading through the dam wall. This may be suitable at all

stages of the TMF. The required discharge capacity is fairly low since the TMF has a rather

small area. The use of a culvert through the dam wall is not a common solution in Sweden,

and regulatory authorities may object.

Whist the final design has not been decided, an allowance in the cost estimate has been made based

on what would be considered the median cost of the range of solutions.

18.3.4 PHASED TMF EXPANSION

Due to the high capital cost, the expansion of the TMF will be done progressively. The design concept

and costs included in the LOM CAPEX are representative of this approach.

18.4 CLARIFICATION POND (LAKE UXATJÄRN)

The use of the small Lake Uxatjärn, located close to the TMF, as a clarification pond for the mine will

continue.

There is a small dam wall for the lake and it is in need of some remediation work as well as expansion

to provide greater operational flexibility, though it is acceptable in the short term to continue its use in

its current state.

A works program is proposed to coincide with Stage 2 of the TMF construction. The dam

embankment will be restored and raised, and the spillway will be redesigned to accommodate

appropriate design floods. The dam embankment is constructed similar to the dams of the TMF, i.e.

with a 1V:2H upstream slope and a 1V:3H downstream slope. The crest of the dam embankment will

be +236 m. The final proposed configuration is shown in Figure 18-5.


Figure 18-5: Clarification Pond Expansion

Water from the clarification pond will be recycled into the plant. If excess water is required to be

discharged, it will undergo conditioning (Figure 18-6) before being released to the river Älman.

Figure 18-6: Water Conditioning Plant


18.5 EMERGENCY PREPAREDNESS

According to Swedish legislation an Emergency Preparedness Plan (EPP) must be filed before start-

up, and dams need to be monitored. Every site is compelled to have an up-to-date Operation,

Maintenance and Surveillance (OMS) manual, which includes information about the TMF as well as

the dam safety organisation

18.6 WATER MANAGEMENT

A water balance has been completed for the site that indicates that under normal circumstances no

additional water will be required (annual net positive water balance). A small amount of raw (fresh)

water is required for the process, for example, gland service, and as such, the cost of installing

equipment to replenish water from the river Älman is included. Water will be pumped from the river

Älman or from one of two boreholes, approximately 1 000 m via Lake Uxatjärn at a rate of 45 m3/h. It

is proposed that the pipe is laid on the ground adjacent to an access road and covered with a thick

layer of woodchips to protect from freezing. This option attracts a much lower capital cost to buried

pipework via the most direct route (550 m) and is in use in other parts of the site with success over

the winter months.

Woxna is already permitted to draw 45 m3/h from the downstream river Älman. No hydrogeology has

been completed to support that option and more boreholes may be required. Study work to verify this

should be conducted in the next stage of project development.


SECTION 19 MARKET STUDIES AND CONTRACTS

19.1 INTRODUCTION

The purpose of this section is to outline current trends and potential within the international natural

graphite market. Particular attention is paid to the supply and demand trends pertaining to flake

graphite of varying size and purity within Europe which is expected to be the principal market for

Woxna graphite. The market outline provided here forms the supporting basis for defining the graphite

prices that are used in a financial evaluation of the Woxna project. Graphite prices in Europe are

usually negotiated between consumers and suppliers on a spot basis reflecting the supply and

demand situation for the particular graphite purity and flake size at the time, however, longer term

contracts, up to a year, can be agreed.

A literature review was completed of the research report Natural & Synthetic Graphite: Global Industry

Markets and Outlook, by Roskill Information Services Ltd (2) to aid in the preparation of this section.

GBM also relied heavily on members of the Flinders management team who have been monitoring

the graphite market and meeting with potential off takers for a number of years.

19.2 MARKET DESCRIPTION

Graphite can be produced from either the natural mineral or through a synthetic process. The global

market for synthetic and natural graphite was estimated to consume 2.4 M tonnes in 2011, worth

USD 9 billion.

Breaking this down, synthetic graphite electrodes is the largest segment by value at USD 4 billion

followed by synthetic graphite blocks and other synthetic products. Natural and synthetic powder,

including flake, are estimated to be worth USD 1.1 billion annually.

Since 2010 the graphite market has attracted increased interest due to rising prices stemming from

growing graphite demand and potential threats to graphite supply. Prior to this the market was

relatively static, as Chinese domination of graphite production allowed China to dictate prices.

Demand has grown steadily in traditional markets, such as refractories, lubricants, friction products

and even pencils while strong growth has occurred in new markets, such as those for expandable

graphite products and lithium-ion battery anodes. Until very recently there has been a lack of

investment in graphite mining ventures outside of China.

Natural (flake, amorphous and lump types) and synthetic graphite can and are used in many of the

same applications. The choice of material is largely determined by their individual properties and/or

their price. Some forms of natural graphite however, particularly flake, have superior performance in

some uses such as refractories and there is little substitution. New technologies, applications and


speciality products are increasing demand for higher specification graphite products. Over the past

few years demand has tended to migrate from synthetic and amorphous graphite to natural flake

graphite due to improved natural graphite products as well as competitive prices.

19.3 NATURAL GRAPHITE DEMAND

Annual demand for natural graphite is estimated at 930 000 t with flake graphite demand (540 000 t)

exceeding amorphous. Refractories, used to line furnaces, are the largest segment for natural

graphite comprising more than 50 % of demand or 485 000 t. Foundries, lubricants, batteries and

friction products then follow in terms of market volume.

Asia has been the largest and fastest growing consumer of natural flake graphite for the last decade,

due in a large part to the rapid economic development of China. Asia accounts for 64 % of global

flake demand. Europe at 85 000 t/a is the second largest market (16 %) globally.

Roskill forecasts that demand for total natural graphite will grow at an average of 3.3 % per annum to

reach 1.1 Mt by 2016 while demand for flake graphite is forecast to grow at an average of 4.3 % per

annum to 2016.

The refractory market is forecast to grow at 3 % per annum driven by increased steel demand in

China. Natural and synthetic flake graphite are also used for anodes in the rapidly growing market for

lithium-ion batteries. This segment is projected to grow at 10 % to 12 % to 2016.

19.4 NATURAL GRAPHITE SUPPLY

China currently dominates world production of natural graphite, accounting for 70 % of supply in 2011.

However, this was as high as 80 % three to four years ago. This reduction is due to the closure of

Chinese amorphous graphite mines in Hunan. China produces both flake and amorphous graphite but

is more abundant in amorphous graphite deposits.

A change in export policy by China in 2011 led to the replacement of export rebates on graphite with

the imposition of taxes and export permits temporarily slowing down in graphite exports and resulting

in a spike in graphite prices.

Flake graphite is also produced in Brazil, India and Canada while Mexico produces amorphous

graphite. Apart from one medium sized graphite mine in Norway, Europe is heavily reliant on imports

to meet its graphite needs.

In the short term Roskill predicts that natural graphite supply is likely to increase as new projects

come on-stream and Chinese amorphous production recovers. It is likely during this time that supply

for flake graphite will meet any growth in demand.


Roskill forecasts that supply will rise by 3 % per annum to 2016, resulting in a supply of natural

graphite of 1.1 Mt. Many existing mines are being exhausted, or costs are rising as they get deeper,

but higher graphite prices since 2011 has encouraged investment in new graphite projects, of which

there were over 70 by mid-2012. Although most of these new ventures are in the early exploration

stage, it is expected that new mines will be responsible for at least 200 000 t of the increased flake

capacity supply by 2016.

In the long term flake graphite demand could potentially exceed supply if production of lithium-ion

battery anodes ramps up to meet an acceleration in sales of hybrid and electric vehicles and because

of the long time it takes to develop new graphite mines. Demand will likely shift towards larger flake,

higher grade graphite, for use in expandable graphite products.

Historically amorphous and flake graphite have accounted for 60 % and 40 % of natural graphite

production respectively. This ratio has altered to around 55 % to 60 % flake graphite and 40 % to

45 % amorphous graphite since 2011, largely due to the closure of Chinese amorphous graphite

mines.

19.5 NATURAL GRAPHITE MARKET PRICE TRENDS

Historically graphite prices were stable, maintained by abundant supply from China and typically at

levels too low to attract new investment in mines outside China. In recent years, however, China’s

prices have risen due to increasing operating costs, governmental policy that increase export taxes

and tariffs, higher domestic demand and diversification of Chinese producers into higher grade

materials.

Graphite prices have fluctuated since 2010, initially with a sharp rise as the market reacted to

increasing global consumption, closure of amorphous mines in China and imposition of Chinese

government taxes. In 2012 prices began to retreat due to buyer resistance to the rapid rise in graphite

prices as well as recession in Europe and the USA and lower than expected growth in China.

Figure 19-1 illustrates the average price of natural graphite by type from 2003 to 2012.

It is expected that prices will bottom out in 2013 and then start to recover in line with economic

activity. Rising costs of production, such as those for labour, meeting environmental regulations and

mining overheads, in conjunction with more accessible reserves being exhausted, are contributing to

pushing prices upwards.

Natural flake graphites command a price premium over amorphous graphites, with price being directly

related to flake size and carbon content. Figure 19-2 illustrates the average market price of medium

size flake graphite by carbon content from 2003 to 2012. Roskill expects that flake graphite prices will

remain higher than those of amorphous graphite. As more new flake graphite mining projects come

on-stream it is expected that the upward pressure on graphite prices will relax somewhat.


Figure 19-1: European Average Price of Natural Graphite by Type from 2003 to 2012 (USD/t) - (2)

Figure 19-2: European Average Price of Medium Size Flake Graphite by Grade from 2003 to 2012 (USD/t) - (2)


19.6 PRICING BASIS FOR THE WOXNA GRAPHITE RESTART PROJECT

The pricing for the project is based on current and forecast prices for relevant grades as compiled by

industry intelligence analyst Industrial Minerals (IM). Graphite prices are not determined on a

publically traded market but instead determined privately between buyers and sellers. IM tracks

industrial minerals prices (including graphite) through regular contact with producers, traders and

purchasers and publishes graphite prices that are generally recognised by the graphite industry to be

the best reflection of current market conditions.

