kevitsa pre-feasibility study · 7/1/2006 · pit mining operation at sml’s kevitsa property in...

Prepared for:

Scandinavian Minerals Limited

Report E705

By

S•t B•a•r•b•a•r•a Consultancy Services 9 John Street London

WC1N 2ES United Kingdom

(In Partnership with CSA International)

KEVITSA

PRE-FEASIBILITY

STUDY

July 2006

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study ii

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

1 Introduction ............................................................................................ 10

1.1 Access .................................................................................................. 10

1.2 Physiography......................................................................................... 11

1.3 Climate................................................................................................. 11

1.4 Mineral rights ........................................................................................ 11

2 Project History ......................................................................................... 13

2.1 Timeline................................................................................................ 13

2.2 Historical Work and Status....................................................................... 13

3 Geology .................................................................................................. 17

3.1 Summary.............................................................................................. 17

3.2 Conclusions ........................................................................................... 18

4 Mining .................................................................................................... 19

4.1 Mine Planning – Whittle Optimisation ........................................................ 19

4.2 Mining Investment and Operating Cost Estimate ......................................... 25

5 Ore Types, “Sulphur Sufficiency” and Representativity of Metallurgical Test Samples ........................................................................................................... 28

5.1 Ore Types ............................................................................................. 28

5.2 Sample Representativity and Prediction of Performance ............................... 28

5.3 Sulphur Sufficiency................................................................................. 29

6 Metallurgy – Test Work and Design ............................................................. 35

6.1 Overview .............................................................................................. 35

6.2 Testwork............................................................................................... 35

6.2.1 Summary of Historical Testwork ......................................................... 35

6.2.2 Bench Scale Test Work Programme: Production of Separate Concentrates 39

6.2.3 Mini Pilot Test Work Programme for Production of Separate Concentrates. 41

6.2.4 Bench Scale Testwork on Low and High Grade Samples ......................... 42

6.2.5 Flotation Tests on Surface Samples..................................................... 43

6.2.6 Estimated Quality and Recoveries of Cu and Ni Concentrates produced in Full Scale Production at Kevitsa ...................................................................... 46

6.3 Plant Design .......................................................................................... 46

6.3.1 Process design................................................................................. 46

6.3.2 Energy Consumption and Cost ........................................................... 47

6.3.3 Flotation Chemicals .......................................................................... 48

6.3.4 Grinding Media and Mill Liners............................................................ 50

6.3.5 Maintenance Supplies ....................................................................... 50

6.3.6 Process Investment Costs.................................................................. 50

7 Tailings and Waste Rock Disposal, Water Balance.......................................... 51

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study iii

7.1 Design Basis for Tailings Ponds ................................................................ 51

7.2 Construction of Tailings Ponds.................................................................. 51

7.3 Decant and Return Water System............................................................. 52

7.4 Cost of Tailings Dam Construction ............................................................ 52

7.5 Waste Rock Piles .................................................................................... 53

7.6 Water Balance ....................................................................................... 54

7.7 Recommendations for Further Work .......................................................... 55

8 Infrastructure .......................................................................................... 56

8.1 Summary.............................................................................................. 56

8.2 Access Road and Bridge .......................................................................... 56

8.3 Potable Water Supply and Sewage Treatment............................................. 56

8.4 Power Supply ........................................................................................ 57

8.5 Maintenance Facilities ............................................................................. 57

8.6 Other Facilities....................................................................................... 57

9 Environmental ......................................................................................... 58

9.1 Key Environmental Issues ....................................................................... 58

9.1.1 General .......................................................................................... 58

9.1.2 Water Effluents Management ............................................................. 58

9.1.3 Closure Design................................................................................. 58

9.1.4 Eagle Nesting .................................................................................. 58

9.1.5 Dust and Noise Emissions.................................................................. 58

9.2 Introduction .......................................................................................... 58

9.3 Regulations and Requirements ................................................................. 59

9.3.1 Finnish Legislation............................................................................ 59

9.3.2 European Union Requirements ........................................................... 60

9.3.3 World Bank Requirements ................................................................. 60

9.4 Current Environmental Conditions............................................................. 60

9.4.1 Local Setting ................................................................................... 60

9.4.2 Population Structure, Livelihoods and Services ..................................... 60

9.4.3 Land Use and Social Structure, Plans Concerning the Area ..................... 61

9.4.4 Landscape and Topography ............................................................... 61

9.4.5 Soil ................................................................................................ 61

9.4.6 Groundwater ................................................................................... 63

9.4.7 Surface waters................................................................................. 63

9.4.8 Climate and Air Quality ..................................................................... 64

9.4.9 Biocoenoses .................................................................................... 64

9.4.10 Conservation Areas .......................................................................... 65

9.4.11 Roads and Traffic ............................................................................. 67

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study iv

9.4.12 Archaeology and Cultural History........................................................ 68

9.4.13 Diverse Use of Nature ....................................................................... 68

9.5 EIA Process ........................................................................................... 68

9.6 Baseline Study and EIA Methodologies ...................................................... 69

9.6.1 Social and Economic Circumstances .................................................... 69

9.6.2 Impacts on General Health and Wellbeing ............................................ 69

9.6.3 Roads and Traffic Connections............................................................ 70

9.6.4 Surface Waters ................................................................................ 70

9.6.5 Groundwater ................................................................................... 71

9.6.6 Soils............................................................................................... 72

9.6.7 Geochemistry .................................................................................. 72

9.6.8 Fauna and Flora ............................................................................... 72

9.6.9 Fish and Fishing ............................................................................... 73

9.6.10 Conservation Areas and Significance of Biological Resources................... 73

9.6.11 Landscape....................................................................................... 73

9.6.12 Air Quality....................................................................................... 74

9.6.13 Noise and Vibration .......................................................................... 74

9.6.14 Archaeology and Cultural Heritage ...................................................... 74

9.6.15 Assessment of Risks to the Environment.............................................. 75

9.7 Proposed Environmental Programme Continuation ...................................... 75

9.7.1 Reindeer Herders Cooperation............................................................ 75

9.7.2 EIA Process ..................................................................................... 75

9.7.3 Environmental Baseline Study............................................................ 76

9.7.4 EIA Report ...................................................................................... 76

9.7.5 Environmental Permit and Water Permit .............................................. 76

9.8 Estimate of Environmental Costs .............................................................. 77

9.9 Closure and Rehabilitation ....................................................................... 77

9.9.1 Closure Planning and After Treatment ................................................. 77

9.9.2 Closure and After-Care Measures........................................................ 78

9.10 Environmental Cost Estimates............................................................... 81

9.10.1 Closure Cost.................................................................................... 81

9.10.2 Project Implementation Environmental Costs ....................................... 81

10 Transportation of Concentrates................................................................... 83

11 Marketing................................................................................................ 85

11.1 Introduction ....................................................................................... 85

11.2 Nickel ................................................................................................ 85

11.3 Copper .............................................................................................. 85

11.4 Cobalt ............................................................................................... 86

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study v

11.5 Platinum Group Metals ......................................................................... 86

11.6 Summary........................................................................................... 86

12 Capital Costs ........................................................................................... 88

12.1 General ............................................................................................. 88

12.2 Full Feasibility Study............................................................................ 89

12.3 Land Acquisition.................................................................................. 89

12.4 Closure Costs ..................................................................................... 89

13 Operating Summary and Manpower Costs .................................................... 90

13.1 Summary........................................................................................... 90

13.2 Manpower .......................................................................................... 90

14 Financial Analysis ..................................................................................... 93

15 Project Implementation and Schedule.......................................................... 95

16 Conclusions, Risks and Opportunities........................................................... 98

17 References .............................................................................................. 99

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study vi

List of Figures

Figure 1-1 Location of Kevitsa in Relation to the Port and Smelter ............................. 11

Figure 2-1 Scandinavian Gold’s claim blocks at Kevitsa ............................................ 15

Figure 4-1 Final Pit design (100m grid interval) ...................................................... 20

Figure 4-2 Bottom 3D View for Pit Shell in final year 16. .......................................... 21

Figure 4-3 Production profile................................................................................ 23

Figure 4-4 Payable Nickel and Copper Outputs and Feed Grades. .............................. 23

Figure 5-1 Estimated Fraction of Sulphide Nickel vs. Sulphur Index ........................... 31

Figure 5-2 Estimated Non S Nickel vs. S Index ....................................................... 32

Figure 5-3 Fraction of Sulphide Nickel vs. S Index................................................... 32

Figure 5-4 Distribution of Sulphur Deficient Blocks (above 0.19% Ni)......................... 34

Figure 6-1 Gold recovery against Copper Recoveries ............................................... 40

Figure 6-2 Pt and Co Recoveries against Ni Recoveries ............................................ 41

Figure 6-3 Electrode Potential .............................................................................. 45

Figure 9-1 Location of the deposit (circled in brown). Map taken from the Sodankylä Municipal Internet Service ................................................................................... 62

Figure 9-2 Typical peat land area at Kevitsa Project. Landscape west of Kevitsa hill (centre). (Photo LVT).......................................................................................... 65

Figure 9-3 Koitelainen conservation area borders by Kevitsa (black line – green hatch) 67

Figure 9-4 Baseline water quality sampling sites and Kitinen River discharge volume measurement sites ............................................................................................. 71

Figure 10-1 Map of Finland Showing Locations of Mine and Port ................................ 84

List of Tables

Table 2-1 Mineral reserve and mineral resources statement by SRK (2000) ................ 16

Table 3-1 Measured and Indicated Resources1 ........................................................ 17

Table 3-2 Inferred Resources ............................................................................... 17

Table 4-1: Metal Prices ....................................................................................... 19

Table 4-2: Whittle vs. Practical Design Results ....................................................... 21

Table 4-3: Kevitsa Mineral Reserve - April 2006...................................................... 22

Table 4-4: Kevitsa Life of Mine Production Schedule ................................................ 24

Table 4-5: Mine Capital Cost Schedule ($000’s unless otherwise stated)..................... 26

Table 4-6: Mine Operating Cost Schedule – Excluding Staff Supervision ($000’s unless otherwise stated) ............................................................................................... 27

Table 5-1 Composition (%) of main sulphide minerals from Kevitsa ........................... 29

Table 5-2 Summary of Important Metallurgical Test Samples.................................... 30

Table 5-3 S Index of Kevitsa Blocks > 0.19% Nickel................................................ 33

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study vii

Table 6-1 Average Grades of Feed, Bulk Concentrate and Recoveries in Bulk Flotation (SRK Consulting, December 2003)........................................................................ 35

Table 6-2 Bench Scale Test Results....................................................................... 35

Table 6-3 Mini Pilot Test Run Results..................................................................... 35

Table 6-4 Average Grades of Ore Feed, Bulk Concentrate and Recoveries in Flotation (SRK Consulting)................................................................................................ 38

Table 6-5 Pipe Zone Feed Assays.......................................................................... 38

Table 6-6 Bulk Rougher Flotation.......................................................................... 38

Table 6-7 Bulk Cleaner Flotation........................................................................... 39

Table 6-8 Feed Grades of Mini Pilot Run................................................................. 39

Table 6-9 Average result of 13 bulk flotation tests................................................... 39

Table 6-10 Flotation Feed Analyses ....................................................................... 39

Table 6-11 Bench Scale Sequential Flotation, Analyses of Final Cu and Ni concentrates of Reference Test 18 .............................................................................................. 41

Table 6-12 Flotation feed analyses........................................................................ 42

Table 6-13 Mini Pilot Sequential Flotation Run - Copper and Nickel Final Concentrates.. 42

Table 6-14 Description of Feed Samples ................................................................ 43

Table 6-15 Analyses of Feed Samples.................................................................... 43

Table 6-16 Bench Scale Sequential Flotation Tests on Three Samples of Various Grades, Analyses of Final Cu and Ni Concentrates............................................................... 43

Table 6-17 Bench Scale Flotation on Near Surface Samples Compared to Reference Test 18.................................................................................................................... 44

Table 6-18 Lime requirement for raising pH to 11 ................................................... 45

Table 6-19 Total Energy Costs.............................................................................. 47

Table 6-20 Flotation Chemical Costs...................................................................... 48

Table 7-1 Pre Feasibility Design Basis for Tailings Disposal ....................................... 51

Table 7-2 Unit Costs of Construction of Tailings Pond............................................... 52

Table 7-3 Summary of Construction of Tailings and Recycle Water Ponds ................... 53

Table 7-4 Pre Feasibility Design Basis for Mine Waste Disposal.................................. 54

Table 7-5 Water Balance Summary ....................................................................... 54

Table 8-1 Summary of Infrastructure Costs............................................................ 56

Table 9-1 Water systems in the vicinity of Kevitsa................................................... 63

Table 9-2 Weather statistics for Sodankylä. Air temperature, annual precipitation and snow cover. (Drebs et al 2002) ............................................................................ 64

Table 9-3 Cost Estimate of Mine Closure ................................................................ 81

Table 10-1 Unit Costs of Transportation................................................................. 83

Table 10-2 Total Costs of Transportation by Destination........................................... 83

Table 12-1 Capital Cost Schedule (Million €)........................................................... 88

Table 12-2 Full Feasibility study costs ................................................................... 89

Table 12-3 Distribution of Different Land Types in the Original Claim Area of Kevitsa ... 89

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Table 13-1 Summary of Operating Costs ............................................................... 90

Table 13-2 Mining Labour Complement and Operating Costs..................................... 91

Table 13-3 Processing Labour Complement and Operating Costs ............................... 91

Table 13-4 General & Administration Labour Complement and Operating Costs ........... 92

Table 14-1 Financial Model Layout ........................................................................ 93

Table 14-2 Project Metal Price Scenario Sensitivity (€M) .......................................... 93

Table 14-3 Kevitsa NPV7% Project Sensitivity (€M)................................................... 94

Table 14-4 Kevitsa IRR Project Sensitivity (%) ....................................................... 94

Table 15-1 Project Schedule ................................................................................ 96

St Barbara Consultancy Services E705 – Kevitsa Pre-Feasibility Study ix

List of Appendices

3A – Kevitsa 43-101 Geological Report – Refer to SEDAR report dated 10th March 2006

4A – Whittle Optimisation Report

5A – Ore Types at Kevitsa

5B – Metallurgy Samples, May 2005

5C – Plan Showing Position of Metallurgy Samples, May 2005

5D – Comparison of Analytical Methods on the core from Kevitsa

6A – Process Flowsheet, Crushing and Grinding

6B – Process Flowsheet, Copper Flotation

6C – Process Flowsheet, Nickel Flotation

6D – Materials Balance

6E – Process Plant Layout

6F – Metal Recoveries

7A – General Site Layout, Including Tailings and Waste Dump

7B – Tailings Pond Cross Section

7C – Process Water Balance

8A – Road and Power Lines Layout

9A – ABA Testing

9B – Acid Drainage Testwork

11A – Specifications of Nickel and Copper Concentrates

11B – Nickel and Copper Marketing Report

11C – Cobalt Industry Review

11D – Platinum Group Metals Industry Review

12A – Investment Costs

13A – Organisational Structure

14A – Financial Model

15A – Qualified Persons’ Letters and Certificates

St Barbara Consultancy Services 1 E705 – Kevitsa Pre Feasibility Study

Executive Summary Introduction and Background

The production of this Independent Technical Report has been managed by St Barbara Consultancy Services (“St Barbara”) at the request of Scandinavian Minerals Ltd. (“SML”). The Report comprises an evaluation, at pre-feasibility level, of a potential open pit mining operation at SML’s Kevitsa property in Northern Finland. The Report has been prepared in accordance with the guidelines set out in Canadian National Instrument 43-101.

The information contained in this Report has either been prepared directly by St Barbara or its partner CSA International. Where others have collaborated then that work has been checked by St Barbara.

The following Finnish entities contributed:

Geological Survey of Finland (“GTK”) – process testwork, design, operating and capital cost estimation;

Lapin Vesitutkimus Oy (“LVT”), a Finnish firm of environmental consultants - environmental information.

The Kevitsa deposit is a large resource of nickel-copper-cobalt-PGE-gold mineralisation, outcropping on surface, which first attracted exploration interest 20 years ago. It is located some 142km north-northeast of Rovaniemi in N Finland in terrain with a plateau level of approximately 230 m with local hills rising to 350 m. The climate is arctic, but there is no permafrost in the area and vegetation consists of bog and birch forest.

Title and Resources

SML holds a total of 32 km2 of claims covering the Kevitsa deposit, including the areas necessary for mine development, processing plant and tailings and waste rock dumps. SML has also applied for a further 2 km2 of claims to cover any likely extension of the waste rock are and granting of these claims is awaited.

The following resource estimates filed according to NI43-101 on 10th March 2006 have been prepared by Mr Dexter Ferreira BSc (Geology), BEng (Mining), Pri.Sci.Nat.

Measured and Indicated Resources1

% g/t Contained

Metal (000 t)

Ni Cut-off %

Mt

Ni Cu Co Au Pd Pt Ni Cu

Measured

0.1 185 0.22 0.29 0.01 0.09 0.13 0.21 408 537

0.2 90 0.31 0.42 0.01 0.13 0.19 0.30 279 378

Indicated

0.1 102 0.22 0.30 0.01 0.08 0.11 0.18 225 307

0.2 51 0.30 0.44 0.01 0.11 0.16 0.25 152 224

Measuredand

Indicated

0.1 287 0.22 0.29 0.01 0.09 0.13 0.20 632 834

0.2 141 0.30 0.42 0.01 0.12 0.18 0.28 422 591

St Barbara Consultancy Services 2 E705 – Kevitsa Pre-Feasibility Study

Inferred Resources

Ni Cut-off %

% g/t Contained

metal(000 t)

Inferred

Mt

Ni Cu Co Au Pd Pt Ni Cu

0.1 544 0.22 0.32 0.01 0.07 0.08 0.09 1,197 1,741

0.2 291 0.29 0.46 0.01 0.09 0.09 0.12 843 1,337

1 Contained metal calculations may not reconcile exactly due to rounding.

Insofar as the present state of knowledge on the Kevitsa deposit stands, there are no known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues which may materially affect the potential development of the resources quoted herein.

Mining and Reserves

A comprehensive Whittle optimisation and subsequent design of the Kevitsa open pit was carried out. As no mining activity has occurred at Kevitsa before, quotations from Finnish mining contractors were to be used to estimate the mining cost. In the event, contractors did not provide an estimate in time for the Whittle optimisation and the rate of €1.27($1.50)/t material moved was used as a preliminary estimate. To adjust the mining cost by depth Mining Cost Adjustment Factors (MCAF) for a similar project were used.

The Whittle optimisation was run on a wall slope set of 45° for the first 30m below surface. The wall angle was then increased to 55° for the remainder of the pit.

In order to assess the mineable reserves and to provide an annual life of mine schedule a mining dilution of 3% at zero grade and a mining recovery fraction of 0.97 were applied.

The specified case design generated from the optimal pit shell is roughly oval. The pit’s longest axis which runs roughly Northwest Southeast has a length of 900m. On its Northeast Southwest axis the design produced a length of 700m. At its deepest point the pit is 330m below surface. At its steepest point the pit wall has an angle of 53.8°.


Final Pit design (100m grid interval)

Mine waste stockpiles are situated to the North and Northwest of the pit.

To maximise the NPV of the project within the constraint of a 4.5Mtpa operation a cut-off optimisation was performed. At a cut-off of around 0.18% Ni the NPV peaked and a feed grade of 0.295% Ni was achieved. This resulted in the following Proven and Probable Reserves.

Kevitsa Mineral Reserve - April 2006

Mineral Reserve

(%) (g/t) Ni Cut-off (%)

Tonnes

Ni Cu Co S Au Pd Pt

Proven 0.18% 56.2 Mt 0.295 0.415 0.014 1.683 0.141 0.201 0.310

Probable 0.18% 10.6 Mt 0.295 0.492 0.015 1.896 0.142 0.171 0.267

Total 0.18% 66.8 Mt 0.295 0.427 0.014 1.717 0.141 0.196 0.303

According to normal practice, Measured Mineral Resources are converted to Proven Reserves and Indicated Resources to Probable Reserves using the ‘modifying factors’ above. Inferred resources were assumed to have zero value for the pit optimisation.

At a throughput of 4.5Mtpa the reserve will be mined over 15 years with one additional year of pre-stripping. The designed open pit contains 223 million rock tonnes of which 66.8 million are above cut off, resulting in an average stripping ratio of 2.34t waste: 1t ore.


The mineralisation outcrops on surface and an oxidised zone of a few meters above the water table has been noted with poorer and less well understood flotation behaviour. This oxidised material, expected to represent less than 2% of the total ore mined will be set aside in the first year of operation. It will then be blended in with the unoxidised ore, enabling the flotation redox (reducing/oxidation) conditions to be adjusted according to normal industry practice.

Processing

In 2005 a successful bench scale test series for producing separate Cu and Ni concentrates was carried out.

Bench Scale Test Results

Product Wt Cu Ni Pd Pt Au

Test 18 % % R-%

% R-%

g/t R-%

g/t R-%

g/t R-%

CuConcentrate

0.8 26.4 45.9 0.5 1.5 3.1 11.8 4.8 14.5 4.2 19.4

NiConcentrate

1.8 6.5 26.1 10.0 65.1 4.0 35.3 5.1 35.9 1.0 10.3

Feed 100 0.4 100 0.3 100 0.2 100 0.2 100 0.3 100

Also, in 2005 a successful mini pilot test run for producing separate Cu and Ni concentrates was conducted.

Mini Pilot Test Run Results


% % R-%

% R-%

g/t R-%

g/t R-%

g/t R-%

Cu, Test 8

1.6 24.5 74.7 0.6 2.7 1.9 16.7 2.6 19.6 2.3 23.3

Ni, Test 8

1.9 2.9 10.7 12.0 67.8 2.3 24.6 1.9 16.8 0.6 7.5

Feed 100 0.5 100 0.3 100 0.2 100 0.2 100 0.2 100

The process design resulting from the above testwork consists of conventional: crushing, SAG and ball milling, sequential copper and nickel flotation with thickening and filtration of the copper and zinc concentrates. It is planned to reject the high sulphide nickel cleaner tailing to isolate this acid generating tailings for disposal in a specially designed facility adjacent to the main non acid generating tailings dam.

Based on a 4.5Mtpa throughput, the following annual concentrate production is expected

74,000t nickel concentrate at 12% Ni

47,500t copper concentrate at 24.5% Cu


Ore Mineralogy

The orebody can be considered to consist of two predominant ore types:

“Main”

“Nickel-PGM”

The main difference between them is that the nickel-PGE type has a higher Ni/Co ratio than the Main type, and has a tendency towards lower sulphur and copper contents. The main ore type makes up 93 % of all ore blocks with more than 0.2 % nickel to a depth of 330 m. Mineralogically the two ore types have the same gangue and sulphide minerals and therefore each responds in a similar way to the flotation process.

With lowering sulphur content the pyrrhotite content diminishes and millerite and heazlewoodite start forming. These minerals are considered to float in a similar manner to pentlandite. However, as sulphur becomes deficient, some of the nickel can occur in the olivine and not be available for recovery by flotation. A Sulphur Index (S/(Ni+Cu)) has been used to calculate such deficiency and it is considered that the metallurgical samples tested are representative of the ore blocks planned to be mined

It could be argued that the block modelling methodology may have “smoothed” the calculation of the Sulphur Index and therefore underestimated the proportion of sulphur deficient zones. Until this argument is refuted and some questionable non sulphide nickel (citrate method) assays repeated, it is recommended that a factor of 0.95 is tested as a sensitivity to the base case average nickel recovery of 67.1%. On the base care financial model this reduces the IRR of the project by 2.2%.

Tailings Disposal

Tailings disposal at Kevitsa will be in separate low and high sulphur tailings ponds of downstream earthfill construction, with the latter having a bentonite cover to the embankment.

Construction material for the 7 m pervious starter dyke will be natural till located nearby. For raising the embankment, mainly natural till and if suitable also coarse tailings will be used. The stability of embankments higher than 7 m will be secured by using low sulphur waste rock for construction of the outer face of the impoundment.

Tailings will be deposited to a total thickness of 22 m and the maximum dam height will be 26 m.

Infrastructure

The overall site layout calls for an open pit, process plant and workshops, stores, canteen, waste rock piles and tailings and return water dams. A new access road, power line and bridge over the Kitinen River are the most important items.

Environment

Collection of data for an Environmental Impact Assessment is ongoing and this is expected to be finalised in the third quarter of 2006.

Recent permitting practise in Finland, discussions with authorities and interaction group meetings in nearby Sodankylä have resulted in the following design principals for planning the Project:

Water Effluents Management - Towards the east of the site lies the Koitelainen conservation area and the nearest water resource is Satojärvi Lake which is considered to be valuable as a migration resting place for birds. Another arm of the Koitelainen conservation area reaches towards the mine open pit area from the north - east.

Any potential changes to the Koitelainen natural environment could cause limitations being put on the mining operations. Therefore the project design philosophy is that all site water will drain to the west, away from the Natura 2000 site. The Kitinen River is


sufficiently large to dilute treated mine effluents to an insignificant level in terms of water quality change.

Closure Design - It is likely that operations after-care responsibility will, in the environmental permit be stipulated to be for at least 30 years after mine closure and the overall mine design will take this into account.

Eagle Nesting - There is a known nest of eagles in the vicinity and all necessary measures will be taken to avoid any harm to the species which could inter alia include building reserve nest platforms.

Dust and Noise Emissions - The Koitelainen Natura 2000 area must be protected from the potential impacts of mining operations and the design and placing of buildings, haul roads and access roads will take cognisance of this.

Market and Price Forecasts

A market study on the sale of nickel and copper concentrates and the revenues to be expected was undertaken based on concentrate analyses from mini-pilot trials undertaken in October 2005.

Ni Concentrate: 12.0 % Ni, 2.9 % Cu, 0.5% Co, 2.3 g/t Pd, 1.9 g/t Pt and 0.6 g/t Au

Cu Concentrate: 0.6 % Ni, 24.5 % Cu, 1.9 g/t Pd, 2.6 g/t Pt and 2.3 g/t Au

The following transportation options were costed:

Kevitsa – truck – Rovaniemi – train – Harjavalta smelter

Kevitsa – truck – Kemi harbour – ship – Pori harbour – train – Harjavalta smelter

Kevitsa – truck – Kemi harbour – ship – Rotterdam harbour

The sale of copper and nickel concentrates from the proposed Kevitsa mine is assessed as being straightforward. Interest has been shown by a number of potential consumers and good estimates for anticipated concentrate revenues have been able to be made. The prospects for reasonable prices for by-product cobalt and PGM credits have also been shown to be strong.

Global market reviews were also undertaken by St Barbara on cobalt and Platinum Group Metals to provide medium term price forecasts.

Nickel and copper prices were provided by a leading base metals consulting firm.

Copper ($/t) $1.21/lb ($2,668/t) Nickel ($/t) $4.68/lb ($10,319/t) Cobalt $15.00/lb Platinum ($/oz) $750/oz Palladium ($/oz) $200/oz Gold ($/oz) $450/oz Silver ($/oz) $8.00/oz


Capital Costs

The capital cost schedule below incorporates the following principles:

Two years initial full feasibility study with on-going test-work in the third year;

Two years construction and commissioning period for process plant;

Mining handled by contractor;

Indirect costs include capital spares and first fill;

EPCM costs set at 5 % of total investment costs;

Value-added tax (22 %) will be returned to the company, and it has therefore not been included in the capital cost list;

A contingency of 20 % has been applied for initial capital up to year 1 only.

Capital Cost Schedule (Million €)

Period Total -3 -2 -1 1 4 7 11

Full Feasibility Study

2.00 1.00 1.00

LandAcquisition

0.35 0.35

Mining Equipment

0.45 0.10 0.35

Process Plant 42.62 18.95 23.67

Tailings Facility

23.69 5.14 5.83 3.91 4.69 4.12

Infrastructure 6.80 6.54 0.26

Indirects 4.50 4.50

EPCM 4.00 2.00 2.00

Contingency 14.34 0.20 0.27 6.55 7.32

Total Capex 98.75 1.20 1.62 39.28 43.93 3.91 4.69 4.12

Total initial capital expenditure is €86.03M.

A closure cost of €6M is allocated to the final year of production.


Operating Costs

The majority of operating costs summarised below were based on budget quotations of materials from two or more companies.

Summary of Operating Costs

Description €/t mined or processed

Mining (contractor) 1.1-2.6

(company) 0.11

Processing (manpower) 0.60

(energy) 1.57

(flotation chemicals) 2.01

(grinding media, liners) 1.02

(maintenance supplies) 0.54

External services 0.10

General & Administration 0.32

TOTAL 7.37-8.87

Financial

The project Net Present Value (NPV) is €91.3 M at a 7% discount rate and the Internal Rate of Return (IRR) is 17.5%. The average operating cost of nickel nett of by product copper credits and payable elements in the nickel concentrate is €1.06/lb. A 2% increase or decrease in discount rate changes the NPV to €62.8M and €129.3M respectively.

The key features of the model are as follows:

100% equity

Cash flow before interest, tax, depreciation and amortisation

2006 money terms (i.e. no adjustments for inflation)

Exchange rate: € = $1.212 (set in February 2006 when the forecast dollar metal prices were provided)

The project economics are seen to be most sensitive to total revenue/price, $/€ exchange rate and operating cost.

Price Scenario NPV7% IRR (%)

Ni $10,319/t Cu $2,668/t 91.3 17.5

Ni $12,500/t Cu $3,300/t 219.5 30.4

Ni $14,000/t Cu $4,000/t 328.2 40.5

Conclusion

The Kevitsa Project at the current prefeasibility stage shows an IRR of 17.5% and an NPV7% of €91.3M using appropriately conservative medium term metal prices and smelter terms based on those expected to be reached following discussions with smelters.


Although the metallurgical performance of the project is based on pilot plant testwork the performance of the differential flotation process is not considered to be fully optimised, particularly with regard to minimisation of reagent use and control of pulp Redox potential at the head of the copper circuit. This represents both risk and the opportunity to improve flotation performance.


1 Introduction The production of this Independent Technical Report has been managed by St Barbara Consultancy Services (“St Barbara”) at the request of Scandinavian Minerals Ltd. (“SML”), formerly Scandinavian Gold Limited prior to January 2006. The Report comprises an evaluation, at pre-feasibility level, of a potential open pit mining operation at SML’s Kevitsa property in Northern Finland. The Report has been prepared in accordance with the guidelines set out in Canadian National Instrument 43-101.

The information contained in this Report has either been prepared directly by St Barbara or its partner CSA International or where others have collaborated then that work has been checked by St Barbara. The following entities contributed:

Geological Survey of Finland (“GTK”) – process testwork, design, operating and capital cost estimation;

Lapin Vesitutkimus Oy (“LVT”), a Finnish firm of environmental consultants - environmental information.

The Kevitsa deposit is located some 142km north-northeast of Rovaniemi, the capital of Finnish Lapland. It comprises a large body of nickel, copper, cobalt, gold, platinum and palladium mineralisation which outcrops at surface and is potentially amenable to open pit mining.

The property was found and subsequently drilled by GTK. A number of studies were undertaken on the deposit in the mid-1990s which at that time concluded that the sulphides could not be economically floated to produce a concentrate suitable for conventional smelting.

In June, 2005 SML announced that bench-scale laboratory tests at the Mineral Technology Laboratory of GTK (“GTK Mintech”) had succeeded in producing separate smelter-grade nickel and copper concentrates at acceptable levels of recovery. SML followed up this work with mini-pilot tests, which confirmed the bench test results.

Following these results, SML decided to embark on a new pre-feasibility study, based on an open pit mining operation at Kevitsa, with production of separate nickel and copper concentrates for sale to smelters

1.1 Access It is possible to travel by road, railway or plane to Rovaniemi from Helsinki, and then it is necessary to travel by road along E75 to the village of Petkula. From there an 8 km long forest road travelling east leads to the deposit. The forest road starts at the crossing of the Kitinen River at the hydro power plant at Vajukoski. The forest roads are owned by the state forest company Metsähallitus. Metsähallitus also owns most of the land at Kevitsa with the exception of a private area on the edge of the claim in square 3C.


Figure 1-1 Location of Kevitsa in Relation to the Port and Smelter

1.2 Physiography The terrain is a plateau at level 220 m to 240 m with local hills rising to 350 metres. The Kevitsa deposit is sited at the water shed between the stream Mataraoja draining towards northwest and west and Viivajoki draining towards east and southeast. The flat terrain creates extensive areas of bog land alternating with slightly raised terrain with pine forest. The original forest at Kevitsa was cut down decades ago.

Bedrock outcrops on the hills but is generally covered by a 1 to 5 metre thin layer of clay and/or sandy till. In boggy land a 1 to 5 metre thick peat layer is developed on top of the till.

1.3 Climate The climate is arctic. Based on long term climatic data from Sodankylä (1971 to 2000) the average temperature was -0.8ºC and average precipitation was 507mm. October to April has negative average temperatures with January being the coldest with an average of -14.1ºC. Half of the precipitation falls during this period as snow. The summer months warm up fast with July being the warmest with an average of 14.3ºC. There is no permafrost in the area.

1.4 Mineral rights Minerals rights in Finland are owned by the state and regulated by an office under the Ministry for Trade and Industry. The Chief Inspector of Mines grants exploration permits according to the legislation. Applications are open for all legal entities within the European Union. Initially one is granted a reservation which is valid for one year which gives exclusive rights for mineral exploration but does not grant access to land.


A claim is considerably more expensive (circa €1450 per km2 per year), but gives the right to conduct activities on the land. Claims are initially given for a period of 5 years and provided activities can be demonstrated, it can be extended for an additional 3 years. During the claim period a mining lease can be applied for provided a potential deposit has been located. In order to obtain a mining lease an environmental impact assessment report shall be presented to and approved by the authorities

SML currently hold a total of 32 square kilometres of claims covering the Kevitsa deposit, including the areas necessary for mine development, processing plants and tailings and waste dumps. SML is also awaiting the granting of a claim for a further 2 square kilometres at the northeastern corner of the existing claims.


2 Project History

2.1 Timeline The Kevitsa deposit is a large resource of nickel-copper-cobalt-PGE-gold mineralisation which first attracted exploration interest 20 years ago. The relatively low grade and complex nature of the mineralogy represented technical challenges which have been addressed over the years:

1930s: The first reports on ultramafic rocks in the area

1969: Systematic geological mapping of the area

1973: first nickel mineralised float sample was found.

1984: Drilling conducted by the Geological Survey of Finland began

1987: Mineralisation was first discovered, when a diamond drill hole intersected 23 m with 0.36 % nickel and 0.46 % copper plus gold and PGE

1995: Surveys carried out by GTK outlined a large low grade nickel-copper resource. Total drilling by the Survey at the deposit and adjoining areas comprised 563 holes for a total of 48,474 m.

1995: The Finnish state put the deposit up for tender, which was awarded to Outokumpu Mining Oy.

1997: Outokumpu decided not to develop the deposit citing difficult metallurgy and low grade, after drilling 15 holes for a total of 2220 m and conducting extensive metallurgical testing. They returned it to the state

2000: Scandinavian Gold Prospecting AB, a subsidiary of Scandinavian Minerals Ltd., obtained a claim reservation covering the deposit. In November, Scandinavian Gold Prospecting AB was granted claims for an area of 8.68 km2.The claims expire on November 10, 2008.

2000: Initially SML (then called Scandinavian Gold Ltd.) pursued a route with production of a bulk concentrate for treatment by a pressure acid leach process known as PlatSolTM. SRK Consulting (UK) Ltd did a pre-feasibility study based on this approach. They identified an open pit, probable reserve of 120 million tons with 0.21 % nickel. 0.30 % copper, 0.012 % cobalt and 0.52 gram/tons PGE plus gold to a depth of 450 m and a waste: ore ratio of 1.8:1. Based on a mining rate of 15 million tons a year the project could under certain assumptions represent a viable project, and SRK recommended continued work on the project.

2003 - 2005: Scandinavian Gold focused on metallurgical test work. Additional drilling for ore delineation and collection of material for testing comprises 25 holes for a total of 4116 m.

2005: GTK Mineral Processing succeeded in making smelter-grade copper and nickel concentrates at reasonable recoveries. The process has been demonstrated at bench scale and in a mini pilot plant.

2006: SML is currently pursuing development of the project based on production of separate copper and nickel concentrates.

2.2 Historical Work and Status The first report on ultramafic rocks at Kevitsa goes back to the 1930s, but systematic mapping of the area was first started by the Geological Survey of Finland (“GTK”) in 1969. In 1973 Tapani Mutanen found the first nickel-bearing float sample with disseminated sulphides. Initial diamond drilling in 1984 by GTK was directed towards geophysical anomalies at the basal contact of the intrusion. Only low grade pyrrhotite-bearing massive sulphides were located.

Systematic ground geophysical surveys (magnetic, gravity and electromagnetic), and till geochemical sampling were conducted by GTK from 1984 to 1987. Anomalies were


investigated by diamond drilling in 1987. Initial findings were nickel-poor massive sulphides but eventually holes no. 23 and 24 intersected 22-24m with 0.36 % nickel, 0.46 % copper and 1.4 gram/tons PGE and gold (Mutanen 1997).

Drilling was continued by GTK in 1990 and 1992-1995, and the outline of a large low grade nickel-copper occurrence was determined. The total amount of drilling conducted by GTK in the nickel-copper occurrence and adjoining areas comprise 563 holes with a total length of 48,474m. Of these, 278 holes totalling 32,845m outline the deposit.

The state auctioned the deposit in 1995, and the project was taken over by Outokumpu. Outokumpu drilled 15 holes for a total length of 2,220m, partly for collection of material for metallurgical testing. Following comprehensive metallurgical testing, they failed to make concentrates of nickel and copper at recoveries, which warranted development of the project, and they returned the project to the Finnish Ministry of Trade and Industry in 1998.

In January 2000, Scandinavian Gold Prospecting AB (“Scandinavian Gold”), a wholly-owned subsidiary of Scandinavian Minerals Ltd, acquired a reservation over the area, and on November 10, 2000 Scandinavian Gold was granted claims for the area. The claims expired after 5 years in November 2005, but extension has been granted for a further 3 year period (Figure 2-1). However, it is Scandinavian Gold’s intention to apply for a mining concession in 2006.

In 2000, Scandinavian Gold engaged SRK Consulting to compile all data and evaluate whether a large scale open pit operation with production of a bulk concentrate and hydrometallurgical treatment by the concentrate using the PlatSolTM process was a viable route for development of the project. . In addition to recovering nickel, cobalt and copper, the process also recovers PGE and gold.

SRK identified resources to a depth of 500 m and reserves for an open pit mining scenario to a depth of 450m.

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t K

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Table 2-1 Mineral reserve and mineral resources statement by SRK (2000)

Cash flow modelling implied that a viable project could exist and SRK recommended continued work. However, the SRK scenario had a short mine life (8 year mine life at 15 million tons a year) and large capital requirement (approx. US$500 million), Therefore Scandinavian Gold has looked into alternative ways for advancing the project.

Scandinavian Gold focused on modelling of the deposit, and it was possible to discriminate various mineralisation types. In 2003 SML drilled 14 holes, totalling 1,557 m, to test the potential for a subordinate nickel-PGE type of mineralisation which could be treated to make high grade nickel-PGE concentrate directly. Subsequent drilling in late 2004 - early 2005 showed that the tonnage potential for this type of mineralisation was too small to support mining in its own right.

In parallel, focus was on metallurgical work on the main nickel-copper-PGE type of mineralisation. Partly for production of bulk concentrates for hydrometallurgical testing by variously newly developed techniques, and partly for production of separate nickel and copper concentrates. Material for metallurgical testing has been obtained by core drilling. Drill holes have been positioned to obtain material and add information on grade distribution. From 2003 to 2005, Scandinavian Gold drilled 25 holes totalling 4116 m.

Scandinavian Gold initiated modelling of the intrusive complex based on various geophysical data. Modelling implies that the ultramafic segment of the Kevitsa intrusion forms a funnel shaped body with basis 1000 to 1200 m below surface. Due to the geometry of the mineralised zone, deep penetrating EM and magneto telluric data can only “see” to a depth of c. 500 m.

Collection of data for an environmental impact assessment has also been started, and this is expected to be finalised in the third quarter of 2006.

During the summer of 2005, metallurgical processing demonstrated that separate, smelter-grade copper and nickel concentrates can be produced at reasonable recoveries. Based on these findings SML is currently investigating the viability of a medium sized (up to 4-5 Mtpa) open pit operation producing separate copper and nickel concentrates for sale to smelters.

SRK Mineral Reserve Statement

Tonnage

Ni* Cu Co* Au Pt Pd S

% % % g/t g/t g/t %

Probable 120 0.21 0.30 0.012 0.11 0.25 0.16 1.40

SRK Mineral Resource Statement

Tonnage

(million Ni* Cu Co* Au Pt Pd S

tonnes) % % % g/t g/t g/t %

Indicated 150 0.18 0.27 0.011 0.09 0.23 0.15 1.23

Inferred 315 0.18 0.29 0.012 0.08 0.20 0.12 1.26

* Ni and Co grades are denoted in terms of metal available as sulphide mineralization

Available Metal Grades

(million

tonnes)

Available Metal Grades


3 Geology The re-estimation of the mineral resource was carried out in 2005/6 by CSAI/St Barbara and this section comprises the summary and conclusions of the 43-101 report filed on 10th March 2006 (included as Appendix 3A).

3.1 Summary The following resource estimates have been prepared by Mr Dexter Ferreira BSc (Geology), BEng (Mining), Pri.Sci.Nat.

Table 3-1 Measured and Indicated Resources1

% g/t Contained metal

tonnesNi Cut-off %

Million tonnes

Ni Cu Co Au Pd Pt Nickel Copper

Measured

0.1 185 0.22 0.29 0.01 0.09 0.13 0.21 408,000 537,000

0.2 90 0.31 0.42 0.01 0.13 0.19 0.30 279,000 378,000

Indicated

0.1 102 0.22 0.30 0.01 0.08 0.11 0.18 225,000 307,000

0.2 51 0.30 0.44 0.01 0.11 0.16 0.25 152,000 224,000

Measuredand

Indicated

0.1 287 0.22 0.29 0.01 0.09 0.13 0.20 632,000 834,000

0.2 141 0.30 0.42 0.01 0.12 0.18 0.28 422,000 591,000

Table 3-2 Inferred Resources

Ni Cut-off %

% g/t Contained metal tonnes

Inferred

Mt

Ni Cu Co Au Pd Pt Nickel Copper

0.1 544 0.22 0.32 0.01 0.07 0.08 0.09 1,197,000 1,741,000

0.2 291 0.29 0.46 0.01 0.09 0.09 0.12 843,000 1,337,000

Note 1: Contained metal calculations may not reconcile exactly due to rounding.

Insofar as the present state of knowledge on the Kevitsa deposit stands there are no known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues which may materially affect the potential development of the resources quoted herein.

It is considered that the above estimate of mineral resources will not be materially affected by mining, metallurgical, infrastructure and other relevant factors. As part of the Pre-Feasibility Study currently underway, two aspects of the deposit have been identified that result in lower than the deposit average metallurgical performance, i.e.:

Surface oxidation in the first few metres of the deposit;

Sulphur deficient zones where some of the nickel is present in olivine and therefore not recoverable by flotation.


Sections 5 and 6 deal with the above topics and it is considered that the oxidation and sulphur deficiency effects only have the potential to deteriorate the Project economics rather than to make certain resource zones uneconomic (and therefore incapable of becoming reserves).

3.2 Conclusions Relevant geological and drillhole data from the operation were utilized to produce a mineral resource and reserve blockmodels. The following list of conclusions and recommendations have been arrived at as a result of this work.

There is a good understanding of the mineralisation behaviour of the Kevitsa deposit given the technical paper by Mr. P. Lamberg. This knowledge has been used to separate the composite data into relevant subpopulations representing separate magmatic events. The extension of this knowledge to develop variography is reasonable and agrees with the findings in his technical paper,

The drill hole database that supports the mineral resource and reserve estimates have been verified by the author using extensive checks and found to be representative of the mineralisation at Kevitsa,

The drilling density is sufficient for achieving a reasonable level of confidence in the resultant models,

Additional deep drilling will improve the confidence of estimates in the deeper areas of the deposit,

The estimation procedures used at Kevitsa, for the 3D block models, are fair and reasonable estimates of the resource and reserve. Experts have substantiated the introduction and application of Ni and Co numbers for the segregation of drillhole data into appropriate subsets,

The resources have been classified using reporting terminology and guidelines given in the CIM Standards, and using accepted practices worldwide. The resources were classified as measured for any block which was estimated using two-thirds of the first range as defined by the fitted double spherical variographic model (50m), approximately defining two thirds of the overall variance, indicated for any block estimated using the larger or entire first variographic range, and inferred for any block estimated using the second variographic range,

The cut-off strategy is reasonable and is similar to industry standard practices,

The methodology utilized for the estimation was ordinary kriging and the resultant blockmodel has proven to be particularly robust for all metals. Cross validation values indicate that there is a good level of confidence in the final blockmodel,

This study attempted to generate a blockmodel with metal values that are representative of the in situ resources of the Kevitsa deposit. Given the amount and quality of data available for resource estimation, it is believed that the resultant model is representative of the in situ resources at Kevitsa.