It has been common practice to use a 24 month trailing average graphite price in graphite PEAs. The

Woxna PEA utilized a 12 month trailing average graphite price so as to exclude the peak price period

occurring in 2011/12.

19.7 PRODUCTION OF SALEABLE GRAPHITE

Applying Woxna’s planned product distribution to IM’s 12 month trailing average graphite prices

produces an average selling price of 1 199 USD/t for the Woxna PEA. This selling price is considered

to be conservative when compared to the 24 month trailing average graphite price of 1 548 USD/t or

prices used in other graphite project PEAs.

Table 19-1: Average Sale Price Estimation

Size (µm) Purity Proportion of Production

1 Year Trailing Average (USD/t)

Annual Quantity (t)

+ 250 95 % 18 % 1 824 2 990

+ 180, - 250 94 % 22 % 1 526 3 650

+ 100, - 180 92 % 28 % 1 009 4 650

- 100 88 % 32 % 787 5 310

Average / Total 1 199 16 600

The production reports of a comparatively small volume of product from historical operations validate

the metallurgical, recovery and saleability findings.

The demand for graphite has weakened during 2013 as a result of reduced demand and slower

economic growth in China, but many analysts believe that the prices have bottomed out and that the

market will see a positive trend from now and throughout 2013/14. Demand for flake graphite

continues to grow led by refractories. Higher growth levels are expected in the battery sector. There

will be a continued shift away from amorphous graphite as emerging applications typically require

large flake and/or high-purity grades. If lower prices continue, companies developing new graphite


projects will face competition for a shrinking pool of investment opportunities. This could be an

advantage for advanced projects as Woxna Graphite, which company will be able to produce and sell

graphite in the middle of 2014.

China will continue to lead the way for international trade and pricing in graphite, as it has done for

many years. In the long term, Chinese export prices are expected to rise because of the increasing

cost of domestic production (as labour, environmental and overhead costs rise) and because of

increasing Chinese control through consolidation.

Meanwhile, the quantity of graphite available for export will decrease as China ramps-up production

and export of value added products. These factors will gradually push consumers in the rest of the

world to look for alternative sources of raw material elsewhere, which is an advantage for Woxna

graphite.


SECTION 20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL

OR COMMUNITY IMPACT

20.1 ENVIRONMENTAL PERMITTING REQUIREMENTS AND STATUS

The Swedish Environmental Code is a combined code which contains general requirements for the

environment, land and water use, environmental impact assessments and the code also enables

European Directives e.g. the water frame work directive to be incorporated into Swedish law. The

code also regulates areas of national interest including Natura 2000 areas, reindeer herding and

mineral deposits. The ESIA chapter in the code summarizes the legal requirements for an EISA. More

information can be found in sub-decrees from different authorities. The Swedish legislation requires

the applicant to undertake an environmental (and social) assessment (in Swedish “MKB”) at two

different stages during the development of a mining project.

The first MKB is produced when applying for an exploitation concession in accordance with the

Minerals Act (SFS 1194:45) from the Mining Inspectorate of Sweden (“Bergsstaten”). Although the

exploitation concession follows the Minerals Act, a MKB must be performed according to the

requirements of the Swedish Environmental Code (1998:808). The produced MKB, however, is to

some extent simplified, e.g. no alternatives are needed to be presented or explored. The emphasis of

the MKB is only to demonstrate that there are no obvious conflicts and that mining operation is

possible with a minimum of impact on the surrounding land uses. The assessment is based on early

project design information (“PEA” information). When granted, the company gets the sole right to the

deposit for 25 years which at the end of the period may be extended further. Stakeholders also have

the right to appeal to a higher court. Stakeholder consultation is in general not required, but is

recommended, and the County Administrative Board as well as the local municipalities (local

environmental authorities) will provide comment on the MKB. Approval of the exploitation concession

is required prior to submitting an application for an environmental permit to mine (“extraction permit”)

to the Land and Environmental Court (In Swedish “Mark och miljödomstol”).

The environmental permit application, under the terms of the Environment Code (SFS 1998:808), also

requires an MKB, but an extended one, which is based on a more defined (definite) project

description (“PFS” or “FS” level) than the MKB prepared for the exploitation concession application.

Several alternative ways of mining, TMF designs and locations etc. must also be evaluated and

described as well as transportation and processing options, noise and vibration etc. The

environmental permit application must also be followed by a technical description, which describes

the future operations in detail. Again, the application is evaluated by the Swedish Land and

Environmental Court with input from various regulatory authorities, the County Administrative Board,

Municipalities, Swedish EPA etc. Stakeholder consultation is required during the application process.


Once granted, mining operations can be started. Stakeholders also have the right to appeal the permit

to a higher court.

Flinders Resources Ltd already has existing and valid exploitation and environmental permits for the

mine and the central deposit since the mine has been in operation. According to the Swedish

legislation the mine is still in operation, although the company has not actively been mining since

2001. All the documentation (Environmental Impact Assessment (“EIA”) etc.) is in Swedish and mostly

originates from the permitting process in 1992 (Kringelgruvan).

The company also had valid and existing exploitation permits for the surrounding deposits (Mattsmyra

and Gropabo). Since those deposits have not been mined, the environmental permits expired in April

2012 (the permits had a life-time of 7 years before mining had to commence).

All existing mining leases and extraction permits are presented in Table 20-1.

Table 20-1: Woxna Graphite Project Tenure

Property

Type of Area Extraction permit Valid Until

Conditions Mineral Tenure (Ha)

(Environmental permit)

(date)

Kringelgruvan Mining Licence

30.76 Permits received and still valid 1992-09-17

and 1992-10-27 31/12/2016

Environmental permit in hand for 100 kt annual production.

nr 1 with exploration

licence 4, 5, 6 & 7

(Exploitation concession)

Mining has depleted 300kt from historical resource.

Gropabo Mining Licence 18.2


21/02/2025 ***


Expired 2012-04-18

Mattsmyra Mining Licence 72.97


21/02/2025 ***


Expired 2012-04-18

Månsberg Mining Licence

24.77 No application filed 27/12/2024 (Exploitation concession)

* Is automatically extended 10 years when a mine is in operation

** The central deposit incl. mine, mill, tailings impoundment, clarification pond and the rest of the facilities.

*** Surrounding deposit (no mine).

20.2 FURTHER PERMITS

Changes to the stipulations in the present permit (from 1992) as needed for an expansion, requires

the authorities to issue a new/updated permit. Under the scenario presented in this report, in order to


process more than 100 000 t/a feed in any given calendar year (i.e. to reach the nameplate tonnage

of 155 000 t/a), Woxna will need to apply for a new permit. This permit will include consideration for

the whole operation including mining, processing, power, water management, tailings and waste

management.

The environmental permit application, under the terms of the Environment Code (SFS 1998:808),

requires an EIA and a technical description which describes the future operation in detail. The

application is evaluated by the Swedish Land and Environmental Court with input from various

regulatory authorities most importantly the County Administrative Board. Stakeholder consultation is

required during the application process and the company have already had meetings with the local

community (residents), the municipality as well as the County Administrative Board. However the

permitting process has not formally started.

In order to prepare the EIA, baseline data is needed as well as technical data. The company is in a

good position with water baseline data from 1997 to the present day. During 2012 the company also

performed both a nature inventory report as well as a study of the recipient (fish, fauna, sediments).

Since the mine is an existing one it is not expected that the company needs to do more studies than

noise and vibration in order to comply with what the authorities expect, and it should not take more

than 12 months to 18 months before a new permit will be granted. It is also expected that the

permitting process should be straight-forward. There are no obvious conflicts in the area, since the

activity at the mine mean it is technically already in operation. If the permit application is filed

early 2014, a new permit should be in effect by late 2015.

Owing to the construction and ramp-up period, it is not expected that cumulative plant throughput

exceeding permitted tonnages will be achieved until late in the 2015 calendar year. If circumstances

lead to delayed permit approval, short term permits can be granted to cover the balance of an

operating year, thus delaying the requirement for a permit until mid to late 2016.


SECTION 21 CAPITAL AND OPERATING COSTS

This section serves to provide information pertaining to the preparation of the capital cost estimate

(CAPEX) and operating cost estimate (OPEX). This includes the assumptions, exclusions, methods

and data sources for the various aspects of the cost engineering exercise.

21.1 BASIS OF ESTIMATE

The CAPEX and OPEX have been estimated based on the following project specifications,

determined through the ore resource study and testwork.

Table 21-1: Operating Inputs

Description Units Phase 1

Crushing Plant Design Dry Tonnage t/h 70

Process Plant Design Dry Tonnage t/h 20.8

Crushing Plant Availability Max - 75 %

Process Plant Availability - 85 %

The level of engineering detail and confidence has a direct effect on the precision and accuracy of the

final estimate. To ensure the required level of confidence is achieved in the preparation of the cost

estimate, several engineering design inputs are required which form part of the overall basis. The cost

estimate for this project is based on an existing plant and infrastructure as well as preliminary

engineering on modifications and additions. The engineering documents which have been used to

develop the CAPEX and OPEX are listed in Table 21-2.

Table 21-2: Supporting Documents

Title Reference

Process Design Criteria (3)

Project Design Basis (4)

Mass and Water Balance (5)

Capital Expenditure (CAPEX) Cost Database (6)

CAPEX and OPEX for TMF (Golder/TCS) (7)

Operational Expenditure (OPEX) Cost Calculation (8)

Mechanical Equipment List (9)

Electrical Load List (10)


Title Reference

Production Schedule and Mining Costs (11)

Site layouts and GA Drawings -

21.1.1 PROJECT AREAS

The project has been divided up into project areas (detailed in Table 21-3), each with an area number

and brief description.

This structure forms the basis for the financial model, in which the capital cost is reported for each line

item.