4 Mining A comprehensive report on the Whittle optimisation and subsequent design of the Kevitsa open pit is given in Appendix 4-1. This section provides a summary of the results of that optimisation and an operating and capital cost estimate for the mining operation to provide a check on budget contract mining cost information received.

4.1 Mine Planning – Whittle Optimisation The starting point for the Whittle optimisation was the detailed geological model resulting from the re-estimation of the resource (reported in Section 3). Waste blocks were added to the model and a density of 3.15 t/m3 applied to both ore and waste.

As no mining activity has occurred at Kevitsa before, quotations from Finnish mining contractors were to be used to estimate the mining cost. In the event, contractors did not provide estimates in time for the Whittle optimisation and the rate of €1.27($1.50)/t material moved was used as a preliminary estimate. To adjust the mining cost by depth Mining Cost Adjustment Factors (MCAF) for a similar project were used. Process recoveries were based on an analysis of metallurgical testwork results (Section 6 and Appendix ) while smelter terms were based on the results of the market study (Section 11 and Appendix 11B).

Metal prices used in the pit optimisation are listed in Table 4-1.

Table 4-1: Metal Prices

Element Price used in Whittle optimisation Units

Ni 10 319 $/t

Cu 2 668 $/t

Co 15 $/lb

Pt 500 $/oz

Pd 200 $/oz

Au 450 $/oz

These prices, and indeed operating costs are slightly different to those finally used in the economic evaluation of the project but they are not considered sufficiently different to materially alter the open pit design. In any case, the Whittle optimisation and pit redesign will be repeated in any Full Feasibility Study.

A discount rate of 7% was used in the Whittle optimisation.

The Whittle optimisation was run on a wall slope set of 45° for the first 30m below surface. The wall angle was then increased to 55° for the remainder of the pit

In order to assess the mineable reserves and to provide an annual life of mine schedule a mining dilution of 3% at zero grade and a mining recovery fraction of 0.97 were applied.

A maximum processing plant capacity was set to 4.5 million tonnes per annum with an administration and processing cost of $7.15 per tonne. Selling costs of $429.10/t of saleable nickel and $208.62/t of saleable copper were applied.

By applying Revenue Factors (more strictly speaking “price factors) of 0.3 to 1.5 Whittle generated 101 pit shells of increasing size and, up to a maximum, NPV.


Whittle optimises using three scenarios: best, worst and specified case:

Best case is defined as mining each successive incremental pit shell exactly to the shell before the next shell is to be mined;

Worst case is defined as mining to the ultimate pit shell from surface down to the deepest point in the shell in incremental benches with a complete bench cut being mined before proceeding to the next bench;

Specified case is defined by the user since the Best and Worst scenarios are both unrealistic in terms of practical mining. The user specifies the pushbacks or major shells and the required bench lead before a successive pushback may begin.

The specified case’s results are chosen since it would be the closest to a practical mining plan and hence real-world results. The other two scenarios then define the variability or risk in terms of mining poorly or within accepted standards and the potential gains to be had if better mining practises are employed.

The specified case design generated from the optimal pit shell is roughly oval. The pit’s longest axis which runs roughly Northwest Southeast has a length of 900m. On its Northeast Southwest axis the design produced a length of 700m. At its deepest point the pit is 330m below surface. A single-ramp system was incorporated into the design of the optimal pit. At its steepest point the pit wall has an angle of 53.8°. Figure 4-1 shows a plan of the pit layout.

Figure 4-1 Final Pit design (100m grid interval)

Mine waste stockpiles are situated to the North and Northwest of the pit and are shown in Appendix 7A.

To aid visualisation of the pit, a 3D sketch is given below


Figure 4-2 Bottom 3D View for Pit Shell in final year 16.

A comparison of the practically designed pit was made with the Whittle pit shell using Datamine software:

Table 4-2: Whittle vs. Practical Design Results

Optimised Ore selected for Processing

Base Case Total Mineral Reserve

Whittle Design Variance (%)

Total Tonnage(Mt)

223.1 223.9 0.36%

OreTonnage(Mt)

104.5 104.3 -0.21%

Payable Recovered Metal

Ni (t) 113,859 110,978

-2.53%

Cu (t) 259,054 253,509

-2.14%

Co (t) 2,595 2,545

-1.93%

Pt (g) 4,089,684 4,004,977

-2.07%

Pd (g) 2,480,565 2,432,052

-1.96%

Au (g) 2,387,794 2,365,595

-0.93%


Feed Grade (Diluted In-Situ)

Ni (%)

0.244% 0.242% -0.96%

Cu(%)

0.350% 0.346% -1.09%

Co(%)

0.013% 0.013% -0.42%

Pt(g/t) 0.246 0.244

-0.92%

Pd(g/t) 0.157 0.155

-1.02%

Au(g/t) 0.115 0.114

-0.65%

Metal Contents (Diluted In-Situ)

Ni (t) 254,979 251,996

-1.17%

Cu(t) 365,799 361,046

-1.30%

Co(t) 13,488 13,402

-0.64%

Pt(g) 25,720,229 25,429,981

-1.13%

Pd(g) 16,375,763 16,174,658

-1.23%

Au(g) 11,964,382 11,861,754

-0.86%

To maximise the NPV of the project within the constraint of a 4.5Mtpa operation, a cut-off optimisation was performed. A cut-off of 0.18% Ni was selected for the schedule and a feed grade of 0.295% Ni was achieved. This resulted in the following Proven and Probable Reserves.

Table 4-3: Kevitsa Mineral Reserve - April 2006

Mineral Reserve

(%) (g/t)Ni Cut-off (%) Tonnes Ni Cu Co S Au Pd Pt

Proven 0.18% 56.2 Mt 0.295 0.415 0.014 1.683 0.141 0.201 0.310

Probable 0.18% 10.6 Mt 0.295 0.492 0.015 1.896 0.142 0.171 0.267

Total 0.18% 66.8 Mt 0.295 0.427 0.014 1.717 0.141 0.196 0.303

According to normal practice, Measured Mineral Resources are converted to Proven Reserves and Indicated Resources to Probable Reserves using the ‘modifying factors’ above. Inferred resources were assumed to have zero value for the pit optimisation.


At a throughput of 4.5Mtpa the reserve will be mined over 16 years (one year of pre-stripping included). The designed open pit contains 222.7 million rock tonnes of which 66.8 million are above cut off, resulting in an average stripping ratio of 2.37t waste: 1t ore.

Figure 4-3 shows the life of mine production profile. Figure 4-4 shows expected Nickel and Copper payable outputs and Feed grades.

Production Profile

-

5,000,000

10,000,000

15,000,000

20,000,000

25,000,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

ag

e

-0.50

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

Ore Waste Strip Ratio

Figure 4-3 Production profile

Payable Ni & Cu Output

-

2,000

4,000

6,000

8,000

10,000

12,000

14,000

16,000

18,000

20,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

es

-0.100%

0.000%

0.100%

0.200%

0.300%

0.400%

0.500%

0.600%

CU NI NI% CU%

Figure 4-4 Payable Nickel and Copper Outputs and Feed Grades.

The full production schedule is provided in Table 4-4 overleaf.

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-4:

Kevit

sa L

ife o

f M

ine P

rod

ucti

on

Sch

ed

ule

Nic

ke

l F

ee

d G

rad

e C

uto

ff :

0

.18

%

Ore

(t)

W

aste

(t)

F

ee

d G

rad

e (

In-S

itu

Dil

ute

d)

%

g/t

1

1

0,0

00

,000

10

,00

0,0

00

N

i C

u

Co

P

t P

d

Au

2

9,4

13

,04

8

18

,54

5,1

40

4

,500

,00

0

14

,04

5,1

40

0

.285

%

0.3

42

%

0.0

13

%

0.3

49

0

.231

0

.135

3

52

,89

4,1

45

1

3,4

67

,628

4

,500

,00

0

8,9

67

,62

8

0.3

20

%

0.3

59

%

0.0

14

%

0.3

71

0

.268

0

.156

4

25

,54

3,9

82

2

0,8

84

,570

4

,500

,00

0

16

,38

4,5

70

0

.272

%

0.3

27

%

0.0

13

%

0.3

25

0

.231

0

.138

5

34

,16

2,1

06

2

0,9

16

,724

4

,500

,00

0

16

,41

6,7

24

0

.296

%

0.3

79

%

0.0

14

%

0.3

09

0

.201

0

.142

6

31

,55

7,0

27

2

0,9

42

,502

4

,500

,00

0

16

,44

2,5

02

0

.290

%

0.3

78

%

0.0

14

%

0.3

14

0

.203

0

.140

7

37

,26

1,0

24

2

1,0

12

,441

4

,500

,00

0

16

,51

2,4

41

0

.283

%

0.4

75

%

0.0

14

%

0.3

26

0

.200

0

.167

8

33

,36

8,1

86

2

0,7

67

,327

4

,500

,00

0

16

,26

7,3

27

0

.291

%

0.4

05

%

0.0

14

%

0.3

42

0

.210

0

.145

9

27

,10

5,8

22

2

1,0

28

,409

4

,500

,00

0

16

,52

8,4

09

0

.294

%

0.3

52

%

0.0

14

%

0.3

03

0

.182

0

.127

10

5

8,8

19

,836

9

,144

,48

8

4,5

00

,00

0

4,6

44

,48

8

0.3

00

%

0.4

66

%

0.0

15

%

0.3

30

0

.196

0

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tal

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4.2 Mining Investment and Operating Cost Estimate In order to provide a check on mining contractor quotations the capital and operating costs for the mining operation at Kevitsa have been estimated from “first principles” based on the mining schedule, in Table 4-4.

The following principles have been applied.

A specific gravity of 3.15 was used for ore and waste; The cycle times and other operational data such as fuel consumption and operating life of the machines are taken from the Caterpillar Handbook Edition 28; Equipment costs are from in house data Equipment operating costs are based on in house estimates and include a provision for tyre wear. Because earth-moving tyres are in short supply, there may be problems in replacing tyres, and the unit costs may be higher in the short/medium term. No allowance has been made for this; Personnel costs are based on the estimates of GTK. Only those personnel directly involved with the earth-moving operation are included. All other personnel are assumed to be common for the contract mining or owner mining cases. Fuel costs are based on the Finnish off-road diesel cost to large users An all diesel fleet option has been chosen because of the added flexibility. As part of the Full Feasibility Study a review of electric shovel operations should be made if owner mining is considered Cost calculations have been made in USD (or USD 000 where indicated) and where converted to Euros an exchange rate of $1.212:€1 has been used

Capital and operating cost estimates by year for the life of the mine are given in Table 4-5 and Table 4-6 respectively overleaf.

The capital replacement schedule is as per the Caterpillar recommendation. This may change with experience in the specific operating conditions. Some of the replacement costs in the latter stages of the operation may be reviewed in the light of experience, but will not have significant present value.

The capital and operating costs apply to the earth moving operation only, and are for the purpose of comparing with a quotation from a mining contractor. Technical functions such as survey, geology, grade control, health and safety, environmental protection are included in the mine owner’s costs (Section 13).

The life of mine average operating cost is estimated at €1.20/t of material moved.

The total capital cost for the mining operation of €42.66M for the equipment is equivalent to €29.6M (discounted at 7%) at today’s value or €0.13/t moved for the initial mining plan considered.

If a profit of 20% is assumed then one can expect a contract mining cost to be of the order of €1.60/t (=(1.20+0.13)*1.20). This supports a verbal quotation of €1.60/t from a Finnish contractor who wished to remain anonymous for commercial reasons.

St

Barb

ara

Consu

ltancy

Serv

ices

26

E705 –

Kevitsa

Pre

Feasi

bili

ty S

tudy

Ta

ble

4-5

: M

ine

Ca

pit

al C

os

t S

ch

ed

ule

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000

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ss

oth

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uro

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52

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57

St

Barb

ara

Consu

ltancy

Serv

ices

27

E705 –

Kevitsa

Pre

Feasi

bili

ty S

tudy

Tab

le 4

-6:

Min

e O

pera

tin

g C

ost

Sch

ed

ule

– E

xclu

din

g S

taff

Su

perv

isio

n (

$000’s

un

less

oth

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ise s

tate

d)

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ment m

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Exp

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es

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St Barbara Consultancy Services 28 E705 – Kevitsa Prefeasibility Study

5 Ore Types, “Sulphur Sufficiency” and Representativity of Metallurgical Test Samples This Section provides a short discussion on those aspects of ore type which can have a potentially material impact on mineral processing by the flotation method and should be read in conjunction with Section 6.

5.1 Ore Types The ore typing methodology employed by Lamberg has been used in the resource estimation methodology already described to separate the composite drillhole data into relevant subpopulations representing separate magmatic events. This has been useful for carrying out variography.

As far as the mineral processing is concerned the orebody can be considered to consist of two predominant ore types:

“Main”

“Nickel-PGM”

The main difference between them is that the nickel-PGE type has a higher Ni/Co ratio than the Main type, and has a tendency towards lower sulphur and copper contents. The main ore type makes up 93 % of all ore blocks with more than 0.2 % nickel to a depth of 330 m (Appendix 5A). Mineralogically the two ore types have the same gangue and sulphide minerals – pyrrhotite, pentlandite and chalcopyrite and therefore each respond in a similar way to the sequential copper and nickel flotation process proposed. With lowering sulphur content the pyrrhotite content diminishes and nickel sulphides without iron (millerite and heazlewoodite) start forming which are considered to float in a similar manner to pentlandite. However, as sulphur content reduces beyond a certain limit (discussed in 5.3 below) some of the nickel can occur in the olivine and not be available for recovery by flotation.

The mineralisation outcrops on surface and an oxidised zone of a few meters above the water table has been noted with poorer and less well understood flotation behaviour (Section 6.2.5). This oxidised material, expected to represent less than 2% of the total ore mined will be set aside in the first year of operation. It will then be blended in with the unoxidised ore, enabling the flotation redox (reducing/oxidation) conditions to be adjusted according to normal industry practice.

5.2 Sample Representivity and Prediction of Performance In any series of metallurgical tests the representativity of the testwork is determined by the closeness of the sample tested to that of the plant feed in the overall life of mine (LOM) plan. Table 5-2 below summarises and compares the overall planned plant feed grade with the three test programmes considered most relevant to the prediction of metallurgical performance in the current sequential flotation of copper and nickel proposed.

St Barbara Consultancy Services 29 E705 – Kevitsa Prefeasibility Study

One can note that for Laboratory Test 18 and the mini pilot test programme for the production of separate copper and nickel concentrates the head grades of the economic elements are broadly in line with the ore proposed to be mined in this study. The LOM platinum grade is somewhat higher than the test samples while, conversely the gold grade is lower. In order to predict the metallurgical performance from any specific feed grade the following correlations were developed from the laboratory and pilot tests:

Copper recovery vs. copper feed grade

Nickel recovery vs. nickel feed grade

Gold recovery vs. copper recovery to copper concentrate

Copper, cobalt, gold, platinum and palladium recoveries vs. nickel recovery to nickel concentrate

This methodology is considered appropriate since the recovery behaviour of the precious elements and cobalt is primarily dependent on their association with the main economic minerals, pentlandite and chalcopyrite. The correlations developed are given in Appendix 6F.

In addition to the obviously important parameters of grade, the concept of “sulphur sufficiency” was introduced to deal with the issue of non sulphide nickel or nickel in silicates which will not be recovered by the flotation process proposed.

5.3 Sulphur Sufficiency The following table summarises the composition of the main sulphide minerals at Kevitsa.

Table 5-1 Composition (%) of main sulphide minerals from Kevitsa

Mineral Sulphur Iron Nickel Copper Cobalt Note

Pyrrhotite 38,3 61,4 0,17 0,01 - Analysis

Pentlandite 32,6 33,2 31,6 0,0 1,4 Analysis

Chalcopyrite 34,9 30,4 - 34,6 - Ideal formula

St

Barb

ara

Consu

ltancy

Serv

ices

30

E705 –

Kevitsa

Pre

feasi

bili

ty S

tudy

Tab

le 5

-2 S

um

mary

of

Imp

ort

an

t M

eta

llu

rgic

al T

est

Sam

ple

s

Hole

Fro

m (

m)

To (

m)

Cu

(%)

Ni

(%)

Co

(%)

S (%)

Pt

(ppm

)Pd

(ppm

)Au

(ppm

)Sulp

hur

Index

1Sulp

hur

Num

ber2

LO

M p

lan

t fe

ed

0.4

03

0.2

93

0.0

14

1.6

40

.32

2

0.2

09

0.1

44

2.3

60

.94

Test

18

xxx

Xxx

xxx

0.4

43

0.2

63

na

1.6

90

.22

2

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83

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92

.39

0.9

8

Min

i p

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t p

lan

t te

st,

sep

ara

te c

on

cen

trate

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2

44

148

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40

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13

1.6

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87

0.1

10

0.1

37

2.3

80.9

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3

24

138

0.5

28

0.2

94

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14

1.9

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33

0.1

55

0.1

62

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9KV-2

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88

200

0.5

42

0.3

67

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17

2.3

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22

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93

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91.4

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54

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63

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67

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A

naly

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0.5

17

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24

na

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50

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6

0.1

92

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82

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41

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t p

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t te

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lk c

on

cen

trate

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5

7

115

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21.9

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8

0.2

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72.4

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6

128

202

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7

0.1

70.1

53.0

11.5

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0

290

353

0.6

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60.0

22.2

10.2

1

0.1

40.1

42.3

11.2

5H

ead 1

Calc

ula

ted

½ c

ore

0.5

10.3

30.0

16

2.1

50.2

6

0.1

71

0.1

53

2.5

61.3

1

An

aly

sed

0.5

50

.33

na

2.4

90

.27

0

.23

10

.16

92

.83

1.6

1

Head 2

Calc

ula

ted

¼ c

ore

0.5

10.3

30.0

16

2.1

30.2

6

0.1

74

0.1

55

2.5

51.3

0

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aly

sed

0.5

20

.34

na

2.1

90

.26

0

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00

.14

92

.54

1.3

3

Note

s: 1.

Sulp

hur

Index =

S/(

Ni+

Cu)

2.

Sulp

hur

Num

ber

= S

-Ni-

Cu


Pederson (Appendix 5A) shows that the weight per cent of nickel is similar to the weight percent for sulphur in pentlandite and weight percent of copper is similar to sulphur in chalcopyrite. The amount of sulphur taken up in pentlandite and chalcopyrite is thus equal to the sum of nickel and copper. The Sulphur Index = (S/(Ni+Cu)) is thus a measure for whether there is enough sulphur available to form pentlandite and chalcopyrite, considering that the formation of pyrrhotite is a later process.

In order to validate the theory that Sulphur Index (or alternatively Sulphur Number) was a reasonable indicator of the amount of non sulphide nickel present the following analysis was done:

Figure 5-1 Estimated Fraction of Sulphide Nickel vs. Sulphur Index

Notes:

1. Data circled in red is near surface and low nickel grade

2. The “bromine method” of extraction assays only sulphide nickel whereas the “aqua regia method” measures total nickel

With the exception of the 3 points thought to be oxidised material the bromine assays appeared to support the sulphur sufficiency theory.

In order to test the theory further, with more data a series of approximately 200 assays was done using the “citrate method”, used in preference to the “bromine method” for health & safety/environmental reasons. However it was discovered that this method as used by the GTK laboratory, when applied to the copper assay did not produce 100% sulphide copper, even on unoxidised samples. This was initially thought to be due to the weaker chemical “attack” not totally liberating the chalcopyrite and therefore the estimated non sulphide nickel assay was corrected as follows:

Non Sulphide Ni assay estimate = Citrate nickel assay x total Cu assay / citrate Cu assay

The fraction of sulphide nickel in a sample can then be determined as:

1-(Non S nickel assay estimate/Total Ni assay by aqua regia)

Ni(brom)/Ni(Aqua Regia) vs. Sulphur Index

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

0.90

1.00

0.00 1.00 2.00 3.00 4.00 5.00

Sulphur Index (S/(Ni+Cu))

Es

tim

ate

d F

rac

tio

n o

f S

ulp

hid

e N

ick

el


This is considered a reasonable first approximation to correcting the citrate nickel assays until a re-assaying by the citrate method is completed, which GTK have undertaken to do.

The following figures demonstrate the relationship between estimated non S nickel assay and fraction of nickel occurring as sulphide and Sulphur Index.

Non Sulphide Nickel (ppm) vs Sulpur Index

0

500

1000

1500

2000

0.00 2.00 4.00 6.00 8.00 10.00 12.00 14.00 16.00 18.00

Sulphur index S/(Cu+Ni)

No

n S

ulp

hid

e N

i (p

pm

)

Figure 5-2 Estimated Non S Nickel vs. S Index

Proportion of Sulphide Nickel vs Sulpur Index

0.00

0.20

0.40

0.60

0.80

1.00

1.20

0.00 2.00 4.00 6.00 8.00 10.00 12.00 14.00 16.00 18.00

Sulphur index S/(Cu+Ni)

Fra

cti

on

Ni a

s S

ulp

hid

e N

i

Figure 5-3 Fraction of Sulphide Nickel vs. S Index


Clearly, one can see that below a Sulphur Index of 2-2.5 the amount of non sulphide nickel starts to increase.

In Figure 5-3 those data points with a S Index above 2.0 have a non sulphide nickel fraction of 0.885 compared to 0.780 of those below, a 12% relative lower expected nickel recovery.

The table below estimates the distribution of S Index within the blocks above a 0.19% Ni cut off.

Table 5-3 S Index of Kevitsa Blocks > 0.19% Nickel

S IndexVolume(Mm3) Mt Co Cu Ni Pt

AverageS Index In class

Cumul. Dist’n. S index

0.0001 -> 1.0 0.660 2.1 0.009 0.071 0.279 0.544 0.66 0.51.0 -> 1.2 1.433 4.5 0.01 0.191 0.426 0.438 1.14 1.11.2 -> 1.5 6.710 21.1 0.01 0.202 0.375 0.388 1.35 4.91.5 -> 2.0 25.419 80.1 0.013 0.448 0.326 0.366 1.81 18.72.0 -> 101.425 319.5 0.014 0.454 0.268 0.183 2.59 74.8Grand Total 135.647 427.3 0.014 0.436 0.286 0.236 2.36 100.0

It is important to note that the overall S Index of 2.36 is the same as that in the Life of Mine Plan and within 0.08 of both the Test 18 sample and the mini pilot scale sample. Therefore one can conclude, on the basis of the forgoing arguments that these two samples are representative, in terms of sulphur sufficiency of the ore to be processed and no discount to the nickel flotation recoveries achieved is warranted at this stage.