Table 21-3: Project Area Breakdown

Area No. Area Name

000 General

100 Mine

200 Processing

300 Waste Disposal

400 Product Handling and Transport

500 Infrastructure and Utilities

21.1.2 METHODOLOGY

The cost estimate is based on the current level of engineering design and has been generated from

supporting engineering quantities and cost information. Cost information has been derived from the

following sources:

Preliminary evaluated tender for the crushing plant

Budget quotations for other plant equipment from suppliers on preliminary process data

Historical cost information sourced from in-house and commercial databases. (Infomine

(12))

Actual expenditures and budgeted figures for site refurbishment works

Data from current and historical operations

Factors for ancillary works have been applied to the Mechanical Equipment. Factors are

derived from the in-house database and from estimating publications (13).

This estimate has been prepared to an accuracy range of ± 30 %.


21.1.3 ASSUMPTIONS

The following assumptions have been made during the preparation of this estimate:

New equipment will be purchased where equipment has to be replaced or additional

equipment is required

Existing equipment will be refurbished

Equipment costs are based on GBM selected suppliers that may not necessarily be the

final equipment supplier for the project

Equipment costs are based on information and testwork available at the time of design.

All required earthworks materials such as fill, sand, gravel, crushed rock, etc. can be

sourced within 2 km of the plant site

21.1.4 CURRENCY AND EXCHANGE RATES

Capital and operating cost estimates are prepared in mixed currencies and reported in United States

dollars (USD). The exchange rates used in calculations are shown in Table 21-4.

Table 21-4: Currency Exchange Rate

Currency Code Rate

United States dollar USD 1.0000

Swedish Krona SEK 6.5401

Canadian dollar CAD 0.9696

Euro EUR 0.7569

Pound sterling GBP 0.6512

Note: exchange rates are taken from http://www.xe.com as of 15 August 2013.

21.1.5 BASE DATE AND REPORTING CURRENCY

The cost estimate has a base date of the 3rd quarter of 2013. The estimate is reported in USD.

21.1.6 CAPEX INCLUSIONS

Mechanical equipment costs.

Some commodity quantities have been derived from basic design documentation.

The balance of the piping, platework, earthworks, civil, structural, control and

instrumentation, electrical, installation and freight forwarding have been calculated as

a factor of the mechanical equipment costs. These factors have taken into account

the nature of the existing operation. Specifics are described further below


Infrastructure including roads, security and fencing.

Any upgrades and modifications required to utilities including power supply, raw

water supply, potable water, compressed air, fuel depot and sewage treatment to

support the restarted operation.

Any upgrades and additions to buildings associated with the beneficiation plant such

as the laboratory, mill office, warehouse, workshop, emergency services building,

and administration office, as required to support the restarted operation.

EPCM as cost for the support of the delivery of the plant and infrastructure as per the

battery limits. However, it is intended that this level of assistance is as low as

practicable.

21.1.7 OPEX INCLUSIONS

Provision of water, fuel and electricity.

Manpower for the operation of the mine, process plant and associated infrastructure.

General administration costs.

Provision of appropriate operating spares.

Necessary maintenance costs (equipment, vehicles, roads, pipelines, etc.).

Lease of relevant mobile equipment for the operations of the beneficiation plant,

support services and personal use.

21.1.8 EXCLUSIONS

Costs incurred to date on site associated with the Woxna Graphite Restart Project,

including consultant and testwork fees.

Provision for demolition works or removal of existing infrastructure.

Allowance for cost escalation or for currency fluctuations.

Allowance for future scope changes.

Land acquisition.

Environmental studies and permitting.

Additional consultants.

Operational insurances.

Community relations and services.

Labour stand down costs.

Demurrage costs.

Any geotechnical work or test work.

Insurance costs.

Management reserve


21.1.9 RISKS AND OPPORTUNITIES

Several risks and opportunities have been identified in the preparation of this estimate as a result of

the estimating methods:

The use of database costs may introduce errors or omissions in failing to allow for

site specific requirements.

No allowances have been made for the event of a labour stand-down or other

significant interruptions to work.

As this is a PEA level study the level of engineering is only sufficient to estimate the

costs to within ± 30 %.

21.1.10 CONTINGENCY

Contingency is an amount added to an estimate to allow for items, conditions, or events for which the

state, occurrence, or effect is uncertain and that experience shows will likely result, in aggregate, in

additional costs.

It covers costs which may result from incomplete design, unforeseen and unpredictable conditions, or

uncertainties within the defined project scope. Contingency will usually include allowances for

planning and estimating errors and omissions, minor price fluctuations other than general escalation,

design developments and changes within the scope, and variations in market and environmental

conditions. The contingency estimate for the project is broken down into the risk drivers identified as

follows:

Project definition.

Estimating methods and estimating data.

Engineering design efforts.

Supplier quotations or database costs.

Site data and test work.

Contingency is estimated using a factored method which is suitable for a Class 4 cost estimate. As a

result of the estimating techniques employed, a 20 % contingency is to be used.

21.2 CAPITAL COST

The capital cost estimate is broken down into direct, indirect and working capital costs shown in

Table 21-5. The capital cost breakdown structure is reported on a project area basis using the

structure shown in Table 21-3.

The total initial capital investment for the start-up is USD 16.64 M, which includes contingency and the

cost of the first tailings expansion.


Table 21-5: Initial Capital Cost (M USD)

Cost Centre Total 000

General 100

Mining 200

Process 300

Waste 400

Product 500

Infrast

Total Capital Investment 16.72 5.69 0.15 6.80 3.37 - 0.70

Fixed Capital 14.33 3.75 0.15 6.36 3.37 - 0.70

100 - Direct 10.28 0.31 - 6.36 2.91 - 0.70

101 - Earthwork & Civils 3.07 0.21 - 0.22 2.64 - -

102 - Structural 0.09 - - 0.09 - - -

103 - Buildings 0.16 - - - - - 0.16

104 - Mechanical Refurb 0.46 - - 0.46 - - 0.01

105 - Mechanical New 4.40 - - 3.83 0.09 - 0.48

106 - Mobile Equipment - - - - - - -

107 - Electrical 1.11 - - 1.11 - - -

108 – Control & Instrumentation 0.20 - - 0.20 - - -

109 - Piping 0.28 - - 0.04 0.18 - 0.05

110 - Platework 0.07 - - 0.07 - - -

111 - Mechanical Installation 0.35 - - 0.35 - - -

112 - Freight Forwarding 0.10 0.10 - - - - -

200 - Indirect 4.05 3.44 0.15 - 0.46 - -

201 - EPCM 1.06 1.06 - - - - -

202 - Owner's Costs 0.20 0.05 0.15 - - - -

203 - Consultants 1.34 0.88 - - 0.46 - -

204 - Field Indirect 0.07 0.07 - - - - -

205 - Insurance - - - - - - -

206 - Contingency 1.38 1.38 - - - - -

Working Capital 2.38 1.94 - 0.44 - - -

301 - Consumables 0.04 - - 0.04 - - -

302 - Initial & Commis Spares 0.40 - - 0.40 - - -

21.2.1 DIRECT COST DEVELOPMENT

The direct costs for the project include:

All labour required for Project construction and management activities.

All material and equipment required for construction and refurbishment

Mechanical, electrical, control, instrumentation, civil works, earthworks and piping

installation services.

Transport and freighting services.


21.2.1.1 STRUCTURAL

An allowance has been made for a nominal tonnage of equipment based on the general arrangement

drawings provided. Costs for fabricated steelwork have been sourced from an estimating guide and

benchmarked against discussions with local fabricators.

21.2.1.2 BUILDINGS

Building cost is predominantly for a laboratory. Budget pricing has been received for the equipment,

with a rate for building construction from an estimating guide applied the area advised by the supplier.

21.2.1.3 ELECTRICAL AND INSTRUMENTATION

Preliminary BOQ’s have been developed for the electrical and control infrastructure envisaged to

service new equipment as per the mechanical equipment list. Major costs are based budget

quotations from suppliers, and the balance on GBM database figures.

Allowance has been made for the refurbishment of relevant electrical equipment, with some input

from local service providers.

21.2.1.4 PLATEWORK

A draft platework BOQ has been prepared based on preliminary sizing of major tanks. A tonnage

based contingency has been allowed for miscellaneous platework items. Costs for platework are

based on database figures.

21.2.1.5 NEW EQUIPMENT - MECHANICAL EQUIPMENT COSTS

The Mechanical Engineering Costs have been sourced using budget quotations for major equipment.

Commercial databases and in-house historical data have been used for the balance of plant

estimating.

21.2.1.6 FREIGHT FORWARDING

An allowance of USD 100 000 for freight forwarding has been included in the estimate, which is

equivalent to approximately 50 trucks from continental Europe.

21.2.1.7 BALANCE OF PLANT COST ESTIMATES

For the balance of the plant, costs have been estimated using both quantity estimates and factors of

Mechanical Equipment costs. These cost centres and their associated factors are listed in Table 21-6.


The factors are sourced from estimating resource (13) with modifications based on comparison with in

house database estimates for which more detailed level costing exercises have been conducted, site

construction techniques and most importantly consideration for the existing infrastructure on site.

Table 21-6: Direct Cost Centre Factors

Cost Centre Factor Comments

Earthwork 2 % Most equipment inside existing building

Civil 3 % Most equipment inside existing building

Initial and Commissioning Spares 9 % Considers locality of service providers (Metso etc)

Mechanical Installation 8 %

These factors have been modified to consider the extent of the existing infrastructure.

21.2.1.8 EXISTING EQUIPMENT – REFURBISHMENT COST

Refurbishment costs have been developed on an equipment level basis and mostly are derived from

an allowance methodology. Local labour rates, known cost of spares and expert experience have

been considered as part of this approach.

21.2.1.9 TAILINGS MANAGEMENT FACILITY COST ESTIMATION

Tailings Management Facility evaluation, design and costing has been undertaken by Tailings

Consultants Scandinavia AB (TCS).

The cost (CAPEX and OPEX) for works related to the TMF and the clarification pond has been

estimated using the following method:

The required available capacity of water in the clarification pond has been estimated to

approximately 120 000 m3.

The required storage capacity in the TMF has been estimated to 2.3 Mt, however only the

phases to contain 1.6 Mt are included.