Parkkinen concludes that the sulphur deficient zones:

Distribution reflects the same or similar structural features as those controlling high sulphide bodies;

Are concentrated in zones next to high sulphide bodies;

Are highest near surface.

The sulphur deficient volumes seem to concentrate in subvertical northerly zones, especially along 3498800 and around the intersection of 3498800 and 7512325. They favour a northerly direction dipping steeply to the east or 100-110/80-85.

Figure 5-4 below gives an indication of the spatial distribution of sulphur deficient blocks.


Figure 5-4 Distribution of Sulphur Deficient Blocks (above 0.19% Ni)

Moreover, the patterns of sulphur deficiency resemble those of lower specific gravity. These features may imply that sulphur depletion has something to do with rock alteration and supergene processes. It seems to be related to the youngest vertical northerly structures rather than to the oldest gently dipping structures.

It could be argued that the block modelling methodology may have “smoothed” the calculation of S Index and therefore underestimated the proportion of sulphur deficient zones. Until this argument is refuted and the citrate assays repeated it is recommended that a factor of 0.95 is tested as a sensitivity to the base case nickel recovery


6 Metallurgy – Test Work and Design

6.1 Overview Historical testwork indicated that by flotation only bulk sulphide concentrates could be produced successfully. Testwork carried out at VTT (Technical Research Centre of Finland), Lakefield Research and Cardiff University, demonstrated that one of the main factors which would influence the flotation performance is the composition of the ore to be processed; factors such as Ni:Cu ratio, proportion of nickel in sulphides and degree of oxidation. SRK Consulting proposed a combined process: production of a bulk concentrate by flotation and dissolving and recovering of the valuable elements by the PlatSol® process.

Table 6-1 Average Grades of Feed, Bulk Concentrate and Recoveries in Bulk Flotation (SRK Consulting, December 2003)

Ni (%) Cu (%) Co (%) Au (g/t) Pt (g/t) Pd (g/t)

Ore 0.21 0.30 0.01 0.11 0.25 0.16

Concentrate 10.3 5.5 2.1 11.0 7.5

Recovery 72 85 69 47 58 58

In 2005 a successful bench scale test series for producing separate Cu and Ni concentrates was carried out. The results from test 18 are shown below.

Table 6-2 Bench Scale Test Results Product Wt Cu Ni Pd Pt Au

Test 18 % % R-% % R-%

g/t R-%

g/t R-%

g/t R-%

CuConcentrate

0.8 26.4 45.9 0.5 1.5 3.1 11.8 4.8 14.5 4.2 19.4

NiConcentrate

1.8 6.5 26.1 10.0 65.1 4.0 35.3 5.1 35.9 1.0 10.3

Feed 100 0.4 100 0.3 100 0.2 100 0.2 100 0.3 100

Also, in 2005 a successful mini pilot test run for producing separate Cu and Ni concentrates was conducted. The results from pilot run 8 are shown below.

Table 6-3 Mini Pilot Test Run Results Product Wt Cu Ni Pd Pt Au

% % R-% % R-% g/t R-% g/t R-% g/t R-%

Cu, Test 8 1.6 24.5 74.7 0.6 2.7 1.9 16.7 2.6 19.6 2.3 23.3 Ni, Test 8 1.9 2.9 10.7 12.0 67.8 2.3 24.6 1.9 16.8 0.6 7.5 Feed 100 0.5 100 0.3 100 0.2 100 0.2 100 0.2 100

6.2 Testwork

6.2.1 Summary of Historical Testwork

Historical testwork indicated that by flotation only bulk sulphide concentrates could be produced successfully. Significant bulk flotation testwork results are summarised below.


“Preliminary Flotation Study with a Low Grade Sample from Kevitsa Deposit (Helsinki University of Technology)" 1994

This work concentrated on two approaches to make saleable copper and nickel concentrates. The first method was the production of a bulk copper-nickel concentrate followed by a selective flotation of the copper from the bulk concentrate. The second approach entailed the sequential flotation of copper and then nickel sulphides from the ore. The selected sample was noted to be a sample containing 0.3 % Cu and 0.25 % Ni. The results were inconclusive.

“Preliminary Study on Concentration of Precious Metals from Kevitsa Ores (Helsinki University of Technology)” 1994

This report was a continuation of the study above. Five ore types were tested for flotation response, but this was done under differing conditions, which coupled with apparent assay and accounting discrepancies, makes the report inconclusive.

“Additional Beneficiation Tests and Preliminary Leaching Tests of the Kevitsa Ore Samples”, VTT, Technical Research Centre of Finland, Research Report KET4010/94, 1994” Three individual drill hole samples (R713, R725 and R333) and a composite sample (R326 and R360) were tested. The study comprised:

Characterisation of the feed material; Comminution testwork to determine estimated power requirements; Rougher flotation tests at differing pHs, grinds, reagent types and dosages; Cleaner flotation tests on the composite concentrates; Magnetic separation of sulphide materials; Pressure leaching tests on flotation concentrates.

The feed grade of the samples varied from 0.01 % Cu to 0.83 % Cu and from 0.22 % Ni to 0.67 % Ni. It was noted that the silicate mineralogy of these samples differs from conventional Finnish nickel-copper ores. The gangue contains very little serpentine and other altered minerals that normally cause difficulties in processing. The samples R713 and R333 contained hardly any iron sulphides and chalcopyrite. The content of PGE was also high in these two samples.

The comminution testwork indicated that an energy requirement of 23 kWh/t was needed to produce a grind with 60 % passing 45 µm, which was considered necessary at the time to achieve the required recoveries.

Flotation testwork indicated that low pH during rougher flotation improved the recoveries of copper, nickel and platinum group metals (PGMs). The nickel recoveries were 70-86 % (except for sample R333, in which part of the nickel was in silicate form), copper recoveries were 75-90 %, and PGM recoveries 50-85 %. The grade of the bulk concentrate was about 10 times that of the feed at 0.31%Ni and 0.44%Cu.

Production of separate copper and nickel concentrates suitable for conventional smelting processes was carried out with the composite sample. A copper concentrate containing 20 % Cu at 70 % recovery and a nickel concentrate containing 4 % Ni at a 60 % recovery were produced.

Pressure leaching test conducted as a part of this study indicated that some 95 – 97 % of copper and nickel was taken into solution, supporting later testwork results. Also, Outokumpu Mining Company conducted testwork programmes, although no information is available on this.


By using electronic microscope and image analysis VTT concluded that the amount of nickel associated with pentlandite is:

Sample R333: 76.8 % Sample R 713: 84.7 % Sample R 725: 94.0 % Composite sample (R326 and R360): 93.9 %

Sample R333 is characterised as being sulphur deficient with a Sulphur Number of just 0.63.

“The Recovery of Copper, Nickel and PGMs from the Kevitsa Project samples (Lakefield Research 10131-001 – Progress Report No. 1), October 2000”

In addition to their literature review of previous reports, Lakefield Research conducted a limited testwork programme, the primary aim being the production of a concentrate for use in Pressure Oxidation (POX) and PlatSol® Process amenability testwork, as described below. The initial criteria for the bulk sulphide concentrate were:

Combined Cu/Ni concentrate grade 14 – 15 % Sulphur grade >12 %

Three batch flotation tests, F1, F2 and F3 were conducted in order to establish the optimum conditions for the bulk flotation test F4. The high level results of the bulk flotation test were as follows:

5.3 % Cu, recovery 79 % 9.9 % Ni, recovery 68 % 10.6 g/t Pt, recovery 59 % 7.2 g/t Pd, recovery 56 % 2.0 g/t Au, recovery 48 %

Lakefield Research regarded the produced concentrate as a suitable feed material for the POX or PlatSol® processes.

“Kevitsa Flotation Testwork – Factual Report (2695), February 2001”, Cardiff University, Wales, UK

Bench-scale comminution and flotation tests on two low grade samples were conducted by Cardiff University (UK). Initial rougher tests were based on the VTT flotation parameters. The results were, however, relatively poor compared with the previous testwork. A series of rougher tests was conducted with a composite sample with a view to optimise the following parameters:

Grind time; Reagent type – using a combination of Sodium Isobutyl Xanthate and Mercaptobenzothiazole / Dithiophosphate collectors; Activation – through the use of copper sulphate.

Despite the optimisation attempt the sulphide recoveries were still lower than those achieved in earlier testwork. The chosen reagent combination was probably too weak for producing a bulk concentrate at an accepted recovery level. Previous VTT and Lakefield Research investigations had concentrated on the stronger amyl and isobutyl xanthate collectors, which are ideal for bulk sulphide flotation. Furthermore the interpretation of the results is complicated, because of assay reproducibility problems, with back calculated head assays showing too much variability.


“Independent Technical Report on the Kevitsa Ni-Cu-co-Au-PGE Deposit, Finland”, SRK Consulting, Cardiff (UK), December 2003

This pre-feasibility study reviews the results of the previous testwork. SRK Consulting proposes a combined process: production of a bulk concentrate by flotation and dissolving and recovery of the valuable elements by the PlatSol® process. This process had been developed at Lakefield by International PGM Limited. The key stage is the leach step, consisting of a high temperature, pressure oxidation in the presence of minor chloride ion concentration. The PlatSol testwork conducted at Lakefield is reported in: “Recovery of Base and Precious Metals from a Sample of Kevitsa Concentrate LR10131-002”

Testwork carried out at VTT, Lakefield Research and Cardiff University, demonstrated that one of the main factors which would influence the flotation performance is the composition of the ore to be processed. In response, although performance will probably be impacted upon, the modification of various operating parameters would likely mitigate these effects. These would include grind size, pH, reagent addition rates and method of addition, use of activating reagents, and conditioning. The average flotation grade and recovery factors used by SRK in the pre-feasibility study are itemised in table 6.4.

Table 6-4 Average Grades of Ore Feed, Bulk Concentrate and Recoveries in Flotation (SRK Consulting)

Ni (%) Cu (%) Co (%) Au (g/t) Pt (g/t) Pd (g/t) S (%)

Ore 0.21 0.30 0.01 0.11 0.25 0.16

Conc. 10.3 5.5 2.1 11.0 7.5

Recov. 72 85 69 47 58 58

“Metallurgical studies on the High Grade Ni-PGE Ore of Kevitsa”, Geological Survey of Finland, Research Report GTK650/118/04, August 2004

In order to decrease the €420M investment proposed in the SRK Study, SML decided to conduct laboratory scale bulk flotation tests on samples taken from high grade pipes/zones at Kevitsa. The tested drill core samples (totally 55 kg) were: KV-9 (74-78 m), KV-10 (24-82 m), KV-10 (36–38 m), KV-10 (70–72 m), KV-11 (6–10 m). The analyses of the combined feed material and results of two rougher and cleaner bulk flotation tests are presented below.

Table 6-5 Pipe Zone Feed Assays

Pt Pd Au Co Cu Ni S

g/t g/t g/t % % % %

1.09 0.82 0.14 0.01 0.14 0.61 1.06

Table 6-6 Bulk Rougher Flotation

Mass Pt Pd Au Cu Ni Test

% g/t Rec. %

g/t Rec. %

g/t Rec. %

% Rec. %

% Rec. %

19 7.6 11.2 89.3 8.7 90.0 1.5 86.8 1.5 90.0 6.4 86.7 21 6.0 14.3 79.2 11.4 79.6 1.5 69.9 1.7 88.6 7.8 83.5


Table 6-7 Bulk Cleaner Flotation

Mass Pt Pd Au Cu Ni Test

% g/t Rec. %

g/t Rec. %

g/t Rec. %

% Rec. %

% Rec. %

6 3.5 22.1 75.5 17.5 72.1 6.0 73.2 3.7 86.2 13.5 77.8 20 2.7 31.6 73.8 23.8 73.0 3.3 63.0 4.1 84.3 15.7 74.2

“Mini Pilot Run on Kevitsa Ore”, Geological Survey of Finland, Research report No. GTK650/108/05, May 2005

As the tonnages of the high grade Ni-PGE pipes were regarded as relatively small, it was decided to concentrate the efforts on the main ore type representing the major part of Kevitsa deposit. A mini pilot test run was carried out with the following core samples: KV-15, KV-16 and KV-20, totalling 910 kg. The analyses of the combined feed sample are presented in the table below.

Table 6-8 Feed Grades of Mini Pilot Run

Cu Ni Fe S Pd Pt Au MgO

% % % % g/t g/t g/t %

0.54 0.33 6.24 2.39 0.23 0.27 0.16 24.00

The ore feedrate was 15 kg/h. The average result of 13 bulk flotation tests is presented in the table below.

Table 6-9 Average result of 13 bulk flotation tests

Mass Cu Ni PGM + Au

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

6.3 8.1 93.7 3.8 68.9 6.9 66.9

As can be seen the copper final recovery is high. Also the nickel recovery was high in the rougher flotation stage, but about 20 % of nickel recovery was lost in cleaning stage. The reason for the high Ni-loss was thought to be the slow flotation response of fine pentlandite after re-grinding. The Ni-cleaning flotation needed more optimisation.

6.2.2 Bench Scale Test Work Programme: Production of Separate Concentrates

“Copper – Nickel Separation Tests on Bench Scale, Phase 3 metallurgical Study on Kevitsa Ore”, Geological Survey of Finland, Research Report No. GTK650/110/05, June 2005

This study was a continuation of the metallurgical test work on the most common ore type of Kevitsa. About 150 kg of drill core samples were selected from drill hole KV-15 (30-80 m). Analyses of the feed sample are presented below.

Table 6-10 Flotation Feed Analyses

Pt Pd Au Cu Ni Fe S MgO

g/t g/t g/t % % % % %

0.22 0.18 0.25 0.44 0.26 5.4 1.7 23.0


In the previous testwork it was found that the highest grade bulk concentrates contained almost 50 % chalcopyrite. Therefore the objective of producing separate smelter grade copper and nickel concentrates was set. The targets were:

a) 70% Recovery to a minimum 8 % Ni concentrate;

b) Ni in the copper concentrate should be less than 0.8 %, preferably less than 0.5-0.6 %.

Two types of tests were carried out: Cu-Ni separation from bulk concentrate and Cu-Ni separation by sequential flotation. Rougher and cleaning flotation tests were conducted on both test types.

The best results were obtained using the sequential method, first floating the copper concentrate and then nickel concentrate from the copper rougher tailings.

The following reagent regime was used:

Collector: potassium amyl xanthate and sodium isobutyl dithiophosphinateDepressant: carboxy methyl cellulose (CMC) Depressant / dispersant: dextrin Depressant for pyrrhotite: tri ethylene tetramine (TETA) Modifier: sodium metabisulphite Frother: methyl isobutyl carbinol (MIBC) Activator: copper sulphate

The highest grade of nickel concentrate was 20.4 % Ni (test 8), but the nickel recovery was then only 38.5 %. In general, the nickel grade of the concentrates varied between 8-12 % with recoveries around 60 %. Although the Ni-recovery target (70 %) for the final concentrate was not quite reached, the grade-recovery curve revealed that the target level was attainable; in some tests the grade-recovery curves crossed the point 8 % - 70 %.

Au Recovery against Cu Recovery

0

5

10

15

20

25

30

35

0 10 20 30 40 50 60 70 80

Cu Recovery (%)

Au

Reco

very

(%

)

Mini pilot lab test results Test 18 Phase 4 Testwork

Figure 6-1 Gold recovery against Copper Recoveries


Pt and Co Recovery against Ni Recovery

0

10

20

30

40

50

60

70

80

0 10 20 30 40 50 60 70 80 90

Ni Recovery (%)

Pt

an

d C

o R

eco

very

(%

)

Mini pilot test results Reference Test 18 Phase 4 Testwork Cobalt (data from mini pilot tests 5 and 8)

Figure 6-2 Pt and Co Recoveries against Ni Recoveries

The highest copper grade in the final copper concentrate was 26.4 % (tests 15 and 18), but the recoveries remained low in these tests. In test 12 the recovery of copper was 74 % and grade 19.8 %.

Ni content in the copper concentrates of the most promising sequential flotation test was less than 0.8 %.

It was decided to choose the test 18 as the reference for the future test work.

Table 6-11 Bench Scale Sequential Flotation, Analyses of Final Cu and Ni concentrates of Reference Test 18


Test 18 % % R-% % R-% g/t R-% g/t R-% g/t R-%

Cu Cl Con 0.8 26.4 45.9 0.5 1.5 3.1 11.8 4.8 14.5 4.2 19.4 Ni Cl Con 1.8 6.5 26.1 10.0 65.1 4.0 35.3 5.1 35.9 1.0 10.3

6.2.3 Mini Pilot Test Work Programme for Production of Separate Concentrates

“Mini Pilot Run for Sequential Flotation, Test Work on Kevitsa Ore”, Geological Survey of Finland, Research Report No. GTK650/119/05, September 2005

The feed material was a blend of drill core samples KV22 (248 kg), KV-23 (375 kg) and KV-24 (357 kg), totalling 980 kg. The head grades of the mini pilot feed sample are presented in Table 6-12.


Table 6-12 Flotation feed analyses

Cu Ni Fe S Pd Pt Au MgO

% % % % g/t g/t g/t %

0.52 0.32 5.78 2.05 0.19 0.22 0.18 23.22

The feed rate of the mini pilot run was 20 kg/h and the results of 6 of the optimum tests are presented below.

Table 6-13 Mini Pilot Sequential Flotation Run - Copper and Nickel Final Concentrates

The following reagent regime was used:

Collector: sodium isopropyl xanthate Collector: sodium isobutyl dithiophosphinate Depressant: carboxy methyl cellulose (CMC) Depressant for pyrrhotite: tri ethylene tetramine (TETA) Modifier: sodium metabisulphite Frother: methyl isobutyl carbinol (MIBC) Activator: copper sulphate

6.2.4 Bench Scale Testwork on Low and High Grade Samples

“Phase 4 Metallurgical Testwork on Kevitsa Ore: Cu-Ni Separation Tests on Three Ore Samples with Different Grades”, Geological survey of Finland, Research Report GTK650/120/05, September 2005

This study was the continuation of the metallurgical test work on the most common ore type of Kevitsa. The aim of the study was to investigate the effect of feed grade on metal recoveries by flotation. Three ore samples from two different drill holes were used as test material. The description and analyses of the feed material are presented in Table 6-14.


% % R-% % R-% g/t R-% g/t R-% g/t R-%

Cu, Test 2 1.3 24.6 62.5 0.6 2.4 2.2 15.0 2.5 13.8 1.7 14.6 Ni, Test 2 4.7 2.9 25.8 4.5 64.6 1.7 39.6 1.6 32.3 0.9 27.1

Cu, Test 4 0.7 28.4 37.4 0.4 0.9 1.7 5.8 2.6 8.8 2.8 12.2 Ni, Test 4 2.7 8.3 42.8 6.8 56.3 2.0 27.1 1.8 23.3 1.0 17.0

Cu, Test 6 0.8 26.5 42.1 0.5 1.2 1.8 6.8 3.0 10.5 3.7 16.6 Ni, Test 6 1.8 7.2 24.5 10.4 53.4 2.4 19.2 1.7 12.9 0.8 7.6

Cu, Test 7 0.9 26.4 44.1 0.6 1.5 1.6 5.7 2.5 9.1 2.8 11.7 Ni, Test 7 2.1 6.4 26.0 9.0 57.6 2.1 17.3 1.7 14.8 0.7 7.2

Cu, Test 8 1.6 24.5 74.7 0.6 2.7 1.9 16.7 2.6 19.6 2.3 23.3 Ni, Test 8 1.9 2.9 10.7 12.0 67.8 2.3 24.6 1.9 16.8 0.6 7.5

Cu, Test 9 1.4 22.5 63.2 0.7 2.6 1.9 13.7 2.9 18.5 2.8 19.6 NI, Test 9 2.3 3.1 13.7 9.9 62.0 2.3 25.6 1.8 18.4 0.8 9.1


Table 6-14 Description of Feed Samples

Sample no. Drill hole no. Range (m) Description

1 KV-22 102 – 104 Metaperidotite 2 KV-25 57 – 59 Metaperidotite 3 KV-25 103 - 105 Olivine pyroxenite

Table 6-15 Analyses of Feed Samples

Sample Cu (%)

Ni (%) Fe (%)

S(%)

Pt (g/t) Pd(g/t

Au(g/t)

MgO(%)

1 0.17 0.11 3.20 0.63 0.07 0.07 0.06 23.9

2 0.15 0.26 4.98 2.72 0.16 0.20 0.12 24.9

3 0.64 0.61 7.78 3.47 0.71 0.52 0.43 26.3

The flotation tests were carried out according to the procedure used in Reference Test 18 of the previous bench scale testwork. The results are shown below.

Table 6-16 Bench Scale Sequential Flotation Tests on Three Samples of Various Grades, Analyses of Final Cu and Ni Concentrates


% % R-% % R-% g/t R-% g/t R-% g/t R-%

Cu, Test 1 0.1 22.4 14.9 0.4 0.4 2.1 3.6 1.7 1.8 2.7 3.9 Ni, Test 1 0.7 9.6 42.3 8.1 49.1 1.9 20.8 4.9 34.2 1.0 9.7

Cu, Test 2 0.1 26.0 19.8 1.1 0.5 3.1 1.6 1.0 0.7 4.6 4.0 Ni, Test 2 6.2 1.3 55.7 3.5 83.0 1.7 48.9 1.7 65.1 0.9 42.6

Cu, Test 3 1.5 28.3 66.5 1.0 2.5 8.3 24.3 9.7 17.9 8.3 30.9 Ni, Test 3 6.5 1.9 18.5 7.4 77.6 4.2 52.7 7.9 61.7 1.0 15.4

Compared to the metaperidotite samples 1 and 2, the sample 3, olivine pyroxenite, was found to have distinctly higher metal grades and recoveries. The flotation performance of copper in the tests 1 and 2 was poor and consequently a large portion of the copper reports to the nickel concentrate. The overall copper recovery in the tests 1 and 2 is at an acceptable level. The conditioning time and probably also the collector dosage in copper flotation was not optimised.

6.2.5 Flotation Tests on Surface Samples

“Metallurgical Test Work on Kevitsa Ore: Cu-Ni Separation Tests on Near Surface Ore Samples”, Geological Survey of Finland, Research Report No. CM/MT/2006/3

The aim of this bench scale test work was to investigate whether the near surface material at Kevitsa responds to flotation in the same way as the ore samples taken from deeper locations. It was thought that possible weathering of surface layers may have an effect on flotation behaviour.