Volume of tailings has been calculated based on the dry density of tailings samples taken

of the deposited tailings.

Dams and their required heights have been modelled using an assumed slope of the

tailings beach (based on references from similar sites in Sweden). To compensate for

variations in deposition, a concave beach profile, uncertainties in beach slope and

uncertainties in dry densities the required volume has been increased by 20 % as a safety

margin.


A preliminary dam design, following Swedish guidelines, has been used to estimate the

required volume of fill materials.

The characteristics of the tailings regarding the upstream construction have not been

considered at this stage, i.e. the grain size distribution and the acidic properties of the sand.

Cost for dams have been calculated using reference values of cost for fill materials from

other TMF projects in Sweden.

To compensate for uncertainties in material cost and material volumes (the later related to

unknown conditions regarding excavation depth to reach till of acceptable quality – here

assumed to be 1 meter) calculated costs are increased by 20 %.

Additional infrastructure has been calculated as follows:

Spillways and channels: Estimated cost for either spillway solution.

Pipes and pumps (slurry deposition): A preliminary design of pipelines and

pumps is available, but for a lower yearly tonnage (reference costs taken from

suppliers.) The cost of slurry deposition in the final stage of production has

been based on these references. An overall upgrade of the deposition system

once in the lifetime of the mine has been included. Pumps are not included.

Dam safety monitoring: Estimation based on preliminary design.

Operating costs are mainly based on estimated cost for electricity and required

operators.

Contingency has been included in the quantities and therefore no overriding contingency is applied to

Tailings costs.

21.2.1.10 MINING CAPITAL COST ESTIMATES

Golder did not undertake an estimate of the mine capital cost estimates assuming that the existing

infrastructure was adequate to execute the mine plan. The contractor supplied mobile and fixed plant

were addressed through the mine operating cost estimates.

21.2.2 INDIRECT COST DEVELOPMENT

Definitions of the project indirect costs are given below in Table 21-7.

Table 21-7: Indirect Cost Centre Definitions

Indirect Cost Centre

Definition

EPCM Costs for engineering, procurement activities, construction management, project management, site mobilisation, site establishment and services during construction.

Contingency A cost element of the estimate which covers the uncertainty and variability associated with the estimate. The method of defining the contingency is described in section 21.1.10


Indirect Cost Centre

Definition

Owners Cost

Costs cover training, IT systems (software), owner project team including their flights, accommodation and per diem whilst on EPCM or OEM visits, client administration personnel, any translations to local language and socioeconomic matters.

Field Indirect

Lump sum costs for items such as geotechnical works, test work and surveying. This could also include site visits by OEM and EPC contractors for bidding purposes during the EPCM tender process if not covered in the EPCM contract.

The costs factors have been applied to the direct costs to generate the indirect costs.

Table 21-8 - Indirect Cost Factors

Cost Centre Factor Comments

EPCM - Estimated – an hours allowance for high level design and supervision assistance is included

Owner's costs - Estimated by Woxna

Field indirect 1 % Considers brownfield site

Consultants 1 % Tailings consultants allowed for elsewhere

Insurance 0 % Assumed

Contingency 20 % Section 21.1.10

21.2.3 WORKING CAPITAL

Allowance has been made for the cost of the purchase of initial inventory and cost of operations until

revenue from product sale is received. 3 months of operational cost has been used to estimate this

figure.

In addition, allowance has been made for insurance spares and consumables on a factored basis.

21.2.4 DEFERRED AND SUSTAINING CAPITAL

There are a number of capital items that have been identified suitable to defer expenditure, or

required to sustain of operations during the LOM. These are summarised in the following table.

Table 21-9: Deferred Capital Investment

Cost Centre Total Year 4 Year 7 Year 13

Deferred Capital Investment 7.89 0.08 1.81 6.00


21.2.4.1 POWER LINE RELOCATION

As described in Section 16.9.2, the proposed pit shell narrowly crosses the path of the existing power

line to site. An allowance has been made in Year 4 to relocate a portion of this line so the mining

operation can advance.

21.2.4.2 TAILINGS EXPANSION

In Year 8, the TMF will reach its Stage 1 design capacity of 1.0 M t. Therefore, the Stage 2 expansion

will be completed in the summer of Year 7. Only the cost for preparation of the fill is included in the

capital estimate as it is assumed that the material will be sourced from and delivered by mining.

21.2.4.3 CLOSURE COST

An order of magnitude estimate has been prepared for the costs of closure of the mine and

associated infrastructure by Woxna. The estimate considers the following key elements:

Sealing of tailings and waste dumps, and covering of a protective layer. Material is

assumed to be locally sourced.

Demolition of the plant and permanent facilities. Rubble will be disposed of in the TMF and

the cost of demolition will be predominantly be covered by the sale of scrap metal.

Decommissioning of the open pit by treatment of pH and safety measures such as access

blocking and signage.

21.3 OPERATIONAL COST

The operational costs are those incurred during the full operation of the mining, process plant and

surrounding infrastructure for the duration of the operation. The operating costs are in reported SEK

since all operating costs are paid in local currency. As almost all constituents of operating cost are

borne in local currency, this is reflective of actual operations and allows for sensitivity of exchange

rate to be performed on the economic model (see Section 22.3). A USD conversion is shown for

clarity.

Unit operating costs have been based on the process plant operating as per Table 21-10.

Table 21-10: Operational Costs

Item Value Unit LoM USD/t ROM Ave

LoM USD/t Graphite Ave

Mining Cost

Ore 26.3 SEK/t ROM 25.8 240

Waste 27.1 SEK/t Waste


Item Value Unit LoM USD/t ROM Ave

LoM USD/t Graphite Ave

Fixed 800 000 SEK/a

Process Cost

Reagents and Consumables 12.6 SEK/t ROM 1.9 18

Labour 22 320 000 SEK/a 23.4 217

Grinding Media and Liners 12.1 SEK/t ROM 1.9 17

Power 26.3 SEK/t ROM 4.0 37

General & Administration costs 3 525 600 SEK/a 3.7 34

Fuel 27.3 SEK/t ROM 4.2 39

Maintenance 34.4 SEK/t ROM 5.3 49

Tailings Management 6.8 SEK/t ROM 1.0 10

Sales Cost

Transport 325.5 SEK/t product 5.4 50

Broker Fee (% fee / % traded) 7.5 % / 20 % 1.9 18

Total 78.5 730

For the nominal plant design, costs are calculated according to the nature in which they will be paid.

Some costs will be incurred equally irrespective of throughput and others will be borne only on

production tonnages (depended on recovery). Average values shows are based on LOM production

averages.

The proportions of each operating cost component are shown in Figure 21-1.


Figure 21-1: Plant Operating Cost Proportions

The follow sections describe the methodology and information source for each component.

21.3.1 OPEX COST DEVELOPMENT

The following sections describe the development of the various operating cost components.

21.3.1.1 MINING OPERATING COST ESTIMATES

The mine operating costs (OPEX) were estimated by Golder from contractor quotes from mining

contractors received specifically for the Woxna Graphite Restart Project. Key contractor mining

operating costs are summarised in Table 21-11 for contractor mining fleet.

Due to the variable nature of waste and PEM ratios, the mining cost changes on a yearly basis, so

totals are based on a nominal strip ratio of 5:1.

Table 21-11: Summary of Contractor Operating Costs

Item Unit SEK Qty p/a SEK/Waste

tonne SEK/ROM

tonne SEK/t total

Overburden removal SEK/m3 36.5 0 0.0 0.0 0.0

Waste rock excavation SEK/t 27.1 500 000 27.1 135.5 22.6

PEM (ROM) excavation SEK/t 26.3 100 000 5.3 26.3 4.4


Item Unit SEK Qty p/a SEK/Waste

tonne SEK/ROM

tonne SEK/t total

Surveying, dewatering (Overburden) SEK/t 0.1 0.6 0.1

Roads SEK/mon 0.8 4.1 0.7

Lighting electricity SEK/mon 0.4 3.2 0.5

Lighting rental SEK/mon 0.4 1.8 0.3

Mine planning grade control SEK/a 1 000 000 1

Overburden has been included this cost with the waste rock excavation, from contractor

quotes

Waste rock excavation includes drilling, blasting, loading and hauling to waste dumps from

contractor quotes

PEM excavation includes drilling, blasting, loading and hauling to ROM pad from contractor

quotes

Dewatering is estimated by experience and surveying is not included

Allowance for road maintenance is estimated by experience and based on the equipment

hourly rate list provided by contractor.

Allowance for lighting power estimated from experience.

Cost provided by contractor for 4 lighting plants.

Woxna has included an allowance of SEK 1 M per year for mine planning and grade control

costs

21.3.1.2 REAGENTS AND CONSUMABLES

Reagent consumption rates for the process plant are taken from the mass balance, which has been

calculated using historical operating data, available test work or industry norms for a plant of this type

and size. The annual reagent consumption costs have been determined using calculated

consumption rates and prices per tonne of reagent based on database prices. Reagent costs are be

expressed as a SEK/t value and applied to the plant feed tonnage.

In addition to those related to wear and maintenance, the major consumable allowed for is product

bags, the cost of which is based on database figures.

21.3.1.3 LABOUR

The required personnel have been estimated by GBM in conjunction with Woxna Graphite. The

manpower requirements are outlined in the organisational chart (Figure 21-2). With consideration for

Swedish labour laws, to operate a under continuous conditions (24 / 7), 5 persons are required to

cover 1 position.


Figure 21-2: Proposed Organisational Chart

For the PEA, the roles have been broken into 5 manpower rate categories based on the

organisational level of the role. The rate for each category has been Client supplied and is inclusive of

all benefits and represents the cost to the company.

Labour costs will be expressed as a SEK/a value applied against the operating period.

21.3.1.4 GRINDING MEDIA AND LINERS

Annual grinding media consumption is estimated using the rate of consumption in kilograms per

tonne, and the cost per tonne of grinding media. The rate of rod consumption is based Aminpro

(metallurgical test work provider) and compared with benchmarked rates.

Ceramic regrind media consumption and cost was supplied by Metso.