The bench scale test results of 6 near surface samples are presented in Table 6-17. The feed samples are described below:

Test 1: DDH001 (2.4 – 3.5 m) Test 2: KV-17 (3.1 – 7.0 m) Test 3: KV-23 (2.7 – 6.0 m) Test 4: DDH001 (3.5 – 7.0 m) Test 5: DDH001 (7.0 – 11.0 m) Test 6: DDH001 (11.0 – 15.1 m) Test 18 (Reference test): KV-15 (30 – 80 m)

Table 6-17 Bench Scale Flotation on Near Surface Samples Compared to Reference Test 18


% % R-% % R-% g/t R-% g/t R-% g/t R-%

Cu, Test 1 0.4 24.8 62.9 0.8 2.2 5.1 21.1 4.9 14.2 6.6 19.9 Ni, Test 1 0.7 1.3 5.6 8.5 40.1 3.4 23.0 6.2 29.2 1.4 6.8

Cu, Test 2 0.5 24.5 45.0 0.6 1.1 3.8 6.8 5.5 7.4 2.6 7.8 Ni, Test 2 2.5 1.7 17.6 5.9 58.0 4.6 45.5 5.4 40.1 1.0 15.9

Cu, Test 3 0.2 25.3 30.7 0.9 0.6 1.7 1.9 15.7 1.2 2.9 6.7 Ni, Test 3 1.42 2.6 21.2 11.5 50.9 6.4 50.1 99.2 52.4 0.9 14.3

Cu, Test 4 0.7 25.1 50.1 1.4 4.0 5.4 16.4 6.7 13.4 6.0 22.2 Ni, Test 4 1.0 4.1 12.7 7.8 33.3 4.1 18.9 9.6 29.8 2.5 14.1

Cu, Test 5 0.9 22.3 50.0 1.0 2.5 3.2 9.7 4.7 9.7 4.5 19.8 Ni, Test 5 3.0 3.1 25.0 7.1 61.2 3.6 39.1 6.2 45.7 1.6 25.3

Cu, Test 6 0.8 24.0 67.6 0.8 2.8 5.9 24.2 5.8 14.9 5.0 23.3 Ni, Test 6 1.4 2.0 10.3 9.2 58.4 4.5 34.4 7.8 37.7 1.54 13.3

Cu,Reference Test 18

0.8 26.4 45.9 0.5 1.5 3.1 11.8 4.8 14.5 4.2 19.4

Ni,Reference Test 18

1.8 6.5 26.1 10.0 65.1 4.0 35.3 5.1 35.9 1.0 10.3

It was found that the sample location had a significant effect on the nickel recovery. For most of the near surface samples (2 – 7 m) the nickel recoveries were clearly lower than for samples taken from deeper sections.

The lime requirements and mineral potential values (Eh) are presented in Table 6-18 and Figure 6-3 below:

Test 18: Reference test Test 1: Corresponds to test 1 in table 6-16 Test 2: Corresponds to test 2 in table 6-16 Test 3: Corresponds to test 3 in table 6-16 Test 7: Corresponds to test 1 in table 6-17 Test 8: Corresponds to test 4 in table 6-17 Test 9: corresponds to test 5 in table 6-17 Test 10: Corresponds to test 6 in table 6-17


Table 6-18 Lime requirement for raising pH to 11 Test no. and sample (test date) pH Cu-flotation (pH 11.0) Ni-flotation (pH 5.0)

after grinding Ca(OH)2 g/t H2SO4 g/t

Test 18: hole KV-15, 30-80m (12.5.05) 9,8 604 9575

Test 1: hole KV-22, 102-104m (8.8.05) 9,9 654 6859

Test 2: hole KV-25, 57-59m (10.8.05) 9,7 694 33572

Test 3: hole KV-25, 103-105m (16.8.05) 10,0 504 12704

Test 7: DDH001, 2.4-3.5m (8.2.06) 9,7 720 2029

Test 8: DDH001, 3.5-7.0m (9.2.06) 8,7 2330 1509

Test 9: DDH001, 7.0-11.0m (10.2.06) 9,1 1490 1292

Test 10: DDH001, 11.0-15.1m (13.2.06) 9,7 640 2272

By studying the lime requirement for raising the pH to 11, and the mineral potential values (Eh) in flotation, it was evident that these surface samples had suffered from surface oxidation.

Keivitsa Electrode Potentials Platinum Electrode

-250

-200

-150

-100

-50

0

50

100

Aftergrinding

pH to 11,0 CMC AP 3418A MIBC CuRF1 AP 3418A CuRF2 AP 3418A CuRF3

Eh

(m

V v

s S

CE

)

Test 18: hole KV-15 (12.5.05) Test 1: hole KV-22 (8.8.05) Test 2: hole KV-25 (10.8.05)

Test 3: hole KV-25 (16.8.05) Test 7: DDH001, 2.4-3.5m (8.2.06) Test 8: DDH001, 3.5-7.0m (9.2.06)

Test 9: DDH001, 7.0-11.0m (10.2.06) Test 10: DDH001, 11.0-15.1m (13.2.06)

Figure 6-3 Electrode Potential

This surface material needs additional investigation to understand and mitigate the oxidation effects and it is proposed to undertake this work in the next phase of investigation.

It is proposed that the oxidised section of the deposit is stockpiled and blended in to the initial 2 years of production while controlling the pulp Eh potential with sodium thiosulphate.


6.2.6 Estimated Quality and Recoveries of Cu and Ni Concentrates produced in Full Scale Production at Kevitsa

The estimated quality and recovery figures of produced Cu and Ni concentrates are based on the sequential flotation test results on bench and pilot plant scale. The quality and recovery figures, which vary with grade and type of the ore feed, are presented in the financial model in Section 14 and Appendix 6F Projected Process Performance.

6.3 Plant Design 6.3.1 Process design

The process design, which is presented in appendices 6A, 6B and 6C is based on the parameters used in mini pilot test no. 8. A complete list of process equipment is presented in appendix 12A.

Mining experience in northern Finland has shown that normal production is not halted due to the cold climate in winter. The only interruptions are planned annual service periods, lasting about 1.5 weeks, and occasional production stops. The normal utilisation rate in Finland is 95 %, which has been chosen in this case. The ore feed rate is 541 t/h, corresponding to 4.5 Mt/a.

In practically every new large scale mineral processing project SAG mill grinding is applied instead of the usually more expensive rod mill/ball mill grinding. For the purposes of the Prefeasibility Study it has been assumed that the ore at Kevitsa is amenable to SAG mill grinding. The basis of this assumption is the MKD5 nickel orebody at Mount Keith in central Western Australia which was discovered in 1968. Nickel mineralisation is low grade and occurs as disseminated sulphides interstitial to former olivine grains in serpentinised and carbonate-altered dunites and peridotites; a situation not dissimilar to Kevitsa. Testing of a metallurgical bulk composite sample representative of the ore to be treated over the life of mine is planned in the early part of the Feasibility Phase in order to confirm the SAG Mill approach.

Prior to the copper rougher flotation, the ore is ground to 88 % finer than 75 µm. The nickel rougher concentrate and copper cleaning tailings are reground to a particle size of 99 % finer than 75 µm. The chosen crushing and grinding technique is as follows:

Crushing in one stage with a jaw crusher SAG mill in a closed circuit with a vibrating screen Ball mill in a closed circuit with hydrocyclones

The energy demand in grinding, which has been measured by a Mergan mill, is 16 kWh/t, which is normal with this type of ore. The total energy demand is 30 kWh/t.

The flotation process starts with copper rougher flotation at pH 11. The retention time is 20 min and solids content 35 %. The copper rougher concentrate is cleaned in two steps at pH 12.

The rougher nickel concentrate is produced at pH 5. Part of the nickel flotation feed water (56 %) is recycled to copper flotation in order to minimize the consumption of sulphuric acid and quick lime. For the water separation a conventional thickener and a Lamella thickener in series are used. No flocculants are added at this stage. The water balance of the whole process is presented in appendix 7C.


The retention time of nickel rougher flotation is 40 min and solids content 35 %. The combined nickel rougher concentrate and copper first cleaner tailings is reground and cleaned in two steps at pH 11.5.

The target in nickel rougher flotation is to collect as much sulphide minerals as possible (pH 5). Consequently the corresponding tailings product has a low content of sulphur. By raising the pH to 11.5 in the nickel cleaning flotation a large portion of pyrrhotite and pyrite can be removed as a tailing product. In this way the major part of the tailings is “low sulphur tailing”, which does not require any unordinary method of tailings disposal. The “high sulphur tailing” produced in nickel cleaning flotation is pumped to a separate tailings pond (see chapter 7). If these two tailings product were mixed, the sulphur content would about 1 % S.

Low sulphur tailings (0.4 % S): 478 t/h High sulphur tailings (7 % S): 48 t/h

The copper concentrate (5.9 t/h) and nickel concentrate (9.3 t/h) are thickened and pressure filtered to moisture content of 9 %, ready for shipping.

6.3.2 Energy Consumption and Cost

The total energy consumption and cost are presented below.

Table 6-19 Total Energy Costs

Item Nominal (KW)

1. Receiving hopper & feeder 37

2. Jaw crusher 200

3. Belt conveyor, 1.4 m/12 m 11

4. Belt conveyor, 1.4 m/50 m, 2 units 70

5. Apron feeder, 2 units 44

6. SAG mill, D 8.5 m, L 4.3 m 5,500

7. Vibrating screen 30


9. Belt conveyor, 1.0 m/17 m 15

10. Ball mill, D 7.0 m, L 11.0 m 6,000

11. Regrinding ball mill, D 2.7 m, L 3.6 m 300

12. Conditioner, 2 x 130 m3, 2 x 70 m3, 2 x 50 m3, 2 x 30 m3,1 x 20 m3

700

13. Flotation cell 130 m3, 6 units 900





18. Flotation air blower, 2 units 400

19. High Rate thickener, 9 m diameter (2 units) 6

20. High Rate thickener, 30 m diameter (1 unit) 11

21. Conventional thickener 40 m diameter (1 unit) 19

22. Pressure filter, 52 m2 30






27. Cyclone feed pump, 3 units duty + 2 units standby 750

28. Froth pump, 6 units duty + 6 units standby 111



31. Thickener underflow pump, 2 units duty + 2 units standby 10

32. Thickener underflow pump, 1 unit duty + 1 unit standby 150

33. Thickener overflow pump, 2 units duty + 2 units standby 150

34. Thickener overflow pump, 2 units duty + 2 units standby 10

35. Tailings pump, 2 units duty + 2 units standby 500

36. Tailings pump, 2 units duty + 2 units standby 150

37. Sump pump, 5 units duty 93

38. Waste rock heap and tailing pond ditch water pumps 148

39. Reagent storages and mixing 200

40. Dust control, heating and ventilation 400

41. Return water pump, pipe line 1 km 55

42. Fresh water pipe line from Kitinen river + return water, 5 km

455

43. Mine water removal pumps 110

Total installed power (kW) 19,172

Total installed power x 0.85 16.3 MW

Price of energy 40 €/MWh

Price of energy transfer to Kevitsa 12 €/MWh

Annual electric energy cost (8322 h/a) 7.05 M€

6.3.3 Flotation Chemicals

The usage and costs of flotation chemicals are presented below. The estimation of dosages of chemicals is based on experience from mini pilot and laboratory tests. The chemical prices are based on budget quotations.

Table 6-20 Flotation Chemical Costs

Sodium isopropyl xanthate:

Dosage (g/ton) 250Annual consumption (ton/a) 1,125Price (EUR/ton) 1,500Annual cost (EUR 1 000) 1,688

Sodium isobutyl dithiophosphinate:

Dosage (g/ton) 10Annual consumption (ton/a) 45Price (EUR/ton) 1,700Annual cost (EUR 1 000) 77

Copper sulphate:


Dosage (g/ton) 65Annual consumption (ton/a) 293Price (EUR/ton) 900Annual cost (EUR 1 000) 263

Methyl isobutyl carbinol (MIBC):


Carboxy methyl cellulose (CMC):

Dosage (g/ton) 450Annual consumption (ton/a) 2,025Price (EUR/ton) 1,100Annual cost (EUR 1 000) 2,228

Tri ethylene tetramine (TETA):

Dosage (g/ton) 200Annual consumption (ton/a) 900Price (EUR/ton) 2,750Annual cost (EUR 1 000) 2,475

Sodium metabisulphite:

Dosage (g/ton) 200Annual consumption (ton/a) 900Price (EUR/ton) 100Annual cost (EUR 1 000) 90

Flocculant:


Sulphuric acid:

Dosage (g/ton) 1,000Annual consumption (ton/a) 4,500Price (EUR/ton) 165Annual cost (EUR 1 000) 743

Quick lime:

Dosage (g/ton) 1,500Annual consumption (ton/a) 6,750Price (EUR/ton) 120Annual cost (EUR 1 000) 810

Total annual cost of flotation chemicals (€ 1 000) 9,025


6.3.4 Grinding Media and Mill Liners

The energy demand in grinding is 16 kW/t. The consumption of grinding media and liners has been estimated by utilizing empirical equations and the abrasion index Ai

developed on basis of plant operation data. Wear rate of balls and liners in wet ball mill grinding:

Balls: Unit wear rate (kg/kWh) = 0.159 (Ai – 0.015)0.34

Liners: Unit wear rate (kg/kWh) = 0.0118 (Ai – 0.015)0.3

Abrasion index Ai for copper-nickel ore: 0.46

By taking into account the lower wear rate of SAG mill compared to ball mill, it was estimated that the total wear rate of grinding media and liners of SAG and ball mill would be 1.7 kg/ton of ore. With an estimated market price of € 600/ton of material, the operating cost will be € 1.02 / ton of processed ore.

6.3.5 Maintenance Supplies

The cost of maintenance supplies at 0.54 €/ton of processed ore, was calculated as 5 % of the estimated process equipment expenditure.

6.3.6 Process Investment Costs

The general lay out of the concentrator building is shown in appendix 6E. The complete capital investment cost is presented in section 12 with detail given in Appendix 12A.


7 Tailings and Waste Rock Disposal, Water Balance

7.1 Design Basis for Tailings Ponds Tailings disposal at Kevitsa will be in separate low and high sulphur tailings ponds of downstream earthfill construction, with the latter having a bentonite cover to the embankment. The main design criteria are as follows:

Table 7-1 Pre Feasibility Design Basis for Tailings Disposal

Design Aspect Specification

Designed mine life 15 years

Annual ore feedrate 100.0 % = 4.50 Mt/a

Annual low sulphur (0.4 % S) tailings 88.3 % = 3.97 Mt/a

Annual high sulphur (7 % S) tailings 8.9 % = 0.40 Mt/a

Total tonnage of low sulphur (0.4 % S)

tailings

58.4 Mt

Total tonnage of high sulphur (7 %) tailings 5.9 Mt

Solids content of consolidated tailings slurry 70% by weight

Total volume of low sulphur (0.4 % S)

tailings

44.5 Mm3

Total volume of high sulphur (7 % S) tailings 4.5 Mm3

Particle size distribution of tailings 87.8 % -75 µm, 67.4 % -45 µm

Considering the specific safety aspects due to the cold climate at Kevitsa, it is considered that the downstream method of embankment construction for tailings impoundment is the most appropriate. Construction material for the pervious starter dyke (height 7 m) will be natural till located nearby. For raising the embankment, mainly natural till and if suitable also coarse tailings will be used. The stability of embankments higher than 7 m will be secured by using low sulphur waste rock for construction of the outer face of the impoundment.

7.2 Construction of Tailings Ponds The layout of the tailings ponds and recycle water pond is presented in appendix 7A.

The tailings ponds will be constructed mainly on a bog area to the west of Kevitsa hill. The area of the low sulphur tailings pond is 2 km2 and the high sulphur tailings pond 0.22 km2. Below the peat there is a natural till layer with a water permeability of 10-9 m/s. The existing peat layer (average 2.3 m thickness) will further decrease the water permeability after being compressed to a thickness of 0.5 m by consolidated tailings. Peat will be excavated away from the embankment foundations.


Certain areas (total 0.4 km2) at the base of the low sulphur tailings pond will be improved by adding 1 m of natural till.

The base of the high sulphur tailings pond will be improved by adding 1 m peat (over 0.22 km2). The embankment impermeability will be improved by adding a cover of bentonite clay.

The total embankment height for both tailings ponds will be 26 m as follows:

22 m for consolidated slurry;

0.5 m for compressed peat;

1.5 m for water level fluctuation and,

2 m safety margin.

The embankment will be approximately at the same height as the hill to the west and considerably lower than the Kevitsa hill to the east. Because the natural ground surface is inclining towards the northwest, the water level (and embankment) of the high sulphur tailings pond will be 2 m lower than that of the low sulphur tailings pond. This means that water from the high sulphur tailings pond will not seep towards the low sulphur tailings pond.

The embankment will be raised after 3 production years from 7 m to 14m, after 6 years to 20 m and after 10 years to 26 m.

A cross section of tailings dam shows the construction method and materials used (Appendix 7B).

7.3 Decant and Return Water System The decanting of pond water will be accomplished by pumping from a floating barge. This is a flexible method and used successfully for many years in the cold climate of Finland. The supernatant water is recycled to the mill via a separate decant pond (0.19 m2 on bog area, 7 m embankment). Seepage water collected from ditches around the tailings ponds is also pumped to the separate decant pond.

7.4 Cost of Tailings Dam Construction The unit costs of construction, based on prevailing rates in Finland are given in the following table.

Table 7-2 Unit Costs of Construction of Tailings Pond

Description Unit costs

Removing of topsoil, 1 km transportation 1 €/m2

Removing of peat, 1 km transportation 2.8 €/m3

Construction of embankment, 3 km transportation of natural till 3.5 €/m3

Construction of bentonite cover 10 €/m2

Based on these unit rates the total construction cost of the two tailings ponds and water recycling reservoir is given below.


Table 7-3 Summary of Construction of Tailings and Recycle Water Ponds

Item Cost/€ 1 000

1. Low sulphur tailings pond (2 km2): Starter dams from 225 m to 232 m abovesea level (masl), total length 3.7 km. Base improvement by adding 1 m till,area 0.4 km2

5 030

2. Low sulphur tailings pond: Raising of embankment to 239 masl, total damlength 4.3 km

2 930


3 430


2 300

5. High sulphur tailings pond (0.22 km2): Starter dams from 223 m to 230masl, total length 1.4 km. Base improvement by addition and compaction ofpeat, area 0.22 km2. Embankment improvement by adding bentonite cover

2 580

6. High sulphur tailings pond: Raising of embankment to 237 masl, total damlength 1.4 km

980

7. High sulphur tailings pond: Raising embankment to 243 masl, total damlength 1.4 km

1 260

8. High sulphur tailings pond: Raising of embankment to 249 masl, total damlength 1.4 km

1 820

9. Recycle water reservoir (0.19 km2): Dam length 1.8 km, height 7 m 1 260

10. Tailings pipe line, 3 km, D 0.35 m (steel) 540

11. Reclaim water pump, (1 + 1), 55 kW/pump, pipe line 1.5 km, D 0.35 m(plastic)

180

12. Fresh water pipe line from Kitinen river + return water pipe line, 2 x 5 km,D 0.5 m (plastic). Kitinen pump station: Duty 2 x 200 kW, standby 1 x 200 kW.Return water pump station: Duty 1 x 55 kW, standby 1 x 55 kW

1 140

13. Ditches around waste rock and tailings ponds area, 15 km 10

14. Waste rock heap and tailings pond ditch pump, 8 units duty + 8 unitsstandby, 18.5 kW/pump

230

Total 23 690

7.5 Waste Rock Piles The waste rock piles are presented in appendix 7A. They are located on bog areas. The low permeability (10-9 m/s) of the till layer together with the compressed peat will minimize the seepage of water under the piles. Water from rain and snow is collected and pumped to tailings pond. The sulphur content of the waste rock is on average 1 % S. Layers with higher sulphur content will be covered with layers of lower sulphur content in order to minimize the potential for acid generation.

The design criteria for the waste rock dumps are given in the table below.


Table 7-4 Pre Feasibility Design Basis for Mine Waste Disposal

Generated over LOM

Total tonnage of waste rock 156 Mt

Specific gravity 3.15t/m3

Bulk density 1.8t/m3

Total volume of waste rock 87 Mm3

7.6 Water Balance The net annual effect of rain, snow and evaporation in the catchment area is 0.35 m (0.6 m – 0.25 m).

The open pit area is higher than the surrounding bog area, and no surface water from bogs will flow towards the open pit.

The bedrock is covered by a compact till layer. It is estimated that about 20 % of the annual rain and snowfall (0.6 m) will penetrate this till layer. The bedrock has fractures that collect ground water from a 5-10 km2 area surrounding the open pit. It is estimated that the open pit will annually generate about 1 Mm3 ground water after 10 years of operation.

The water balance flowsheet is presented in appendix 7C with a summary given below.

Table 7-5 Water Balance Summary

Description Water flow (Mm3/a)

Flotation 8.36

Pumps and filtration (fresh water from river Kitinen) 0.80

Recycle water from thickeners’ overflow 4.72

Recycle water from tailings ponds 3.64

Surplus water pumped to river Kitinen 1.89

The process pumps and filtration section of the plant require 0.8 Mm3/a of fresh water from the river Kitinen, 5 km distant.

There is a small brook (Mataraoja) with breeding fish between Kevitsa and the river Kitinen, and therefore surplus water from the tailings pond cannot be discharged freely through the bog areas to the river Kitinen. Surplus water will be pumped back to the river Kitinen via dedicated 5 km pipeline.

During the start-up of operations process water cannot be recycled from the tailings pond as the water level is built up and the corresponding amount of water will be pumped from the river Kitinen (3.64 Mm3/a). Thereafter, the major part of the process water will be reclaimed from thickener overflows (4.72 Mm3/a).


7.7 Recommendations for Further Work The following work is planned in the Feasibility phase:

Foundation investigation and material testing;

Permeability testing;

Hydrogeological investigation, monitoring borehole siting, drilling and test pumping;

Further acid rock drainage investigation, contaminant transportation;

Stability and consolidation analysis.


8 Infrastructure

8.1 Summary The overall site layout plan is given in appendix 7A. This illustrates the proposed open pit, process plant, waste rock piles and tailings and return water dam locations. The new access road, power line and a new bridge over Kitinen River are shown in appendix 8. The following table summarises the proposed investment on infrastructure items

Table 8-1 Summary of Infrastructure Costs

Item Cost/€ 1 000

1. Access road, 6 km 1 200

2. New bridge over river Kitinen, 100 m 1 000

3. Potable water supply + sewage treatment plant 200

4. Electric power line, 110 kV, 5 km 500

5. Connection to Vajukoski power station 200

6. Electric transformer at Kevitsa, 25 MVA, 6 connections for 20 kV 1 070

7. Mine electric power line, 1 km 70

8. Maintenance shop (20 m x 15 m), overhead crane 35 ton 680

9. First aid, changehouse and warehouse (30 m x 15 m) 430

10. Canteen (15 m x 15 m) 350

11. Administration offices (15 m x 12 m) 210

12. Truck service workshop (15 m x 15 m) 290

13. Internal roads, 1 km x 6 m, 1 km x 12 m, 0.5 km x 20 m 340

14. Plant yard, 2 ha 240

15. Fence around open pit 20

Total 6 800

8.2 Access Road and Bridge There is a high quality highway 6 km to the west of Kevitsa, running north from Rovaniemi. The road connecting this highway to the site is currently an unsurfaced logging road that will be upgraded. A new bridge (length 100 m) will be constructed downstream of the hydroelectric dam to cross the Kitinen river, as the road across the dam was not designed for the routine use of heavy vehicles. The route across boggy areas will be crossed using causeways as at present.

8.3 Potable Water Supply and Sewage Treatment The daily need for potable water is 100 – 200 m3. The potable water will be supplied by either ground water or purified river water.

The sewage water (100 – 200 m3/day) will be pumped to a water treatment plant separating 80 % of the total phosphorus. The purified water will be pumped to the tailings pond and the remaining sewage sludge will be transported by tanker to the municipal water treatment plant at Sodankylä.


8.4 Power Supply Finland has a national grid, which is connected to the grids of Sweden, Norway and Russia. Power is generated by hydroelectric dams, coal, peat and nuclear power stations. A fifth nuclear reactor is currently under construction. There is a hydroelectric dam generating 21 MW at Vajukoski, 5 km to the west of the mine site.

The estimated power demand is 16 MW. A new 110 kV power line, 5 km long, will be built from the Vajukoski power plant. A 25 MVA transformer will be constructed at Kevitsa. The estimated market price of electric energy is 40 €/MWh and the price for energy transfer to Kevitsa 12 €/MWh, giving a total electricity price of 52 €/MWh.

8.5 Maintenance Facilities Routine maintenance work will be carried out on site. More complicated tasks will be carried out at Sodankylä (36 km) or Rovaniemi (167 km). Urgent spare parts can be transported by daily flights to Rovaniemi.

The ore transportation trucks will be serviced in a separate truck workshop.

8.6 Other Facilities It is practice in Finland to have a substantial lunch during day shift and also lighter meals may be served in the morning and evening. Therefore a canteen facility is planned.

The facilities are heated by electricity.

A nurse will be available on site during dayshift to give first aid in a small clinic. Routine health care and monitoring will be arranged at Sodankylä. An emergency helicopter service is available from Rovaniemi and ambulance and fire fighting services exist in Sodankylä.

The internal road between the open pit and the crushing station will be 20 m wide (0.5 km). The other internal areas will be served by 12 m wide (1 km) and 6 m wide (1 km) roads. The yard in front of the plant, workshops and administration buildings has a total area of 2 ha.

Local reindeer will be protected by a fence around the open pit.


9 Environmental

9.1 Key Environmental Issues

9.1.1 General

Recent permitting practise in Finland, discussions with authorities and interaction group meetings in Sodankylä have given understanding about important key environmental issues connected to the Kevitsa mining project. According to these a number of keynote rules have been taken as a basis for planning.

9.1.2 Water Effluents Management

Towards the east of the project site there lies the Koitelainen conservation area and the nearest water resource is Satojärvi Lake which is considered to be valuable as a migration resting place for birds. Another arm of the Koitelainen conservation area reaches towards the mine open pit area from the north - east.

Any potential changes to the Koitelainen natural environment could cause limitations being put on the mining operations. Therefore the project design philosophy is that all site water will drain to the west, away from the Natura 2000 site. Groundwater level and fluctuations between the pit area and Koitelainen will be monitored during pit development.

The choice between two water effluent receiving options is proposed to be made in favour of the Kitinen River, which is sufficiently large to dilute treated mine effluents to an insignificant level in terms of water quality change. The other option is Mataraoja Brook, which is a trout production area and is too small to accept mine effluents without harm to the trout population and other fish.

9.1.3 Closure Design

It is likely that according to the Finnish government landfill decision, operations after-care responsibility will, in the environmental permit be stipulated to be for at least 30 years after mine closure. The overall mine design will take this into account.

9.1.4 Eagle Nesting

There is a known nest of eagles in the vicinity and all necessary measures will be taken to avoid any harm to the species and its nesting. Mitigation could include building reserve nest platforms, modifying construction timing and eagle behaviour monitoring.

9.1.5 Dust and Noise Emissions

As mentioned in 9.1.29.1.2 the Koitelainen Natura 2000 area must be protected from the potential impacts of mining operations and the design and placing of buildings, haul roads and access roads will take cognisance of this.

9.2 Introduction This chapter is mostly compiled from existing environmental documents for the Kevitsa project, which are Baseline Study reports, Environmental Impact Assessment (EIA) Programme and preliminary Environmental Management Plan (EMP).

Current environmental work includes progressing the EIA report in close connection with the Project Pre-Feasibility Study and on-going investigations on concentration process development, ore testing and other fields.


The EIA contact authority, Lapland Environmental Centre has on 21.2.2006 given their comment on the Kevitsa EIA programme, amended in October 2005. The statement gives guidance to EIA reporting.

9.3 Regulations and Requirements

9.3.1 Finnish Legislation

Finland has adopted European Union regulation to its own legislation either by changing existing laws or by taking new degrees or acts directly from EU directives.

The EIA procedure for the Kevitsa mining project will comply with the act (468/1994) and decree (268/1999) concerning the procedure for an environmental impact assessment.

The environmental impacts for clarification, as intended by the EIA act, are the direct and indirect impacts caused by a project or operation within and beyond Finland that focus on:

a) The health, living conditions and well-being of people;

b) The soil, water, air, climate, flora, organisms and diversity of nature;

c) Urban structure, buildings, landscape, cityscape and cultural tradition;

d) The use of natural resources, and

e) The mutual interdependency of the factors mentioned in subsections a-d.

In order for mining operations to get underway in the Kevitsa area, an environmental permit that complies with the Environmental Protection Act is required; this is granted by the Northern Finland Environmental Permit Authority based on the submission of an application for the project. An environmental permit covers such matters concerning waste management and groundwater, and soil and air conservation. A permit that complies with the Water Act is required for extracting and channelling water. The application referred to in the Water Act, and a permit application concerning the pollution of water that complies with the Environmental Protection Act and that pertains to the same activity, must be processed together and a joint decision made, unless this is deemed unnecessary for some special reason.

An application for a mining concession for the control and use of the land and water areas required for mining operations is conducted in accordance with the Mining Act, as a result of which a mining concession and the right to mine are granted by decision of the Ministry of Trade and Industry. The mining concession application and the EIA process are being implemented simultaneously in the case of the Kevitsa mining project.

In order to commence mining operations, the concession holder needs the approval of the Safety Technology Authority for the general plan of the mine in accordance with decision 921/1975 of the Ministry of Trade and Industry. The Safety Technology Authority processes and grants licences and notifications concerning industrial handling and storage referred to in the decree 59/1999 on the industrial handling and storage of dangerous chemicals. The Safety Technology Authority is also responsible for inspecting and approving dams in tailing management facilities and other dams as well as the documentation pertaining to them. Mining operations must also comply with the applicable sections of the guidelines for dam safety that have been published by the Ministry of Agriculture and Forestry.

Constructing the necessary buildings at a mine requires planning permission and a building permit in accordance with the Land Use and Building Act; these are granted by the municipal authority. The prerequisites for granting a building permit include municipal planning and municipal decisions concerning allocating the use of areas. The Finnish Road Administration must grant a permit if a decision is made to change the access road to a mine into a public road (the law on public roads). Constructing a power


line (110 kV) requires a permit from the electricity market authorities in accordance with the Electricity Market Act. Land usage at the mine must also take into consideration the regulations prescribed by the Antiquities Act.

9.3.2 European Union Requirements

The proposed Mining Waste Directive will be made effective after several years of preparation and decision making in a few years time. The Kevitsa environmental permit application is expected to take place sooner than that and hence the basis for waste regulation will come from Finnish government landfill decision (4.9.1997/861).

Any operation applying for environmental permits faces the need to meet common Best Available Techniques (BAT) reference documents requirements.

9.3.3 World Bank Requirements

World Bank funding requirements become fulfilled along with the EIA process and Environmental Permitting process. The EIA law and Environmental Protection law are applied in a way which pays full attention to socio-economic factors, especially the effects on indigenous communities in remote areas, the importance of continuing baseline measurements over 12 months so that seasonal variations are recorded, the need for a detailed environmental management plan for the mine, and a comprehensive closure and rehabilitation plan and programme.

The Kevitsa area is classified as a ‘reindeer herding area’, but is not included in the ‘Sámi people area’ which is north of the Sodankylä municipality.

9.4 Current Environmental Conditions

9.4.1 Local Setting

The mining project is situated in the Municipality of Sodankylä in the Province of Lapland approximately 35 km north-northeast of the municipal centre of Sodankylä on the northern side of Kevitsavaara arctic hill (Picture 9-1).

The Project site landscape is dominated by Kevitsa hill which is 1 km from the planned pit area to the south. Otherwise the area consists of forest and bogs in a moderately even terrain. More than 50 % of planned project land use will take place on bog lands (Picture 9-2). Lowest areas are situated on the north-western part of project site, generally some 210 m above sea level. The highest places to be constructed upon are some 240 m above sea level.

The area is 7 km from permanent dwellings in Petkula village in the west and some 2 - 3 km from the nearest holiday cottages by Saiveljärvi Lake to the south of the project site.

9.4.2 Population Structure, Livelihoods and Services

The closest population concentration in the area is the central population cluster of Sodankylä. Habitation in the vicinity of the mining project is concentrated in the villages of Petkula and Moskuvaara. Moskuvaara has a forest machinery contractor, a guesthouse, an earthmoving company and taxi services. Petkula has one farm and it is changing over to goat husbandry.

Kevitsa is part of the extensive Oraniemi Reindeer Herding Cooperative, which is defined by the Reindeer Herding Act as a so-called reindeer husbandry area with special status. However, Oraniemi is not a part of the Sámi Homeland.

The closest reindeer corral to Kevitsa is in Iso Vaiskoselkä, approximately 3.5 km north of the mining area. The site is used 2 - 3 times a year to round up reindeer.


9.4.3 Land Use and Social Structure, Plans Concerning the Area

There are no communities in the vicinity of Kevitsa but there are two villages in the district: Petkula (population about 90 according to the village committee) and Moskuvaara (population 55 according to the municipal web pages). The Vajukoski hydropower plant on the River Kitinen is located in Petkula; it has transformed the village landscape and altered the ways the river is used.

There are a number of “claims” (exploration licences) at the heart of the mining project area, i.e. surrounding the future open pit, which with the exception of exploratory diggings covering a few hectares, are being used by the forestry industry. It is possible to practice reindeer husbandry throughout the area in accordance with the Reindeer Herding Act.

There is a regional land use plan in force for the area: the fourth regional land use plan for Lapland, approved by the Regional Council of Lapland. The plan designates the project area as forestry land. A snowmobile route (kr2021) is marked on the western side of Saiveljärvi lake and there is a power line (220-756) on the northern side.

Most of the land at Kevitsa is owned by the State but there are privately owned strips of meadows at the Satojärvi lake outlet and on the upper reaches of the Sato-oja ditch as well as at the Saiveljärvi lake outlet and on the southern shore. There are two renovated old cottages at Saiveljärvi lake that are still used; they are not accessible by car.

In the Lokka-Koitelainen-Kevitsa section of the municipal plan, Kevitsa and its environs are designated as a soil extraction site, EO, covering an area of more than 80 km2. The designation is used to refer to an area intended for mining activities and soil extraction. A nature economy (berry and mushroom collecting, hunting etc.) as well as agriculture and forestry will be permitted in the area if the activities are not prohibited or limited by other designations in the plan or decisions and regulations based on separate laws.

A new territorial plan is proposed by Lapland council and is open to official and public comments at the moment. The proposed plan keeps the Kevitsa mine area designated to mining, and includes new sections about exploration on conservation areas.

9.4.4 Landscape and Topography

The area is typical of the gently sloping marsh, hill and Arctic hill landscape of Central Lapland, which is characterised by predominantly mixed forests, marshes and smallish Arctic hills. The height of Kevitsansarvi hill and the mining area surrounding it varies between around 200-240 m above sea level. On its southern side is Kevitsavaara Arctic hill with a height of 313 m. On the eastern side of Kevitsa is the more extensive Arctic hill area of Satovaarankuusikko with a height of around 300 m. Satojärvi lake with a surface height of 222 m lies between Kevitsa and Satovaara Arctic hill; Saiveljärvi lake (219 m) is on the southern side of Kevitsa. The landscape in the immediate vicinity of the mining project is thus divided into low gradient slopes that fall westwards and southwards. Diverse marshlands lie in the areas between the hills and the Arctic hills. According to the Internet-based environmental information system maintained by Finland’s environmental administration, the area has no landscapes of any national significance.

9.4.5 Soil

Soil samples from different parts of the area surveyed by the Geological Survey of Finland were taken for testing. The bottom moraine in the Kevitsansarvi area is predominantly sandy moraine, which is a suitable material for building dams. Studies of moraine stratigraphy showed that two moraine beds of different ages occur in the area. The lower bed, i.e. older moraine, is a grey, massive material. It is dense and has a high rock content. The surface bed of younger moraine is looser than the older one and brownish-grey in colour. The grain size in the surface moraine is finer than that in the lower moraine. The contact between the moraine beds is distinct. In the Kevitsansarvi


area, the pits that were dug penetrated through the entire soil cover to the soft rock below. It was not possible to perforate the moraine blanket elsewhere. The moraine and soft rock contact is distinct and sharp. The results for moraine stratigraphy and graininess correspond with those obtained in explorations in the autumn 1993 and spring 1994 (Manninen et al 1996).

Figure 9-1 Location of the deposit (circled in brown). Map taken from the Sodankylä Municipal Internet Service

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9.4.6 Groundwater

According to Manninen et al (1996), the Kevitsa research area can be divided into several small catchment areas in which the direction the groundwater flows in can be worked out with the help of the topography and bedrock surface formations. Due to the thinness of the soil cover, it can be assumed that the topography follows the rock surface formations. Studies have shown the directions the groundwater flows in, based on the relative height of the area and the surface water currents. The riverbeds are relatively even due to their meandering beds and for the most part, the current is slow. Due to the effect of humus, the water is dark in the streams and rivers flowing thorough the marshy areas.

There is no knowledge of classified groundwater areas in the estimated impact area of the project (Britschgi et al 1996). The soil in the project area is generally relatively thin and there are apparently no large groundwater reserves.

Indicatively, there are no natural springs and no artesian exploration drill holes in the area and groundwater outflows only appear on peat lands.

9.4.7 Surface waters

The project is located in the Kemijoki river catchment area and due to the topography described in Section 9.4.4 its waters drain west and south finally to end up in the river Kitinen.

The largest lakes in the vicinity of the project are Saiveljärvi and Satojärvi. Saiveljärvi lake is evidently shallow, only 2-3 m at its deepest. The surface level of Satojärvi lake was once lowered and later a temporary light dam was used to adjust its surface to the present level. Satojärvi is shallow and partially overgrown with vegetation. There are no other lakes or ponds in the immediate area of impact of the project but the Viivajoki river that runs out of Saiveljärvi flows into the lower Moskujärvi lake. The Mataraoja ditch headwaters and ditches and the Sato-oja ditch that runs from Satojärvi lake and drains into Saiveljärvi lake are located in the vicinity of the mining area (Table 9-1).

NB “ditches” in Finland are 1-2m wide are not classed as waterways. They may be man-made but not necessarily.

Table 9-1 Water systems in the vicinity of Kevitsa

Water system No. Catchment area km²

Lakes %

Moskujärvi lake catchment 65.893 103.99 6.40

Mataraoja ditch catchment 65.829 54.72 0.02

Ala-Liesijoki river catchment 65.828 221.19 0.37

Kitinen, Madetkoski river catchment, Vajukoski*

65.822 3428.99 -

Allemaoja ditch catchment 65.926 59.63 0.54

*The catchment area for the Lokka artificial lake is not included.


9.4.8 Climate and Air Quality

Data concerning the climatic conditions in the proposed area of operations (air temperature, relative humidity, cloudiness, precipitation and wind) is available from the Sodankylä Observatory Weather Station, which is about 50 km south of Kevitsa (Table 9-2).

Table 9-2 Weather statistics for Sodankylä. Air temperature, annual precipitation and snow cover. (Drebs et al 2002)

The prevailing wind comes from the south (Drebs et al 2002). The location of the project is remote and according to the Environment Centre Information Service, the area has no significant sources of air emissions. Factors affecting air quality mainly travel come from long distances within Finland and abroad. From the point of view of the closest air emissions, the more important areas are Rovaniemi, where emissions mainly result from traffic and energy production, and Kemijärvi, where there is a pulp mill. The air emissions from the Pahtavaara mine approximately 30 km west-southwest do not reach Kevitsa. The quality classification for the air in Sodankylä is generally good.

9.4.9 Biocoenoses

The area is part of the northern boreal coniferous forest zone. According to Finland’s marsh vegetation area division, the area is located in the Peräpohjola aapa marsh zone. According to the complex mire/bog division, “flark aapas” (ponds between parallel hummocks of mosses) are common to the area. (http://www.oulu.fi/northnature)

Tree stands are sparse, stunted and slow-growing; there are few tree species. There are few species of undergrowth but it is often dense and thick (shrubs, mosses and lichens are common). There are many evergreen species in the coniferous forests in particular (coniferous trees, many shrubs, mosses and lichens). The soil is poor in nutrients, the cycle of nutrients is slow and there are comparatively few grove forests. Podsol soil is common. The phytomass and annual productivity of the forests are rather small. Climatic factors play a large part in all of this. (http://www.oulu.fi/northnature)


Figure 9-2 Typical peat land area at Kevitsa Project. Landscape west of Kevitsa hill (centre). (Photo LVT)

9.4.10 Conservation Areas

The EIA area includes a small part of the extensive Koitelainen conservation area (Figure 9-3) that belongs to the Natura 2000 network.

The nature value of Koitelainen is described on Finland’s environmental administration website as follows:

Koitelaiskaira is one of the most outstanding aapa marsh areas. It has remained as a wilderness area. It is an extremely significant nesting area and habitat for endangered birds and mammals. There is diverse avifauna. The area is included in the list of internationally important wetlands, i.e. a so-called Ramsar site and a globally important bird area (IBA).

Specification of the conservation situation and the means for implementation: the Koitelainen area belongs to the development programme for national parks and nature reserves. In addition, it includes three sites under the name of the Koitelainen extension which are part of the conservation programme for old-growth forests. Conservation of the area will be implemented by means of the nature Conservation Act and the Land Use and Building Act.


KoitelainenNature Directive biotypes: Humic ponds and lakes 1% Mountain lower plains rivers, with Ranunculion fluitantis and Callitricho-Batrachium vegetation <1% Alpine and boreal fell heath 1%

*Raised bog <1% Transition mire and shore bog 5% Fennoscandinavian springs and spring bogs <1% Fens <1%*Aapa marshes 60% *Boreal natural forests 24% Boreal copses <1%*Fennoscandinavian flood meadows 1% *Tree-stand marshes <1% *Alnus glutinosa and Fraxinus excelsior flood forests (Alno-Padion, Alnion incanae, Salicion albae) <1% * prioritized biotype

Nature Directive Appendix II species: *wolverine, otter, Lapland buttercup, yellow march saxifrage Bird Directive Appendix I birds: Merlin, Tengmalm’s owl, hawk owl, eagle owl, golden plover, crane, great grey owl, arctic tern, whooper, sandpiper, capercaillie, black woodpecker, three-toed woodpecker, hazel grouse, bluethroat, northern harrier, ruff, short-eared owl, black grouse , smew, pygmy owl, northern phalarope, 4 endangered species in the area Migratory birds appearing regularly but not mentioned in Bird Directive Appendix I: Little bunting, kestrel, spotted redshank, Other species: Parrot crossbill, Siberian jay, redstart, common crossbill, common sandpiper, pine grosbeak, waxwing, lynx, bear, wolf


Figure 9-3 Koitelainen conservation area borders by Kevitsa (black line – green hatch)

Mining project marked with brown circle (diameter approx.3 - 4 km). (map source: National Land Survey of Finland, LVT permit myy/16/05)

9.4.11 Roads and Traffic

The only road traffic in the proposed operational area is for forestry on the forest roads and traffic related to the diverse use of nature, which serves local needs. There is little traffic on public roads in the district. The number of vehicles on highway 4 comes to 1000 vehicles per day and 72 on the Petkula village road (Pt19809).

A limitation exists by Kitinen River at the Vajukoski hydroelectric power station, where the road crosses the river by the power station bridge, which is private. There is also a public connection via Matarakoski hydropower station 15 km downstream, but closer connection across Vajukoski is preferred. SML has obtained permission to use the private road section by Vajukoski power station.

A snowmobile route goes along the northern shore of Saiveljärvi lake. It continues along the eastern shore of Satojärvi lake and from there, it heads through the Koitelainen conservation area to Lokka village.

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9.4.12 Archaeology and Cultural History

There used to be a heap of stones shaped in the form of a person (a sacred object) that was built on top of a split conical base rock on the top of Kevitsa. In 1907, it was found that the pile of stones had been destroyed but the base rock still remained. Studies in 1993 came to the conclusion that the triangulation on the top of the hill was set up on the base rock of the sacred object.

It was said that man-sized depressions some 30 cm deep could be distinguished at Ristikenttä field on the north-eastern shore of Saiveljärvi lake. These were not found in 1907. It has to be the “unchristened field” mentioned by Tallgren.

There are trapping hollows at the western end of Saiveljärvi lake. Hollows and paths to the hollows can be distinguished on the top of the ridge. Some of the trapping hollows appear to be unfinished.

There are no other cultural environments listed in the Kevitsa area.

The municipal website mentions a monument in the village of Petkula, the “Mataraoja ditch millstones” that rotated powered by water from the ditch and were used for grinding barley in the early 1900s (information from Kauko Äärelä). They are on the lower reaches of the Mataraoja ditch on the eastern side of the River Kitinen.

9.4.13 Diverse Use of Nature

Most of the area needed by the mining project is used by the forestry industry. Ditches and felling have altered the natural state, but valuable natural habitats and areas for conservation under the Forest Act are also found.