Annual costs of the liners for the mills have been calculated using the wear rates for each mill liner in

grams per tonne. These rates have been sourced from database figures. This rate is applied to the

plant feed tonnage and multiplied by the cost of liners in SEK/t.


21.3.1.5 POWER

The electrical power operating costs have been estimated from the average power consumption

expected during normal plant operation. The average power consumption is determined by load

calculations at all project sites including the process plant, village and all infrastructure facilities. The

load calculation is based on estimated loads for the major equipment, and allowances for lighting and

small power. Reduced loads from duty/standby and intermittent motors are taken into account and

maximum demand diversity calculations are used to allow for the intermittent nature of the lighting

and small power loads. A summary of the power requirements of the plant is shown in Table 21-12.

Table 21-12: Summary Power Requirements

Item Values

Maximum Demand (kW) 2070

Average Power Demand (kW) 1340

Estimated site power factor 0.85

Largest motors to be started (kW) 350 (Rod Mill, VSD)

Existing Process Equipment Average Power Demand (kW) 477

Distribution Voltages on site (V) 400

The electrical consumption per annum is multiplied by a number of fixed and variable tariffs as

provided by Elektra, the local supply authority. At the average demands, the average cost of power is

0.50 SEK/kWh.

Equipment name plate kW’s have been used as the reference figures, however, as a large proportion

of the equipment was purchased second hand, the actual motors maybe larger than necessary and

hence the running loads of this equipment maybe overstated using the assumed factors.

21.3.1.6 TRANSPORT

The annual costs for the transport of product from site to Germany have been estimated using a rate

per trucked load, as determined through discussions with industry brokers. It has been assumed that

each truck has a 22 t load, providing a calculated rate in SEK/t.

Due to the graphite sale pricing assumption being under a ‘CIF European Port’ basis, Woxna has

conducted preliminary discussions with off takers, and believe a 30 EUR/t premium will be obtained

for delivered product. To simplify the additional revenue, a transport credit to the full transport cost

has been applied as part of the OPEX.


21.3.1.7 GENERAL ADMINISTRATION

The yearly general administration costs including the process plant, mine and ancillaries were based

on predominantly Client input, which are heavily influenced by current and historical operational costs.

The GA cost is applied at a fixed rate (SEK/a), regardless of the feed tonnage of the plant, or amount

of product produced.

21.3.1.8 FUEL AND LUBRICANTS

Diesel consumption for the dryer is estimated using a rate of consumption in litres per tonne (L/t)

which is applied to the plant feed tonnage. The rate of consumption has been estimated from the

equipment manual.

Diesel consumption for the loader has been estimated using the rate of diesel consumption provided

by the supplier in SEK/h. This rate is applied to the equipment operating hours as estimated.

Lubricant consumption is included in the maintenance rate.

21.3.1.9 MAINTENANCE

The annual maintenance costs cover the repair and replacement of piping, electrical components and

cabling, hardware, general wear and tear and spare parts. The values are based on GBM experience

and are applied to the replacement cost of mechanical equipment installed in the plant. A slightly

higher maintenance cost allowance is made for the reused equipment.

Due to the limitations of the organisational structure proposed, also included in the maintenance costs

is allowance for contractors. This annual cost has been determined by estimating the hours required

for shutdowns, specialist works, electrical works and call outs. This requirement in h/a is multiplied by

the Client supplied contractor rate in SEK/h.

21.3.1.10 TAILINGS AND ENVIRONMENT

Environmental maintenance is made up of allowances for materials and consumables required for the

upkeep of the TMF and discharge of excess water to the environment. These costs have been

estimated by TCS and are expressed in SEK/a.


SECTION 22 ECONOMIC ANALYSIS

This report contains certain "forward-looking statements" and "forward-looking information" as defined

under applicable Canadian and U.S. securities laws. Forward-looking statements are based on

forecasts of future results, estimates of amounts not yet determinable and assumptions that, while

believed by management and the Consultant to be reasonable, are inherently subject to significant

business, economic and competitive uncertainties and contingencies.

The base case financial model prepared is a constant dollar type model which assumes the

purchasing power does not change with time. This means the cost of capital, operating costs and

revenue are constant through time in a like-for-like manner. A scenario analysis is undertaken to

assess the impact of time varying changes cost and revenue drivers.

METHODOLOGY 22.1

Table 22-1: Depreciation and Loss Parameters

Parameter Method

Asset depreciation method straight-line

Depreciation carry-forward infinite

Operating loss carry-forward Infinite

Parameter Value Unit

Fixed Plant, Equipment & Infrastructure 5 / 20 years / % per annum

Asset End-of-Life Salvage Value 0 USD

Table 22-2: Base case financial inputs

Description Value Unit Reference

Corporation tax rate 22 % Client

Discount rate 10 % Client

Finance rate 0 % A

Royalty rate 0 % Client *

Base currency code USD - Client

Base case graphite price 1 199 USD/t See § 19.6

A royalty rate of 0.2 % is commonly applicable to mining projects in Sweden, however as the mining

license was granted after 2005, Woxna have advised that no royalty is payable on the revenues from

Kringelgruvan nr1.


GBM believes the depreciation and taxation methods are applied appropriately to this project. GBM

however are not accounting professionals and the post-tax economic performance should be taken as

a guide.

22.1.1 CARRIED FORWARD DEPRECIATION AND LOSSES

As Woxna Graphite have existing infrastructure and have operated historically, some depreciation

carry forward and historical losses have been incorporated into the economic model as summarised

in the following Table 22-3.

Table 22-3: Historical Book Assets

Item Value Unit

Operating Loss 2.75 M USD

Depreciation 3 years 0.20 M USD (total)



Total 8.70 M USD

22.2 ECONOMIC PERFORMANCE

The following sections describe the results of economic modelling for the production schedule under

the capital and operating cost, economic parameters as described in the relevant sections.

Table 22-4: Economics Summary


LOM 13.0 years

Average Sale Price 1 199 USD/t graphite

Revenue 121.5 USD/t ROM

1 130 USD/t graphite

OPEX 71.1 USD/t ROM

662 USD/t graphite

Cost of Sales 7.4 USD/t ROM

68 USD/t graphite

Margin 50.4 USD/t ROM

468.6 USD/t graphite

Payback Period 3.86 years



IRR 34.0%

NPV 26.61 M USD


22.3 DISCOUNTED CASH FLOW ANALYSIS

The full discounted cash flow schedule is shown below.

Table 22-5: Discounted Cash Flow Analysis

Item LoM Y 0 Y 1 Y 2 Y 3 Y 4 Y 5 Y 6 Y 7 Y 8 Y 9 Y 10 Y 11 Y 12 Y 13

CAPEX 24.62 16.72 - - - 0.08 - - 1.82 - - - - - 6.00

OPEX 135.1 - 7.76 10.54 11.20 11.52 11.59 11.29 11.35 11.37 11.13 11.37 9.34 9.20 7.38

Revenue 230.7 - 12.76 18.58 16.65 18.45 18.53 18.16 19.05 19.49 18.86 19.49 19.17 19.15 12.35

Sales 244.6 - 13.53 19.70 17.65 19.56 19.65 19.26 20.20 20.67 20.00 20.67 20.33 20.31 13.10

Cost of Sales 13.96 - 0.77 1.12 1.01 1.12 1.12 1.10 1.15 1.18 1.14 1.18 1.16 1.16 0.75

Annual Operating Income 95.63 - 5.00 8.04 5.45 6.92 6.94 6.87 7.69 8.12 7.73 8.12 9.83 9.94 4.97

Annual Depreciation Claimed 21.18 - 3.65 3.65 3.65 3.59 3.60 0.16 0.16 0.52 0.52 0.51 0.51 0.51 0.14

Annual depreciation allowance - - 3.65 3.65 3.65 3.59 3.60 0.16 0.16 0.52 0.52 0.51 0.51 0.51 0.14

Depreciation carry forward - - - - - - - - - - - - - - -

Working capital take out -2.38 - - - - - - - - - - - - - -2.38

Taxable Operating Income 76.83 - 1.34 4.39 1.80 3.34 3.34 6.71 7.53 7.60 7.20 7.61 9.32 9.44 7.21

Taxes 16.30 - - 0.65 0.40 0.73 0.73 1.48 1.66 1.67 1.58 1.67 2.05 2.08 1.59

Corporations Tax 16.30 - - 0.65 0.40 0.73 0.73 1.48 1.66 1.67 1.58 1.67 2.05 2.08 1.59

Loss Carry Forward - 2.75 1.41 - - - - - - - - - - - -

Loss Carry Forward Claimed - - 1.34 1.41 - - - - - - - - - - -

Annual Post Tax Income 60.53 - 1.34 3.73 1.40 2.60 2.60 5.24 5.88 5.93 5.62 5.94 7.27 7.36 5.62

Annual Cash Income 81.71 - 5.00 7.38 5.05 6.19 6.21 5.40 6.04 6.45 6.14 6.45 7.78 7.87 5.76

Annual Cash Flow 57.09 -16.72 5.00 7.38 5.05 6.11 6.21 5.40 4.22 6.45 6.14 6.45 7.78 7.87 -0.24

Cumulative Cash Flow -16.72 -11.72 -4.34 0.72 6.83 13.03 18.43 22.65 29.10 35.24 41.68 49.46 57.33 57.09


22.4 SENSITIVITY ANALYSIS

A scenario analysis was conducted to estimate the magnitude of the sensitivity of the model as measured by NPV. The parameters utilised and results of the

analysis are described in Table 22-6. The sensitivities are show also graphically in Figure 22-1.