The utilisation of natural products has recreational and domestic applications that include hunting, fishing, picking berries, mushrooms and gathering other natural products; the area is ideal for these activities.

Tourism in Sodankylä is concentrated in Saariselkä in the northern part of the municipality and in Luosto to the south. Areas for holiday homes and other sites that bolster water nature and the wilderness are planned for the artificial lakes and the Vajunen basin.

There are no tourist services or outdoor amenities in the environs of the Kevitsa mining project. The Camping Vajusuvanto camping ground operates in summer on the western shore of the River Kitinen in Petkula.

There are duckboards from highway 4 to the birdwatching tower on the shore of Ilmakkijärvi lake west of the village of Petkula.

The water systems in the Kevitsa area belong to the Sodankylä fishing area, to the Petkula statutory local fishery association and the State water area.

The area of the proposed mine is part of the Lapland Game Management District and the Sodankylä Game Management Association’s area. Kevitsan Erä ry’s game management and hunting grounds fall within the sphere of impact of the planned mining project. There are also other hunting clubs in the area (Luiron Erä ry, Koitelaisen Erä ry, Kuritun Jahti ry, Moskun Erä ry, Sodankylän erämiehet ry). The people of Lapland enjoy the free right to hunt on State-owned land in their home municipalities and the private hunting associations hunt some of the game. Alongside moose, hunting for forest gallinaceous birds is popular.

9.5 EIA Process According to Paragraph 2a, Section 6 of the EIA decree, projects for which an environmental impact assessment procedure is applied include “mining metal ore or

other mine minerals, concentration and treatment until the total amount of the separated


substance comes to at least 550,000 tonnes per year or open pit mining where the surface area is in excess of 25 ha.” The Kevitsa mining project fulfils these criteria.

The assessment procedure began when SML submitted the assessment programme to the contact authority (the Lapland Environmental Centre) for the first time in December 2004. The contact authority announced the project and the conditions to be seen in the programme and it is organising in the project area of impact the necessary public meetings where citizens and communities can express their opinions about the project that is the focus of the assessment. Based on the statements, opinions, factors and other additional information expressed about the programme during the public meetings, the contact authority will issue a statement concerning the programme and will state which parts of the assessment programme should be checked.

The environmental impacts of the project will be assessed based on the EIA programme and the statement issued by the contact authority, and the results of the assessment will be presented in the environmental impact assessment report. The contact authority will announce the assessment report in the same way as the programme and it will organize public meetings. It will request the necessary statements concerning the report, and it will provide the opportunity for opinions to be expressed concerning the adequacy of the report. The contact authority will draw up its own statement concerning the report. The EIA procedure will end when the contact authority provides SML with its statement concerning the assessment report. The assessment report and its related statement by the authorities will be appended to applications when applying for permits or their equivalent decisions.

9.6 Baseline Study and EIA Methodologies

9.6.1 Social and Economic Circumstances

The impacts of the project on the economy, employment and regional development can be split into direct impacts that result directly from the added benefit produced by the project and indirect impacts that result from the multiple impacts of the project such as homes for staff, health care, education and new business activities created to serve the project. Large industrial investments will also create positive conditions for business activity in the project area and generate an increase in investment in other industrial fields. However, it is impossible to evaluate numerically the economic and social impacts these positive conditions will generate.

SML has carried out the economic impact assessment which is based on the estimated budget of the mine and existing background data on the impacts of mining projects on the regional and local economies. The economic impact assessment also included an assessment of the project impacts on the livelihoods and services in the area as well as their sphere of influence.

Existing information on other natural resources as well as the plans for their utilisation was collected during the assessment procedure. Data concerning economic and socio-economic matters is collated for synthesis in the EIA report. Based on the SIA study for the Pahtavaara mine (Nieminen 1998), an estimate is presented on the way in which the Kevitsa mining project will affect employment and livelihoods.

9.6.2 Impacts on General Health and Wellbeing

Impacts of the project focusing on health and wellbeing may arise from emissions into the air, groundwater and surface water as well as from noise and vibration. The bulk of the impacts on such things as health will be assessed in the natural environment through the impacts on surface water and air quality. For instance, a change in the quality of the surface water is one criterion for assessing the impact on the usability of household water. Various reports on the expected health impacts of the mine will be combined for presentation in the EIA report. The knowledge of a healthcare expert, such as a doctor


from the Municipality of Sodankylä, on local conditions will be used to assist in assessing the impacts on health. The impacts of the proposed mine on public health are expected to be similar in Sodankylä as those experienced from the mining operations at Pahtavaara.

9.6.3 Roads and Traffic Connections

The report on traffic will produce the basic information for implementing traffic planning and transport at the proposed mine. The EIA will assess the alternatives for transport whilst paying attention to the detriments, risks and other impacts posed to the environment as well as to energy and cost efficiency. The orientation of traffic leaving the mining area and an estimate of the impact of the amount of traffic and accident density will be calculated in the EIA report.

9.6.4 Surface Waters

The major stream in the area is Kitinen River, which is dammed by Vajukoski hydroelectric power station 7 km to the west of the project site. The Kitinen River catchment at Vajukoski is 3430 km2 and the average discharge volume variation is from 15 m3/s to 120 m3/s over a year. Other waterways by the project site are small. There are two lakes – Saiveljärvi Lake, which is 3 km long and 1,1 km wide and Satojärvi Lake which is 1,6 km long and 0,6 km wide. The project site is located on a watershed border area and hence a few headwaters of brooks and ditches have their starting point in bogs.

Impacts of water emissions on surface water will be assessed using a mathematical dilution formula for which the following information will be collected:

• Project water balance and water management plan,

• An estimate of the timing, amount and quality of water emissions, and

• Data on the flows and water quality of the water systems of the area.

The calculation of the impacts of emissions will be performed at the principal junction of the discharge water systems as far down in the water systems as the impacts can be ascertained. The calculations will pay attention to essential changes in the general quality of the water, to impacts caused by nutrient emissions, to changes in metal concentrations in receiving water systems and to the impacts from the possible residue of chemicals used. A change in quality in the ecological state of the water will be assessed according to data on the current state of the water systems and recognized impact mechanisms.

The impacts of the project on the flows and surface levels of the water systems will be assessed by calculating the data on run-off water, flow and surface levels and by using the project plans (channelled and recycled water, water extraction, changes in the catchment area).

An attempt will be made to describe and assess in writing any other possible impacts from the project on the water systems and to present the results, as far as is possible, in a map. The comparison of alternatives will take into consideration the ways in which the water systems are used and their long-term capability to recover after impacts from mining operations have ceased. Methods for reducing the impacts of water emissions will be presented in connection with an assessment of the impact on the water systems. The impact assessment will form part of the EIA report.


Figure 9-4 Baseline water quality sampling sites and Kitinen River discharge volume measurement sites

‘Vajukoski virt. mitt.’ and ‘Matarakoski virt. mitt.’ (map source: National Land Survey of Finland, LVT permit myy/16/05).

9.6.5 Groundwater

As previously stated, there are no natural springs and no artesian exploration drill holes in the project area and groundwater outflows only appear on peat lands.

The assessment of the possible impacts of the mining project on the groundwater will be performed by an expert hydrologist who is familiar with the natural conditions of the area and who is experienced in studying groundwater and in assessing its flow, especially with respect to mining projects. The amount of water arising from the open pit will be considered using model calculations from other mines. The impacts of storage areas for products on the quality of the groundwater will be estimated based on the results of laboratory tests on the materials. If the test results indicate that a significant risk to polluting the groundwater can be expected, the change in the quality of the groundwater can be estimated either through mathematical calculation or by utilizing groundwater modelling.

The EIA has no need to assess the impact of operations on the quality of household water extraction points or wells because none exist in the area.

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9.6.6 Soils

Confirming the natural state of the soil and bedrock makes it possible to assess the impacts from the project. The main impacts on the soil and bedrock that will be assessed are:

• removing soil from areas to be constructed and piling masses,

• mining and waste rock piling,

• physical changes in the soil and bedrock, and

• chemical changes in the soil (clarifying the solubility of metals in the materials and the formation of acidity in process water).

Assessing of the impacts on the soil and bedrock will be performed based on the amount of soil removed, the amount mined and the geochemical environmental characteristics of the rock materials. The assessment will utilize mass balance calculations. The assessment of the impacts of the different alternatives on the soil and bedrock will be performed by a soil expert who is familiar with the natural conditions in the area and the means for implementing the project. Top soils are mostly peat or moraine, with moraine on higher places occasionally covered with boulders. There is a glacial silt body east of Hanhilehto hill and south-west of the open pit area. No sand layers have been found. There are several bedrock outcrops on Kevitsansarvi in the pit area and on Kevitsa hill. Peat thickness varies from 0,5 m to 5 m, average thickness on areas available for waste storage and water storage is 1 - 2 m. Mineral soil thickness is only 0 – 2 m on Kevitsansarvi (future pit area) and up to 10 m on bog areas.

9.6.7 Geochemistry

The Kevitsa ore body outcrops and geological processes have led to some elements migrating into the soil above the ore body. Knowledge of historical glaciation activity leads us to believe that no or very little ore has been mixed mechanically with the soil or transported from the site.

Test results have been used to estimate the risk of acidification in the waste rock and tailings masses to be handled, and possible further measures, such as kinetic solubility tests, will be recommended if short-term tests so indicate. Based on the research results obtained so far, planning of tailings disposal facilities and waste rock dumps takes into consideration the measures to be taken in order to prevent the risk of the formation of acidified water. A programme for monitoring the effectiveness of these measures will be devised.

The report on geochemical risks will provide the basis on how to divide the potential waste rock according to minerals that are completely risk-free, or that pose minimal risk or potential risk as well as their relative ratios; this information will also be used in the environmental planning for the project in both the operational and post-care phases.

Appendices 9A and 9B show recent geochemical testing reports.

9.6.8 Fauna and Flora

The assessment of the project impacts on vegetation will be performed in such a way that areas considered valuable in the report on the natural state and occurrence of endangered plant species will be presented on a map in order to illustrate their location in relation to the areas to be constructed. The impacts of the project on vegetation, such as the drying of marsh areas, will include a written description. The impact assessment for the EIA report will be performed by a biologist.

The assessment of the impacts on avifauna will be performed by an expert in such a way that the significance of areas classified as valuable in the report on the natural state and the importance of the areas to be constructed when implementing the project will be


assessed with respect to the areas surrounding the project. Separate attention will be paid to protected bird species and the nesting territory of large birds of prey that could be disturbed by mining operations and traffic over a more extensive area than that of the mine. The impact assessment for the EIA report will be performed by a biologist.

9.6.9 Fish and Fishing

The assessment of the impacts of the project will be based on the project plans, the assessment of the impacts on water systems and the results of the report on the natural state. EU and Finnish legislation, the suitable classification of water systems as fishing waters, the possible selective detrimental impacts of emission substances on fish and the various combinations of impacts will form the criteria used in the assessment.

The impacts on fishing will be assessed from the point of view of the ability of fish stock to withstand changes in the water quality through paying attention to the prerequisites for reproduction and the consequences of changes in water systems as a result of the project plan. The assessment will pay particular attention to water systems that have considerable value for recreational and domestic fishing or where, from the perspective of preserving biodiversity in the water systems in the region, there are valuable fish stocks. The impact assessment will be performed by a fish biologist or an ichthyologist.

9.6.10 Conservation Areas and Significance of Biological Resources

Koitelainen, which belongs to the Natura 2000 network, is in the estimated impact area of the project. Ilmakkiaapa marsh, which belongs to the Pomokaira Natura conservation area, lies on both sides of highway 4 and the Petkula village road and also alongside traffic route alternative (RT4) for the mine project.

The assessments will examine the natural assets used to justify connecting the areas to the Natura 2000 network and the possible impacts of the project on those natural assets. The impacts of the project on the Natura 2000 network will be assessed by utilising the estimates of the impacts on the natural environment that are put together when performing the EIA. The Natura assessment refers to the obligations imposed in Sections 3 and 4 of Article 6 of the Nature Directive (92/409/EU) to clarify the impacts of projects and plans in certain situations on the natural assets included in the Natura 2000 network. Sections 65 and 66 of the Finnish Environmental Protection Act correspond to the regulations in the Nature Directive.

9.6.11 Landscape

The impacts on the landscape divide into impacts during operations and longer-term impacts as well as into visual and functional impacts. During its operating period, the mine will change areas of marsh or peatland forest, which had been used by the forestry industry, into industrial areas. When operations are wound up, the mine buildings will be torn down but scenic elements such as waste rock dumps, open pit and tailings disposal facilities will remain permanently in the area.

The assessment of impacts on the landscape will be performed as follows:

performing an analysis of the landscape of the area based on local topography, land use and vegetation,

identifying the visual details relating to the project that may change the landscape.

The assessment of visual impacts will be performed in those areas surrounding the mining area from where the mining area is visible in the landscape. Image applications will be presented for the assessment. Such areas include along the access road and other points determined by measuring the relative height of the terrain and the height of the structures for mining operations in relation to the height of the terrain.


The steering and monitoring groups for the EIA process will become an important forum for evaluating the impacts on the landscape, and the opinions of the local residents and the people using the area will be taken into consideration in advance when doing the detailed planning of the mining project.

9.6.12 Air Quality

The study of the emissions resulting from the project will be carried out with regard to dust, traffic and blasting during mining. The emissions will be estimated based on the experience at other mining projects (dust) and calculated from machine output and the amount of explosives used.

The impact area of dust emissions will be determined based on the emission sources and the direction of the prevailing winds. Based on experiences and follow-up results obtained from other mining areas, the sources and extent of emissions can be reliably forecast and limited.

Monitoring, from the point of view of nature and people, the occurrence of detrimental substances and indicators will provide information on the environmental impacts of the mine.

Determining the impact area of air emissions will be performed before the impact areas in the sections below are determined and, for example, before the monitoring point network is planned.

9.6.13 Noise and Vibration

The amplitude of noise emission, the chronological duration and time of day and the spread of noise into the environment will be clarified in the EIA report. The noise estimate will be calculated using generally available models that have been proven to work correctly

Noise will be studied in each zone and special attention will be paid to noise levels in areas for recreational use and holiday homes. The government’s decision (993/1992) on the reference values for noise levels will be used as a basis for comparing the impact values of noise with the maximum permissible noise levels such as at different residential locations.

Vibration from explosions and the impacts of those vibrations will be calculated based on the detailed plans for mining. The results will be used to produce the speed values of the vibrations, which can be compared with the vibration values known to be detrimental to people and buildings. The calculations will be made under the supervision of a person who has made corresponding calculations and comparisons to measurement results at mines currently in operation.

Since the mining project is distant from the population, it is unlikely that vibration would be a problem and therefore, it is probably unnecessary to inspect the buildings within the impact area before mining operations begin in order to ascertain unforeseeable damage resulting from mining operations.

9.6.14 Archaeology and Cultural Heritage

If valuable sites are found in the area during the clarification of its natural state, an estimate of the impacts on these will be drawn up as necessary in cooperation with experts from the Provincial Museum of Lapland. The estimate will pay attention to the way in which the sites are located in relation to the areas to be constructed. Attention will be paid to relics when comparing the alternatives and efforts will be made to find the means to alleviate impacts.


9.6.15 Assessment of Risks to the Environment

The final detailed risk analysis of the project, which is included in the environmental quality management system, will be carried out after the plans have been completed and the decisions for implementing the mine have been made. The safety inspection of the dams to be built and maintained will require drawing up a separate risk analysis. However, the risk will already be assessed during the environmental permit phase on the basis of the preliminary plans for the project.

A preliminary risk assessment, the aim of which is to produce information to support the project plan, will be performed during the EIA procedure. The assessment of the risks will be taken into consideration in the EIA report when comparing the alternatives for implementation. Here, the points to be examined that generate risks include:

transport, warehousing and use of chemicals, fuel and explosive substances;

dam safety;

risks related to the increase in traffic and their prevention;

risks of impacts to the health of the environment;

risks to the environment resulting from a state of emergency;

risks resulting from environmental accidents and their consequences, and

risks resulting from exceptional weather conditions.

9.7 Proposed Environmental Programme Continuation

9.7.1 Reindeer Herders Cooperation

As soon as the project has developed a final mine lay-out and an approved operations strategy, SML will start discussions with Oraniemi reindeer herder’s cooperation. Mine construction and operation will have some influences on reindeer herders’ practices and these will be mitigated.

9.7.2 EIA Process

In accordance with the act, responsibility for arranging official notification of the EIA will fall on the contact authority. This will include notifications of hearing procedures and public meetings concerning the EIA in newspapers and on municipal notice boards. After the EIA programme was submitted for the first time to the contact authority in December 2004, it arranged a public meeting. The views of the environmental authority concerning the EIA procedure and the implementation of the project will be conveyed to interest groups during steering group meetings. The groups will include representatives from other authorities, and the provision of information between SML, the residents in the area of impact and the authorities will become regular and all-inclusive.

SML will also have an active role in providing information. Information on the development of planning for the mining project will be provided publicly in local newspapers and on local radio. Moreover, information will be provided to residents and interest groups through representatives of the municipality, letters to each property and through other forms of contact. Through their representatives, the above-mentioned steering group will become an important spokesperson in associations and the municipality.

A social impact assessment will be performed in order to clarify the expectations, fears and other attitudes of the local residents with respect to the mining project and will during interviews provide the necessary basic information concerning the project.


Bulletins concerning the project will take into consideration at least the following parties:

Municipality of Sodankylä;

Regional Council of Lapland;

Employment and Economic Development Centre for Lapland;

Finnish Road Administration, Lapland District;

Metsähallitus (formerly the Forest and Park Service);

Oraniemi Reindeer Herding Cooperative;

Finnish Game and Fisheries Research Institute;

Sodankylä Fishing District;

Sodankylän riistanhoitoyhdistys ry [Sodankylä Game Management Association];

Lapin luonnonsuojelupiiri ry [Lapland Nature Conservation District Association];

Kemijoki Oy (Kitinen River hydropower company).

Information on the launch of the EIA procedure will be provided through public notification in local newspapers (Lapin Kansa regional newspaper) and on the public municipal information notice board. Other forms of providing information (other newspapers, radio, TV, public meetings) can be used as and when necessary.

Upon its completion, information on the EIA report will be provided publicly in newspapers and on notice boards as with the EIA programme. The contact authority will request the parties and communities concerned to give their statements and opinions concerning the report.

9.7.3 Environmental Baseline Study

EIA baseline studies planned are partially completed at present.

Project development may require more information on soil types and geochemistry, and these studies can be included into environmental baseline data sets.

9.7.4 EIA Report

The EIA report will be drawn up as the result of the studies, environmental impact assessment and interaction in accordance with the EIA programme. It will be drawn up by Lapland Water Research using the assistance of experts in different fields and it will be based on reports of the studies conducted. The matters required by the act and decree concerning an EIA will be presented in the report in order to describe and understand the environmental impact of the Kevitsa mining project in full in such a way that it will provide the basis for granting the necessary permits and for drawing up and approving the plans.

The EIA report will also present an estimate of the compatibility of the best available technology (BAT) for achieving the best method of implementation (BEP) from the perspective of published information and the environment for implementing the alternatives in the project. These factors are some of the criteria for evaluating the comparison of alternatives.

Furthermore, the EIA report will recommend a programme to monitor the impacts of the project during the different phases of mining operations.

9.7.5 Environmental Permit and Water Permit

After completion of EIA reporting an application for the Environmental Permit will be compiled from EIA reports and project design documents. The Environmental Permit and


Water Permit will be dealt with together and decided at the same time. The decision making authority is the Northern Finland Environmental Permit Office.

9.8 Estimate of Environmental Costs Land ownership in Kevitsa is simple and damages estimation to properties could be avoided by leading water effluents to Kitinen river. Hence, shoreline damages do not exist and no financing need be reserved.

The same applies to fish and fisheries compensations, which could be onerous in the Mataraoja brook case. Assuming that the Kitinen river is selected, the estimated annual reservation to project funding for fish compensations is 5,000 euros after construction start.

The reindeer herders’ cooperative is a commercial operating unit and they usually ask for payments against cooperation in reindeer safety, land use, eagle, and waters issues. The annual compensation is estimated at 5,000 euros.

Environmental monitoring is a life of mine activity covering air, waters, fish, tailings, waste rock and other issues. The Environmental Permit decision will form an obligation to monitoring and the monitoring programme will provide the necessary back-ground information to mine management for environmental quality system purposes.

Monitoring will continue after the operational phase and may last in some form to the end of any noticeable effluent from mine site during the after–care period. Environmental monitoring and reporting costs are estimated at 30,000 euros pa during construction and in the operational phase, and 5,000 euros pa after it.

9.9 Closure and Rehabilitation The mine closure and rehabilitation plan will form part of the Feasibility Study and it will be integrated into the environmental permit application. It is prepared with a plan of actions required for the closing of the operation and for the costs associated with them. Later the detailed implementation plans for the after-treatment will be made during the operation of the mine.

9.9.1 Closure Planning and After Treatment

Regulations and Instructions

The basis for the mine aftercare plan is the regulations in the Mine Law and the decision of the Ministry of Trade and Industry on the safety regulations of mines. The orders on the measures to which the operator of the mining activity must adhere when the operation ends have been presented in these regulations. The after-treatment plan includes the closure plan which is based on the Environmental Protection Act (4.2.2000/86) and on Environment Protection Decree (18.2.2000/169). They require regulations from the closure, after-treatment and observation of the operation to be given in environmental permit process.


Decision of the Ministry of Trade and Industry on the safety regulations of mines 28.11.1975/921 sets the following demands for the after-treatment of mines:

128 §

After the mining activity has been discontinued, a mine area must be brought into the

condition required for general safety.

129 §

The open pit must be fenced in a suitable way when necessary and must be equipped

with warning notices. The location of the open pit, the quality and permanence of the pit rim, the water level and water quality must be taken into consideration for people's and

animals' use. Where applicable the regulations of paragraph 1 must be observed if a subsidence risk area exists above an underground mine.

130 §

The slumping of waste rock dumps must be prevented or they must be fenced when

necessary. If there remains a muddy waste area then drying or placement of warning signs and when necessary, fencing is required. Furthermore, care must be taken that a

danger to the environment is not caused from the waste.

Objectives

The purpose of the after-treatment plan is to present the principles of the after-treatment and to describe the scope, development of the measures and costs. Detailed plans are drawn up and updated during the operation of the mine. In the mine area the observation of emissions and their environmental effects will be continued after the cessation of operations and the after-treatment measures have been performed. A separate monitoring programme is drawn up. The following objectives have to be secured with the mine closure plan:

• General securing of safety and health;

• Reduction and prevention of harmful environmental effects, and

• Return of the area to the land use which has been agreed on.