Table 22-6: Sensitivity Analysis Results

% Change Value NPV IRR % Change Value NPV (USD) IRR

Exchange Rate (SEK:USD) CAPEX (M USD)

70 % 4.58 394 002 8.4 % 70 % 11.7 32 182 639 49.5 %

80 % 5.23 11 624 751 20.3 % 80 % 13.4 30 326 762 43.1 %

90 % 5.89 20 021 702 28.3 % 90 % 15.0 28 470 885 38.1 %

100 % 6.54 26 615 008 34.0 % 100 % 16.7 26 615 008 34.0 %

110 % 7.19 32 009 531 38.4 % 110 % 18.4 24 759 131 30.5 %

120 % 7.85 36 498 331 42.0 % 120 % 20.1 22 903 254 27.5 %

130 % 8.50 40 292 041 44.9 % 130 % 21.7 21 047 377 25.0 %

Labour (M USD Avg) Fuel (M USD Avg)

70 % 2.39 32 473 631 39.1 % 70 % 19.1 27 657 102 34.8 %

80 % 2.73 30 520 757 37.4 % 80 % 21.8 27 309 738 34.5 %

90 % 3.07 28 567 882 35.7 % 90 % 24.6 26 962 373 34.2 %

100 % 3.41 26 615 008 34.0 % 100 % 27.3 26 615 008 34.0 %

110 % 3.75 24 662 134 32.2 % 110 % 30.0 26 267 643 33.7 %

120 % 4.10 22 709 260 30.5 % 120 % 32.7 25 920 279 33.4 %

130 % 4.44 20 756 385 28.8 % 120 % 32.7 25 920 279 33.4 %


Mining (USD/t ROM) Price (USD/t product)

70 % 18.1 32 979 398 39.1 % 70 % 839 -6 460 785 -0.1 %

80 % 20.6 30 857 935 37.4 % 80 % 959 4 994 425 13.7 %

90 % 23.2 28 736 471 35.7 % 90 % 1 079 15 938 007 24.7 %

100 % 25.8 26 615 008 34.0 % 100 % 1 199 26 615 008 34.0 %

110 % 28.4 24 493 545 32.2 % 110 % 1 319 37 283 748 42.6 %

120 % 31.0 22 372 081 30.4 % 120 % 1 439 47 934 355 50.8 %

130 % 33.5 20 250 618 28.5 % 130 % 1 558 58 584 191 58.7 %

Recovery (%)

80 % 9 720 410 18.6 %

85 % 15 071 553 23.9 %

90 % 20 327 644 28.6 %

96 % 26 615 008 34.0 %

98 % 28 710 796 35.7 %


Figure 22-1: Sensitivity Analysis


SECTION 23 ADJACENT PROPERTIES

There are no known operators of any relevant activities on directly adjacent properties or locally

adjacent properties.


SECTION 24 OTHER RELEVANT DATA AND INFORMATION

24.1 EXECUTION SCHEDULE

Due to the small scale of the project and advanced nature of the infrastructure, and readiness of the

operating team at the existing site, project execution is expected to be comparatively rapid.

The following schedule presents an execution schedule indicative of the timing believed achievable

for the Woxna Restart Project.

Table 24-1: Indicative Execution Schedule

Whilst a winter construction period is limiting the advancement of tailings construction (and hence

places it on the critical path), it should be considered that engineering and procurement lead times

would prevent any significant improvement in the overall schedule beyond the 11 months shown.


SECTION 25 INTERPRETATION AND CONCLUSIONS

25.1 MINERAL RESOURCE ESTIMATES

A Mineral Resource estimate, using an IDW interpolation method, was completed by Reed Leyton.

The Mineral Resource estimate in this Technical Report is reported using cut-off grades which are

deemed appropriate for the style of mineralization and the current state of the Mineral Resources.

Reed Leyton considers the estimated Mineral Resource to be in accordance with NI 43-101

Guidelines for Resource Estimates. Of importance for mine planning, the model accommodates in situ

and contact dilution but excludes mining dilution. Block size is similar (5 m x 25 m x 5 m) to expected

small-mining units conventionally used in this type of deposit, and appropriate for an open pit mine.

It is the opinion of Reed Leyton that the Mineral Resources estimates for Kringelgruvan satisfy the

definition of Mineral Resource as per the CIM Definition Standards of June 2011 (became law) or

November 27 2010 (published).

Potential for increasing of the Mineral Resources are good, with mineralization open down dip, which

requires further drilling to investigate potential.

25.2 MINING METHODS

The Kringelgruvan graphite deposit, located in Central Sweden, is at an advanced stage of

exploration. Drilling and geological resource block modelling to date has defined a resource that

forms the basis for the mining methods section of the preliminary economic analysis. The purpose of

this section of the PEA was to meet the scope of work as requested by Woxna.

Pit optimisation using Gemcom Whittle™ 4x optimisation industry standard software was completed

for Kringelgruvan deposit. The pit was analysed for optimal pit limits and economic sensitivity and

demonstrated that the deposit could be economically exploited based on the financial and operating

assumptions used in this study. The optimisations included only Measured and Indicated mineral

resources categories within the provided geological models.

The LOM schedules were prepared to support the preliminary economic evaluation of the potential of

Kringelgruvan deposit to develop into a feasible mine. If the preliminary economic analysis shows

encouraging cashflow further work is justified to proceed toward a pre-feasibility/feasibility study.


25.3 METALLURGICAL TEST WORK AND PROCESSING

The test work performed by Aminpro, backed up by historical production data, has been used to

predict the production of graphite concentrates. Variability samples were also assessed by Aminpro

and exhibited little difference in processing kinetics.

There are significant CAPEX savings arising from utilising the existing supporting plant infrastructure.

Importantly however, the majority of the core processing units will be purchased new for the project.

This reduces process uncertainty, as the equipment can be specified ‘fit for purpose’ with test work to

support the purchase. The major plant units to be reused are the dryer and the rod mill, and both have

some flexibility in throughput.

25.3.1 EXISTING PLANT

The extent of the established plant infrastructure (particularly the supporting utilities) is a major

strength of the project. The success of recent operations at site involving reprocessing stockpiled

product further increases confidence in the adequacy of the legacy facilities.

Significant parts of the process critical parts of the plant utilise existing equipment. A refurbishment

program is planned and considerable work has been done on site by way of inspections to support

the costs reported.

25.4 ORGANISATIONAL STRUCTURE

The organisational structure proposed utilises a small crew on the day to day operations. Whist this is

not unique for operations this size, the low level of plant automation mean that the operators will be

required to be attentive to the whole plant operation. Failure to notice some processing anomalies

could impact graphite recovery for extended periods of time.

As the organisational structure has no maintenance personnel, operators will be required to maintain

as well as operate the plant where possible. Contract maintenance personnel will be used for major

maintenance.

25.5 TAILINGS REFURBISHMENT AND EXPANSION

The design proposed above is presented as an upstream dam, as this this design is approved under

the current permit and suits the existing dam wall construction. The present design is also the

cheapest one due to the fact that tailings are used as construction material. Hence there is no need

for a large amount of construction material for every expansion.

There are a number of other alternatives that may be more suitable to the operation under certain

circumstances. These include downstream dam, thickened tailings disposal or dry-stacked tailings.


25.6 PRODUCTION AND GRAPHITE SALE

As the average sale price has been calculated through a combination of various product recovery

grades, sizes (ant their relative proportions) and research house prices, assessing product pricing for

graphite is more difficult than exchange traded commodities.

The production reports of a comparatively small volume of product from historical operations give

further credibility to the assessment of revenue.

25.7 INFRASTRUCTURE

The existing infrastructure at the mine site is a major strength of the project. Very little is required to

be constructed or refurbished in order to bring the mine and process plant back into production.

Those improvements that are proposed (such as the replacement transformer) will serve to improve

the reliability of the plant.

25.8 FINANCIAL PERFORMANCE

The financial performance of the base case indicates a positive performance of the project.

The total initial investment capital cost estimate for the required infrastructure, mining and mineral

processing is USD 16.7 M including contingency. The LOM capital cost estimate is USD 24.6 M

including contingency and closure cost.

Based on the capital investment and operating cost estimation the Project will yield an internal rate of

return (IRR) of 34.0 % and a net present value (NPV) of USD 26.6 M under the mine production

schedule proposed.

The project has been evaluated using a weighted average graphite sale price of 1 199 USD/t of

graphite sold from the Woxna process plant.

The operating costs have been estimated at a LOM average of 78.5 USD/t ROM, or

730 USD/t graphite sold.

Economic performance is most sensitive to price. The project is next sensitive to the US dollar to

Swedish krona exchange rate as revenue is indexed to the US dollar whist OPEX is borne in Swedish

krona.

This economic performance would typically be described as attractive and a signal that the Board

should consider the project for further investigation.


25.9 RISKS AND OPPORTUNITIES

25.9.1 RISKS

The mine production schedule is aggressive with respect to the number of benches that are

mined per annum. This is being driven by the stripping ratio coupled with the need to release

graphite mineral. To achieve this sinking rate, bench heights may have to be increased which

could potentially create additional dilution and impact mining recovery;

The size and location of the waste storage area needs to be confirmed through a

geotechnical investigation. This activity should also assess the annual maximum lift heights,

the batter and final slope angles

Waste characterisation has not been completed and the extent of any acid problem is not

known.

The mine economics and financials are based on a set of conditions including prices and

costs that may change and adversely affect the commercial case for the project

No geotechnical stability analysis has been conducted on the pit design, waste dump stability

and foundation conditions. A geotechnical analysis should be conducted to determine safe

design parameters for these facilities as well as, the ultimate open pit slopes wall geometries.

Operating in severe winter conditions could affect the efficiency of the mining, crushing and

transport. The mine has operated in winter before and the historical production records do

not indicate that it was a problem to operate during the winter.

Commissioning of the flotation plant might take longer than anticipated to produce the

required concentrates. The flotation plant is sized for 155,000 t/a and hence there is excess

capacity to either allow for additional flotation time or to catch-up production. In addition, lower

grade concentrates could either be sold at the best price for that grade or recycled back to the

regrind and flotation circuits.

Failure of the existing, refurbished, equipment. Commission existing ball mill to replace rod

mill at lower throughput. Stockpile wet or dry concentrate until dryer or concentrate handling

plant repaired/replaced.

Whilst effort has been made to investigate critical items, there is some risk that the

refurbishment program will identify problems that the cost estimate did not allow for. In

addition to the CAPEX risk comes lead time risk, with certain components such as mill

gearing requiring manufacturing time in excess of the project execution timeline.

In addition to CAPEX risk, failures that might occur during start-up and operations present

potentially more serious implications. Due to the low operating margins, extended ramp-up

delays may result in cash flow problems.


An 85 % availability has been assumed, which is lower than a usual mill process to

accommodate this however, the absence of an owner maintenance team and potential for

long outages could quickly bring the average plant availability below this number.

Several risks and opportunities have been identified in the preparation of this estimate as a

result of the estimating methods:

o The use of database costs may introduce errors or omissions in failing to allow for

site specific requirements.

o No allowances have been made for the event of a labour stand-down or other

significant interruptions to work.

o As this is a PEA level study the level of engineering is only sufficient to estimate the

costs to within ± 30 %.

25.9.2 OPPORTUNITIES

Owner operated mining operating and capital costs should be evaluated against the

contractor quotes to assess mine economics;

In-pit waste rock storage should be examined in greater detail;

Increasing the maximum depth of the open pit would increase the productive life of the

operation and increase the recovery of the known in-situ resources; and

A higher production rate would see better financial returns on the project, especially if the

mine is owner-operated rather than operated by mining contractors.

An existing mill could be used as a ball mill in the primary grinding circuit. This would allow an

increase in throughput and improve grinding efficiency. However, the limiting factor could be

the capacity of the dryer and bagging plant.

A tailings thickener could be used to improve water usage and potentially assist with tailings

disposal. Tests have been carried out and indicate that it should be possible to thicken the

tailings to approximately 65 % solids. The water would be recycled. Further tests are

required to assess high rate thickening. Additional assessment of the overall water balance of

the project is required.

It might be possible to coarsen the primary grind to produce a coarser graphite concentrate,

that is, greater than 300 µm. This product would have a premium on the 250 µm product.

Further test work is required to assess the possibility of producing a 300 µm concentrate at

94 % C, and the effect on the overall graphite recovery, flowsheet and economics.

The throughput could be increased by improving the plant availability from the estimated

85 %. This could be done by improving maintenance and hence reducing downtime of

equipment

Whist not currently permitted, processing strategies such as leaching to achieve synthetic

graphite have been shown in past to show some compatibility with the Woxna graphite


products. According to Industrial Minerals (14), the price is highly sensitive to very small

variations (that approach 99.95 % C) of the grade. As such, comprehensive testwork and

study would be required to justify a decision to pursue this processing strategy.

There are a number of other alternatives that may be more suitable to the operation under

certain circumstances. These include downstream dam, thickened tailings disposal or dry-

stacked tailings.


SECTION 26 RECOMMENDATIONS

26.1 MINERAL RESOURCE AND RESERVE ESTIMATES

Even though no inferred resources are considered in this project, an in-fill drilling program to increase

the Mineral Resource confidence categorization of areas currently defined as Indicated to Measured

should be completed. Reed Leyton estimates an additional 3 000 m of in-fill and extensional drilling

would be recommended, tightening the drill spacing to 25 m sections and infilling some sections to

25 m spacing to confirm inter-hole continuity in and around faulted zones. Deep drilling to ascertain

the depth of the Kringel graphite is also recommended.

The following table presents some indicative costs for such work.

Table 26-1: Indicative Resource Development Costs

Follow Up Drilling Unit Price Cost (CAD)

15 x 200 m DDH on infill sections 120 CAD/m 360 000

1 x 300 m DDH 150 CAD/m 45 000

Geology, logging, core cutting, support 215 000

TOTAL 520 000

The deposit shape and surrounding areas show evidence of additional resources beyond those

defined. Further exploration work should be conducted to increase the size of the resource.

26.2 METALLURGICAL TEST WORK

If significant quantities (more than several kilograms) concentrate samples are required for marketing

purposes then a pilot plant is required. This will require several tonnes of bulk sample.

Test work needs to be carried out on neutralising acid rock drainage to enable optimising the design

of the neutralising circuit.

Neutralising test work is required on the water in the TMF to properly specify and design the process

and equipment for recycling the water to the plant.

It is believed a coarser concentrate, that is, greater than 300 µm, would improve the economics.

Further laboratory test work is required to assess the possibility of achieving the coarser concentrate.

The effect on grades and recoveries for all fractions needs to be assessed, as well as an appraisal of

the effect on the flowsheet and equipment sizing. A pilot plant test would be required to produce

coarser samples for marketing purposes under the new conditions.


26.3 MINING

26.3.1 EXPLORATION RECOMMENDATIONS

It is recommended drilling to the extent necessary be conducted with a focus on upgrading the

Potential and Inferred Resources to the Indicated and Measured categories and evaluating additional

nearby exploration targets that could add mineral resources to the project.

26.3.2 MINE DESIGN AND PIT STABILITY GEOTECHNICAL STUDIES RECOMMENDATIONS

The pit limits are based on mineral resources reported by Reed (2013) however the pit design

incorporates limited geotechnical information regarding the rock strength or any structural

orientations. The pit walls were assumed at 55˚ for two benches and then 65˚ for the remaining

design for all the walls in both waste and mineralisation. To complete a design for preliminary

feasibility study additional geotechnical information needs to be gathered and analysed.

The following actions are recommended in order to increase confidence in the pit and other earthwork

geotechnical design parameters:

A campaign of bedrock exposure mapping and core logging with the sole purpose of

capturing the geotechnical properties of the rock masses in and around the pit designs;

The benches exposed by previous mining activities should be mapped and logged, to gather

data on the orientation and spatial distribution of the joint sets and faults, such data will form a

key component of the structural modelling and can be performed with minimal expenditure;

Develop a geotechnical model of the site to determine the pit wall structural sectors and to aid

in determining suitable sites for waste storage and other infrastructure;

Assess the blasted and fragmented rock properties to assist in developing the waste rock

storage facility designs and staging plans;

Determining the intact uniaxial and modulus of waste and graphite mineralisation material

through laboratory testing; and

During mining an observational method for fracture mapping especially for the tension

cracks should be used.

26.3.3 MINE PRODUCTION AND OPERATIONS RECOMMENDATIONS

The proposed mine design and operation should be reassessed with the availability of the information

collected for geotechnical purposes outlined in the preceding section. Recommendations related to

the mine design and operations are as follows:

The mineral resources extend below the permitted depth of 65 m (below topography). Woxna

should apply for a new permit to mine deeper;


Woxna should apply for an increase in its mine production rate which is currently set at

100 000 t/a ROM to take advantage of economies of scale associated with the proposed

equipment fleet;

The size and location of the waste disposal area is limited by the available space within the

mining licence. Either the boundaries should be adjusted or the routes and roads within the

current site need to be adjusted to accommodate the storage area used in this study;

Woxna should arrange for the timely movement of the power lines in the western part of the

ultimate pit design; and

A single drill rig was selected that would be capable of keeping drilling operations well ahead

of loading operations.

26.4 PROCESSING

Detailed engineering design is required for installing the flotation cells, the regrind mills, the filtration

plant and the neutralising equipment. This can be started during the pilot plant stage (should it be

required).

Whilst investigation work has been completed to confirm the suitability of reuse of existing equipment,

as soon as project approval is received, work should commence immediately following process sign-

off on refurbishment to ensure no long lead time items affect the overall execution schedule.

As part of the flowsheet simplification, a number of mills are now excluded from the flow sheet, though

the condition of at least one of them is quite good. Considering the under-utilisation of the crushing

plant, utilising this mill for progressive grind would greatly increase the throughput of comminution

circuit. Flotation would require expansion, but the modular nature of flotation plants will make this

process relatively simple. Additional plates can be added to the concentrate filter. The most significant

cost of further expansion will likely be seen in the bagging area, with much of the equipment already

approaching its known or estimated maximum capacity.

26.5 ORGANISATIONAL STRUCTURE

The organisational structure is quite lean GBM believes there is little scope to reduce employee

levels. On the direct (operations) labour, it will be important to employ skilled or semi-skilled operators

(sometimes called technicians) to support the lack of dedicated maintenance personnel. This will

ensure plant availability is maximised where possible and reduce the cost of contract workers, which

at present are allowed for in the OPEX on a minimal basis.

Careful staff acquisition is therefore required in the process of building the team and a clear

understanding of what duties will be undertaken by each position should be defined prior to

employment.


There is also some potential government assistance that can be gained that will lower the cost of

labour borne by the employer which should be investigated further.

26.6 TAILINGS

More geotechnical analysis is recommended to adequately asses the ground condition. This will

assist in verifying and optimising the design proposed, decrease cost over-run risk during

construction.

As part of design optimisation, further geohydrology investigations are needed in order to finalise the

design and to check if the upstream dam is a real alternative.

The desulphurisation of tailings should be investigated as it would be advantageous to use the tailings

as construction material. However this would require a separate impoundment for the sulphides and a

new permit.

If acid generation by the new tailings is considered significant then options for mitigating this will have

to be look at including floating off the sulphides from the scavenger tailings.

If thickening of the tailings is required then it is advised that laboratory tests are performed specifically

for high rate thickening either by vendors or by a laboratory acceptable to the vendors.

26.7 PRODUCTION AND GRAPHITE SALE

The results of testwork and the most recent prices available should be continually used to update the

forecast average price.

Whilst engagement with off-takers can be difficult, efforts to gauge actual prices should be continued.

Production of a bulk sample for analysis by off-takers will benefit this process.

Full analysis of the concentrates from the test work is required for marketing purposes.


SECTION 27 REFERENCES

1. Aminpro Chile. Woxna Graphite Metallurgical Test work and Front End Engineering - Final. 2013.

2. Roskill Information Services Ltd. Natural & Synthetic Graphite: Global Industry Markets and

Outlook. London : s.n., 2012.

3. GBM Minerals Engineering Consultants. 0482-SPC-001 - Process Design Criteria. 2013.

4. —. 0482-SPC-002 - Project Design Basis. 2013.

5. —. 0482-CAL-002 - Mass Balance. 2013.

6. —. 0482-DTB-007 - Capex Database - Phase 1. 2013.

7. Golder Associates. TMF Capex and Opex - Stage 1, 2 and 3.

8. GBM Minerals Engineering Consultants. 0482-CAL-012 - OPEX Calculation. 2013.

9. —. 0482-DTB-006 - Mechanical Equipment List. 2013.

10. —. 0482-CAL-013 - Electrical Load List. 2013.

11. Golder Associates. Mining Methods: Preliminary Economic Analysis Woxna Graphite Restart

Project, 12512450160-R-Rev0. 2013.

12. InfoMine USA, Inc. Mining Cost Service. Spokane Valley, WA : s.n., 2011.

13. Peters, Max and Timmerhaus, Klaus. Plant Design and Economics for Chemical Engineers.

Singapore : McGraw Hill, 1991.

14. Industrial Minerals. Graphite Prices. Industrial Minerals. [Online] [Cited: 23 November 2012.]

www.indmin/prices/prices.

15. The Ontario Securities Commission. Repeal and replacement of National Instrument 43-101

Standards of Disclosure for Mineral Projects, Form 43-101F1 Technical Report, and Companion

Policy 43-101CP, Supplement to the OSC Bulletin, Dated April 8, 2011. Toronto : Published under

authority of the Commission by Carswell, a Thomson Reuters business., 2011. NI 43-101.

16. GBM Minerals Engineering Consultants. 0482-DTB-008 - Capex Database - Phase 2. 2013.

17. ReedLeyton Consulting. o Technical Report for Kringelgruvan Graphite Deposit, Part of The

Woxna Graphite Project, Central Sweden. 2012.


18. Bonde, L. Registerblad - Område av riksintresse för naturvård i Gävleborgs län. Ovanåkers

kommun, Gävleborgs län, Sweden. . s.l. : Retrieved 05 24, 2013 from

http://www.lansstyrelsen.se/gavleborg/SiteCollectionDocuments/Sv/djur-och-natur/skyddad-

natur/riksintresse-naturvard/XN55%20%C3%96stermyroma.pdf, (1998, 11 16).

19. Heidbach, O., Tingay, M., Barth, A., & Reinecker, J. The World Stress Map database release

2008 doi:10.1594/GFZ.WSM.Rel2008. (2008).

20. kommun, O. (Ed.). Översiktsplan. Retrieved 05 24, 2013, from Edsbyn och Alfa, Ovanåkers

Kommun: http://www.ovanaker.se. (2012, 08 20).

21. Kuchta, M., & Hustrulid, W. A. Open pit mine planning & design. London: Taylor & Francis.

(2006).

22. Lawrence, R. W., & Scheske, M. A method to calculate the neutralization potential of mining

wastes. Environmental Geology, 32(2), 100-106. (1997).

23. Wyllie, D. C., & Mah, C. W. Rock slope engineering, (4th ed.). London: Institute of Mining and

Metallurgy. (2004).

24. Hoek, E. Practical Rock Engineering, http://www.rocscience.com/education/hoeks_corner. 2000.

25. Jacobsson, L. Borehole KFM01C. Uniaxial compression test of intact rock. Forsmark site

investigation, SKB P-06-69. 2007.

26. Stephansson, O. Rock Stress in the Fennoscandian Shield, in Comprehensive Rock Engineering

(ed. J.A. Hudsson), Pergamon Press, Oxford, Chapter 17, Vol. 3, pp. 445-59. 1993.

27. CostMine, Infomine USA Inc. Mine and Mill Equipment Costs. An Estimators Guide. .

28. Taxation of Cross-Border Mergers and Acquisitions. KPMG AB. Stockholm : KPMG, 2012.

29. Tailings Consultants Scandinavia. Kringelgruvan TMF - Data for PEA . 2013.


APPENDIX A LOM SCHEDULE

CERTIFICATE OF QUALIFIED PERSON

CHRISTOPHER STINTON

As the individual who has co-authored or assisted in the preparation of Sections 1, 2, 3, 13, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26 and 27 of the Technical Report prepared for Flinders Resources Limited (the “Issuer”) entitled “Woxna Graphite Restart Project, Preliminary Economic Analysis” dated effective 11 October 2013 (the “Technical Report”), I hereby certify that:

I am a Senior Process Engineer of GBM Minerals Engineering Consultants Limited (“GBM”) of Regal House 70 London Road, Twickenham, Middlesex, TW1 3QS, England.

I am a graduate of Birmingham University of Birmingham, United Kingdom, with an Honours Bachelor of Science (Minerals Engineering).

I am a Member of the Institute of Materials, Minerals and Mining (CEng). I have worked as a Professional Engineer in the mining and minerals processing industry for a total of 34 years since my graduation. My relevant experience for the purpose of the Technical Report is metallurgy, process design and engineering, plant operations and management.

I have read the definition of “qualified person” set out in the National Instrument 43-101 (“NI-43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

I have visited the Woxna Site in November 2012.

I have supervised the work carried out by other GBM professionals for GBM’s contribution to the Technical Report, and take responsibility for Sections 1, 2, 3, 13, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26 and 27 of the Technical Report.

I am independent of the Issuer applying the test set out in Section 1.5 of NI 43-101.

I have no prior involvement with the property that is subject of the Technical Report.

I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 and Form 43-101F1.

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.

Dated this 11 October 2013

Original Document signed by Christopher Stinton, BSc, MIMMM, CEng

_____________________________ Christopher Stinton, BSc, MIMMM, CEng


GEOFFREY REED

As the individual who has authored or supervised the preparation of Sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 14, 25.1 and 26.1 of the Technical Report prepared for Flinders Resources Limited (the “Issuer”) entitled “Woxna Graphite Restart Project, Preliminary Economic Analysis” dated effective 11 October 2013 (the “Technical Report”), I hereby certify that:

I am the Principal of ReedLeyton Consulting (“ReedLeyton”) of PO Box 6071, Dural, NSW, 2158, Australia.

I graduated with a degree in Geology with a Bachelor of Applied Science from the University of technology, Sydney, NSW, Australia, awarded in 1997.

I am a Member of the Australasian Institute of Mining and Metallurgy since 1998. My relevant experience for the purpose of the Technical Report is drawn from working as a geologist for a total of over 15 years since my graduation from University


I have visited the Woxna Site on June 12 to 13 2012.

I have completed the work in contribution to the Technical Report, and take responsibility for Sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 14, 25.1 and 26.1 of the Technical Report. My input is based in large part on examination of the material presented to me by Flinders Resources Limited during June 12, 2012 to October 20, 2012. First hand impressions about the style of mineralisation are based on examinations of drill core from representative drill holes during June 12 to 13, 2012.


I have had prior involvement with the Property, namely, I am a “qualified person” responsible for the preparation of the technical report titled” Technical Report for the Kringelgruvan Graphite Deposit, Part of the Woxna Graphite Project, Central Sweden” dated November 2, 2012 (the “Previous Technical Report”) which is superseded and replaced by the Technical Report I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 and Form 43-101F1.



Original Document signed by Geoffrey Reed, B App Sc, MAusIMM

Geoffrey Reed, B App Sc, MAusIMM


BRYAN PULLMAN

As the individual who has co-authored or assisted in the preparation of Sections 15, 16, 25.2 and 26.3 of the Technical Report prepared for Flinders Resources Limited (the “Issuer”) entitled “Woxna Graphite Restart Project, Preliminary Economic Analysis” dated effective 11 October 2013 (the “Technical Report”), I hereby certify that:

I am a Senior Mining Engineer of Golder Associates (UK) Limited (“Golder”) of Cavendish House, Bourne End Business Park, Cores End Road, Bourne End, Buckinghamshire, SL8 5AS, UK

I am a graduate of Queen’s University of Kingston, Ontario, Canada, with a Bachelor of Science (Engineering) in Mining Engineering.

I am a Professional Member of the Association of Professional Engineers and Geoscientists of Alberta (APEGA) (P.Eng.). I have worked as a Professional Engineer in the mining and minerals processing industry for a total of 11 years since my graduation. My relevant experience for the purpose of the Technical Report is mine design and engineering, mine operations and management.


I have not visited the Woxna Site.

I have supervised the work carried out by other Golder professionals for the Golder contribution to the Technical Report, and take responsibility for Sections 15, 16, 25.2 and 26.3 of the Technical Report.


I have no prior involvement with the property that is subject of the Technical Report.




Original Document signed by Bryan Pullman, B.Sc. (Eng.), P.Eng.

Bryan Pullman, B.Sc. (Eng.), P.Eng.


HENNING HOLMSTRÖM

As the individual who has authored or supervised the preparation of Sections 18.3 to 18.6, 25.5 and 26.6 of the Technical Report prepared for Flinders Resources Limited (the “Issuer”) entitled “Woxna Graphite Restart Project, Preliminary Economic Analysis” dated effective 11 October 2013 (the “Technical Report”), I hereby certify that:

I am the Director of Woxna Graphite AB of Skollallén 2B, Bollnäs Sweden.

I graduated with a degree in M.Sc. in Geotechnology from Luleå University of Technology, Sweden awarded in 1996 and a Ph.D in Applied Geology from Luleå University of Technology, Sweden awarded in 2000.

I am a Member of Australiasian Institute of Mining and Metallurgy (MAuIMM no 308735) as well as the Australian Institute of Geoscientists (MAIG no 5628). My relevant experience for the purpose of the Technical Report is drawn from working as a engineer and environmental scientist (several roles and positions) for a total of over 17 years since my graduation from University.


I have visited the Woxna Site regularly between 2011-2013, most recently on September 3, 2013.

I have supervised the work carried out by Golders, TCS (Tailings Consultant Sweden) and SWECO (all contracted Swedish sub-consultants) for my contribution to the Technical Report, and take responsibility for Sections 18.3 to 18.6, 25.5 and 26.6 of the Technical Report.

I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101.

I have prior involvement with the property that is subject of the Technical Report since I am the former Manager Environment and now a Director of the Swedish Company.




Original Document signed by Henning Holmström, M.Sc., Ph.D, MAusIMM, MAIG

Henning Holmström, M.Sc., Ph.D, MAusIMM, MAIG

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