Also, one objective of the after-treatment plan is to ensure that any closure measures taken do not prevent a potential re-start of operations in the future. The closure measures to be taken have to be designed so that they will not, unnecessarily, make the probable later mining activity more difficult.

It is the after-treatment's long term objective to bring the area into such a condition that its care and monitoring require only minor actions by the owner and the area can be given up. To reach this situation the after-treatment contains usually two intermediate stages:

• Active care when the after-care plan is carried out and the executed solutions are maintained;

• Passive care when only minor maintenance actions and monitoring actions are carried out so that the reaching of the set objectives can be secured.

9.9.2 Closure and After-Care Measures

Concentrating Plant Area

The buildings, equipment and structures of the concentrating plant area and electricity lines will be dismantled after the mining activity has ended and the area will be levelled according to the existing contours of the surrounding areas.


The equipment and construction wastes will be transported to designed waste management facilities or sold. The objective is that on in the mine area as few permanent structures as possible will remain so that the area adapts itself as well as possible to the environment. However, the road access is left to serve forestry and other use of the land. Also, some of the buildings in the concentrating plant area may be left another use is found.

The roads which lead to the pit and waste rock dump area will be dug up and the terrain levelled. The road surfacing material can also be used for covering the waste rock dump area. Other parts of the road network may remain for later use.

Water Arrangements

Process water, open pit water system and sewers

The process water pumping plant and water storage structures will be dismantled and the changes that have been made to the shore-line of the river restored. Also all the pumping stations and surface pipelines in the pit and in the concentrating plant area and water and tailings piping systems will be dismantled

Pipelines that have been placed in the countryside will be left in place because they do not cause an environmental hazard and, furthermore, the cost of their removal is large and the disturbance of soil and vegetation would result from their removal again. Unnecessary structures on the lines and all useful devices and materials will be utilised by as far as possible by recycling.

Other dismantled material will be delivered to designed waste management facilities.

Trenching, clarifier basins and surface runoff areas

The water reticulation in the waste rock dump area and of the ditches around the tailings storage facility will be checked and, if necessary, the ditches will be opened and left in use.

The diversion ditches around the open pit will be directed to the pit so that it fills faster. The surface runoff area planned to the north side of the tailings storage facility will be improved, if necessary, and left in operation.

Open pit

The open pit will be allowed to become full of water after the operation has ended. It will take several years to fill. In the after-care period the filling will be accelerated by directing the ditches to the pit. The pit occupies 50 ha size when full the pond will be about 300 m at it’s deepest. Ground level on the north-east side is at its lowest at about +225 masl. The surrounding areas are swampy terrain so the water level in the pit will rise to near this level. Around the pit the ground is mainly level at +230-235 masl, so it will be easy to make slopes and to level the shores of the pond from above the water surface to make the beach safe and useful. The difference in elevation to the water surface on the east edge of the pit is about 20 m over a short distance. To guarantee safety, barriers will be installed around this part of the pit. Moving in the area could be restricted with permanent stone obstacles and with warning signs.

Management of pit waters

The quality of the pit water is determined mainly by the quality of the rock ground water. Pit water is almost totally rock ground water and only a small share is soil ground water or surface runoff and rain. Pit overflow water will be conducted into the clarifier basin and from there into the swamp and on into the Mataraoja through the surface runoff area that has originally been built for the handling of the drainage water from the waste rock dump area.


Waste Rock Dump Area

Amount of waste rock, and heap area

The total size of the waste rock heap area is about 220 ha. Altogether it has been estimated that waste rock production will be about 156 Mt. A filling thickness of about 40 m has been designed in the central part of the area WR1 and a maximum elevation of about +260 masl. The filling will be made so that the slope inclinations will become about 1:3-1:10. The ramps will be steep (1:3) from the edges, but become gentler when going upwards. The filling thickness of the area WR2 becomes about 45 m at around 270 masl maximum altitude. The waste rock area is shaped so that as large a part as possible of the drain water will be made to flow to the west side of the area. The waste rock dump area will be situated towards the swamp area on the north side of the open pit.

Landscaping methods

The waste rock area will adapt themselves scenically better to the surrounding areas if their peaks and slopes are made in varying forms. The peaks of waste rock dumps will be shaped quite evenly. They will be convex shaped generally and slightly hilly.

The choice of the after-treatment method of the surface of waste rock area has an essential effect in the adapting of areas to the surrounding landscape. Because the waste rock does not have harmful geochemical properties, water penetration does not need to be prevented with the surface structure. The after-treatment of the waste rock area can be used to imitate the natural rugged environment. Waste rock will be covered by 0.5 m of moraine.

Before the beginning of the after-treatment, planting tests which serve the final design of the after-treatment will be started in the second waste rock dump. The objective of the tests will be to clarify a suitable covering layer and planting method to carry out the after-treatment efficiently. There are alternative solutions, for example:

• Ploughing of the surface, spreading of the topsoil and planting of the vegetation in grooves, which have been created, and

• Covering of the surface with a thin peat covering and planting of vegetation.

In the planning of planting tests an expert will be used with whom the alternatives for final test arrangements are designed. If the excavation methods do not change, the design and after-treatment of the waste rock area can be done progressively during the operation.

Tailings storage facility

Amount and quality of tailings

About 64 Mt tailings will be stored in the facility, covering an area of about 220 ha. The dam walls are constructed from moraine and waste rock.

To confirm the quality of tailings the contents of the metals will be analysed and solubility tests performed.

Closure method

The tailings storage facility will be allowed to dry. The water surface will lower gradually and at the final stage of the operation waters will be conducted through the water treatment system to the environment. The drain water which comes from the tailings storage area will be directed to a planned surface runoff area. The operation of the field and the distribution of waters will be modified, if necessary.

Tailings will be covered by 0.5 m of moraine. At the same time as the waste rock dump planting tests, tests will be made at the tailings storage facility. As alternative test arrangements the following are considered:


• Light ploughing of the sand surface (to about 30 cm depth), thin topsoil spreading of the layer to sand surface and planting of vegetation;

• Spreading of peat, ploughing of the surface part of the sand and planting of vegetation.

The objective of these tests will be to clarify the most useful method for establishing the vegetation on the surface of the dam. Furthermore, the method chosen in the short run has to prevent dust emission. With the ploughing it should be possible to prevent snow being swept away and to create conditions favourable for revegetation.

9.10 Environmental Cost Estimates

9.10.1 Closure Cost

The costs of after-care treatments have been calculated according to the available plans and according to current costs. The preliminary overall cost estimate is about €6M, detailed below.

Table 9-3 Cost Estimate of Mine Closure

Target Work Amount Unit €/unit €

Plant area - buildings

- machines and devices

- levelling of the area

- electricity line

- demolish

355 000

6 000

m2

m

0,50

2,00

177 000

12 000

Watermanagement

arrangements

- raw water pumping plant

- ditches

- demolish

- restoration 9 000 m 0,10

5 000

900

Open pit - design of edges

55 000 m2 1,25 70 000

Waste rock

dump area

- design of edges

- surface

200 000

1 100 000

m3

m3

1,25

2,50

250 000

2 750 000

Tailings facility - surface 1 100 000 m3 2,50 2 750 000

TOTAL 6 000 000

In the passive after-care period relatively minor costs will arise from monitoring.

9.10.2 Project Implementation Environmental Costs

Cost estimate given after EIA completion costs for the project development.

Permitting

Environmental permit application production after full EIA report and administrative statements cost is estimated to 60,000 euros according to budget quote by LVT. Permitting authority permit decision standard fee in respective cases has been from 20,000 to 35,000 euros. Permit decision regulation include periodical reviews by every 5


to 10 years and permitting authority permit decision fee is then estimated to 20,000 euros. It is assumed that first period will be 5 years and the periods after it 8 – 10 years.

Damages and Compensations

Kevitsa project will not cause any harm to dwellings or to recreational activities.

The only livelihood suffering from the mining will be reindeer herding. The cost of damages and extra work to the mining company will be settled at around €5,000 to €9,000 pa during mine construction and operation.

In the environmental permitting decision the authorities will oblige Project to pay damages for fish and fisheries. Costs are estimated at €2,000 pa throughout mine life.

Vajukoski hydropower station will lose part of its water power through Project water take. It is not possible to give an estimate for compensation at the moment, while permit decisions are not made and the compensation is a matter of agreement.

Monitoring

Environmental permit decisions include addressing of environmental loading and impact monitoring. Monitoring cost €60,000 pa include basic monitoring effort by an independent consultancy.


10 Transportation of Concentrates In conjunction with the marketing and sales studies carried out in section 11 the following options for the transportation of copper and nickel concentrates were costed:

Kevitsa – truck – Rovaniemi – train – Harjavalta smelter

Kevitsa – truck – Kemi harbour – ship – Pori harbour – train – Harjavalta smelter

Kevitsa – truck – Kemi harbour – ship – Rotterdam harbour

Costs were estimated based on the following journey legs:

Table 10-1 Unit Costs of Transportation

Description Cost (€/ton)

Kevitsa – Rovaniemi: Truck 167 km 8

Kevitsa – Kemi harbour: Truck 284 km 13

Rovaniemi – Harjavalta smelter: Train 735 km 18

Kemi harbour – Pori harbour: Ship 610 km 8

Pori harbour – Harjavalta smelter: Train 43 km 2

Kemi harbour – Rotterdam harbour: Ship 22

Table 10-2 Total Costs of Transportation by Destination

Destination Total costs (€/ton)

1. Kevitsa – Rovaniemi – Harjavalta smelter 8+1+18 = 27

2. Kevitsa – Kemi – Pori – Harjavalta smelter 13+1+8+1+2 = 25

3. Kevitsa – Kemi – Rotterdam harbour 13 +1+22 = 36

Notes:

1. 1 €/ton has been added for reloading at Rovaniemi, Kemi and Pori

2. The freight costs are estimated for wet tons, 9 % moisture

3. Ship cargo: 5 000 t

The map in Figure 10-1 shows the approximate locations of the deposit, Rovaniemi, Kemi, Pori and the Harjavalta Smelter.


Figure 10-1 Map of Finland Showing Locations of Mine and Port


11 Marketing

11.1 Introduction A market study was undertaken by John Shaw of St Barbara Consultancy Services to investigate the potential for sale of the Kevitsa nickel and copper concentrates and the revenues to be expected.

The study was based on concentrate analyses provided by Jukka Karhunen of the Geological Survey of Finland from mini-pilot trials undertaken in October 2005 (see Appendix 11A).

The levels of the value metals in the concentrate specifications were:

Ni Concentrate: 12.0 % Ni, 2.9 % Cu, 0.5% Co, 2.3 g/t Pd, 1.9 g/t Pt and

0.6 g/t Au

Cu Concentrate: 0.6 % Ni, 24.5 % Cu, 1.9 g/t Pd, 2.6 g/t Pt and 2.3 g/t Au

Global market reviews were also undertaken by St Barbara Consultancy Services on cobalt and Platinum Group Metals to provide a medium term price forecast for use in the revenue section of the financial model.

Medium term nickel and copper prices were provided by a reputable base metals consultancy firm.

11.2 Nickel The market summary for nickel and copper concentrates is shown in Appendix 11B.

The nickel concentrate is expected to find a ready market in Europe and Canada, although China at this stage looks to be uncompetitive due to lack of payments for the minor metals and the potentially higher freight costs.

Sales to Russia remain a possibility, although their lack of response is probably a reflection of their lack of experience in dealing with non-Russian materials.

The net payable metal contents come to approximately $1,050/t. This net figure is approximately 72% of the payable metal contents. The values derived are used in the revenue model.

11.3 Copper It is assessed that the proposed copper concentrate production from Kevitsa would certainly be able to find a market. In Europe, demand remains very strong for concentrates as a number of smelter expansions are planned including those by Norddeutsche Affinerie, KGHM, Atlantic Copper and Boliden. The un-tapped Russian market is also still a possibility.

The Far East market, however, has become a possible option in recent months as the precious metals prices have strengthened and freight costs have fallen. From a position of not expecting their smelters to be competitive, there is a real possibility of them being in a position to purchase some of the material. Much will depend on the prices of precious metals remaining at high levels.

The nickel in the copper concentrates will require to be kept at the lowest possible levels to avoid penalties.


11.4 Cobalt An overview of the cobalt industry is given in Appendix 11C.

The overview concludes that in spite of the relative abundance of cobalt and improved processing techniques, at present demand for cobalt is exceeding annual production. Reasons for this have been the strong increase in demand, the continuing low production from D R Congo and the legacy of problems associated with the first three nickel/cobalt laterite pressure leach plants in Australia, which have slowed further developments of this technology.

This deficit is likely to continue in the future; although perversely if all the proposed new cobalt developments came to fruition, this could lead to the market being in significant surplus. With the increasing use of cobalt in batteries and chemical products, however, demand is set to grow further over the next few years. Accordingly, perceived wisdom for the medium-term cobalt market is based on the assumption that not all projects will proceed and the supply-demand balance will achieve equilibrium. The caution on behalf of new projects reflects in particular continuing concerns about stability in D R Congo where many of the large projects are located.

The growth in demand for cobalt for batteries for hybrid electric vehicles could increase cobalt demand substantially in the next five years. Future demand will depend on not only economic growth in developed countries but also on the rapidly developing economies of China and India. Demand for cobalt for superalloys is also likely to be strong, with aircraft orders at record levels. The demand for spare parts is also high as airplane capacities increase.

Prior to late 2003, the price of cobalt was drifting down to around $6/lb. The large surge which took cobalt to over $25/lb is now in the past, but the profile of the price decline since this point, coupled with buoyant demand predictions, indicates that a medium term forecast of around $14/lb is not unrealistic.

The St Barbara review recommends that a price range of $12 - $15/lb Co be considered for the medium term and this has been used in the revenue model.

11.5 Platinum Group Metals An overview of the Platinum Group Metals (PGM) industry is given in Appendix 11D.

The report concludes that the medium-term forecast for the PGM market is difficult to predict. New platinum developments are occurring in Southern Africa and Kazakhstan, but information regarding these developments is scarce and it is likely that there are more developments occurring for which there is no information. The industry is extremely secretive, and this has made efforts to investigate the industry difficult.

Currently, the demand for platinum is outweighing the supply. The difference between the two traditionally occurs as a cycle. The price of platinum is dependent upon this cycle, but it is also dependent on the investment side of the platinum market. High investment will cause the platinum price to rise, but this will have a negative effect on the consumers of platinum, who will try and switch to cheaper alternatives, generally palladium, in order to save costs. This will then lead to the price of platinum decreasing, especially as the decrease in price will lead to the selling of investment stocks.

A consensus price forecast for the next five years is in the range $800 - $900/oz and this has been used in the revenue model.

11.6 Summary The sale of copper and nickel concentrates from the proposed Kevitsa mine is assessed as being straightforward. Interest has been shown by a number of potential consumers and good estimates for anticipated concentrate revenues have been able to be made.


The prospects for reasonable prices for by-product cobalt and PGM credits has also been shown to be strong.


12 Capital Costs

12.1 General The capital cost schedule is summarised in table 12.1 below. This incorporates the following principles:

Two years initial full feasibility study with on-going test-work in third year;

Two years construction and commissioning period for process plant;

Mining handled by contractor;

The indirect costs include capital spares and first fill;

EPCM costs set at 5 % of total investment costs;

Value-added tax (22 %) will be returned to the company, and it has therefore not been included in the capital cost list;

A contingency of 20 % has been applied for initial capital (to Year -1) only.

Table 12-1 Capital Cost Schedule (Million €)

Period Total -3 -2 -1 1 4 7 11

Full Feasibility Study

2.00 1.00 1.00

LandAcquisition

0.35 0.35

Mining Equipment

0.45 0.10 0.35

Process Plant 42.62 18.95 23.67

Tailings Facility

23.69 5.14 5.83 3.91 4.69 4.12

Infrastructure 6.80 6.54 0.26

Indirects 4.50 4.50

EPCM 4.00 2.00 2.00

Contingency 14.34 0.20 0.27 6.55 7.32

Total Capex 98.75 1.20 1.62 39.28 43.93 3.91 4.69 4.12

Total initial capital expenditure is €86.03M.

A closure cost of €6M is allocated to the final year of production.

The complete list of capital costs is shown in appendix 12A. For most of the items in the list, budget quotations were asked from two or more companies.


12.2 Full Feasibility Study The full feasibility study cost estimate consists of the following major elements.

Table 12-2 Full Feasibility study costs

Description Cost (Million €)

Further exploration drilling -

Pilot plant testing -

EIA -

Feasibility study 2.0

Total2.0

The required costs of further exploration drilling, pilot plant testing and EIA have already been planned and funded and they are consequently not included in the full feasibility costs.

12.3 Land Acquisition The original total claim area of Kevitsa (squares 1C-9C) was 867.3 hectare (ha). The estimated unit prices of different land types have been provided by the Forest Management Associations (FMA) at Sodankylä, Kittilä and Rovaniemi. The distribution of different land types in squares 1C-9C is shown in table 12.3.

Table 12-3 Distribution of Different Land Types in the Original Claim Area of Kevitsa

Land type Area (ha) Area (%)

Open bog 106.2 12.2

Pine bog 388.1 44.7

Forest 351.4 40.5

Swampy forest 9.8 1.1

Rock field 11.8 1.4

SML intends to purchase a total of 1,200 ha, including squares Kevitsa 12, 13, 15 and 17.

The estimate of the total price for land purchase is € 350 000.

12.4 Closure Costs Closure costs can be found in Section 9.10.1


13 Operating Summary and Manpower Costs

13.1 Summary The operating cost are based on the information presented in sections 5 and 6 of this report and are summarised in table 13.1 below.

Table 13-1 Summary of Operating Costs

Description €/t mined or processed

Mining (contractor) 1.1-2.6

(company) 0.11

Processing (manpower) 0.60

(energy) 1.57

(flotation chemicals) 2.01

(grinding media, liners) 1.02

(maintenance supplies) 0.54

External services 0.10

General & Administration 0.32

TOTAL 7.37-8.87

13.2 Manpower

It is possible to work 365 days a year. However to operate 7 days a week the standard practice has been to operate a 5-panel roster, as compared to the 4 –panel roster found in many countries. Furthermore, operations on Saturday must pay 150 % basic salary, whilst on Sunday this increases to 200 %.

Overtime rates are based on a daily scale: first two hours are charged at 150 % basic rates, thereafter 200 % basic rates. There is a strong preference towards 8-hour shifts as compared to 12-hour shifts.

The standard leave arrangements are 4 weeks per year plus one week winter leave. In addition, the union has negotiated a reduction of 12 days per year, which is usually applied as one day’s leave per month. There are 9 public holidays. The net result is that the standard number of days worked per employee is some 215 days per year.

The required labour complement has been based on the actual labour complement of Finland’s biggest mine, Siilinjarvi phosphate mine, producing 10 Mt/a of ore.

In the tables 13-2, 13-3 and 13-4 the annual costs of mining, processing and support staff are presented. The social cost per person to be paid by the employer is 60 % of the employee’s salary. A management structure chart is given in Appendix 13A.

(All cost figures in EUR 1 000s)


Table 13-2 Mining Labour Complement and Operating Costs

Position No. Cost/ Total

Person/a Cost/a

Mine Production Manager 1 110 110

Geologist 2 90 180

Mine Surveyor 3 65 195

Total 6 81 485

Table 13-3 Processing Labour Complement and Operating Costs


Person/a Cost/a

Process Production Manager 1 120 120

Process Development Manager 1 90 90

Shift Foreman 5 73 365

Process Operators 10 65 650

Chemist/Environmental Officer 1 90 90

Laboratory Technician 2 65 130

Maintenance Manager 1 90 90

Electrician 5 65 325

Instrument mechanic 5 65 325

Mechanic 8 65 520

Total 39 69 2 705


Table 13-4 General & Administration Labour Complement and Operating Costs


Person/a Cost/a

General Manager 1 200 200

G.M. Secretary 1 65 65

Administration Manager 1 90 90

Secretary 1 65 65

Nurse 1 65 65

Accounting Supervisor 1 85 85

Accounting Clerk 1 48 48

Logistics Supervisor 1 85 85

Security Guards 5 48 240

Cad Draftsperson 1 65 65

Canteen Person 4 48 192

Cleaning Person 2 48 96

Warehouse Supervisor 1 65 65

Warehouse Man 1 48 48

Purchasing Agent 1 48 48

Total 23 68 1 457


14 Financial Analysis Appendix 14-1 contains the full discounted cash flow (DCF) financial model for a 4.5Mtpa operation. The project Net Present Value (NPV) and Internal Rate of Return (IRR) is calculated. The key features of the model are as follows:

100% equity

Cash flow before interest, tax, depreciation and amortisation

2006 money terms (i.e. no adjustments for inflation)

Exchange rate: € = $1.212 (set in February 2006 when the forecast dollar metal prices were provided)

Metal Prices

o Copper ($/t) $1.21/lb ($2,668/t) o Nickel ($/t) $4.68/lb ($10,319/t) o Cobalt $15.00/lb o Platinum ($/oz) $750/oz o Palladium ($/oz) $200/oz o Gold ($/oz) $450/oz o Silver ($/oz) $8.00/oz

For ease of understanding the model is annotated with assumptions and comments which can be viewed by hovering the cursor over those cells containing a red triangle.

Cells coloured pink can be adjusted to test sensitivities.

The layout of the model is summarised in the following table:

Table 14-1 Financial Model Layout

Sheet Description

1 Production & Summary, including sensitivity parameters 2 Mine Production Plan 3 Nickel Concentrate Revenue Detail 4 Nickel Concentrate Revenue Detail 5 Mine Cost Summary 6 Operating Cost Summary 7 Capital Expenditure Summary

Sheet 1 provides the Net Present Value (NPV) and Internal Rate of Return (IRR) of the project. The project economics are seen to be most sensitive to total revenue/price, $/€ exchange rate and operating cost as indicated in the following tables.

An illustration of the sensitivity to nickel and copper prices is given in the following table.

Table 14-2 Project Metal Price Scenario Sensitivity (€M)

Price Scenario NPV7% IRR (%)

Ni $10,319/t Cu $2,668/t 91.3 17.5

Ni $12,500/t Cu $3,300/t 219.5 30.4

Ni $14,000/t Cu $4,000/t 328.2 40.5


Table 14-3 Kevitsa NPV7% Project Sensitivity (€M)

+10% Base Case 4.5Mtpa

-10%

Operating cost 49.5 91.3 133.3

Capital cost 83.9 91.3 98.8

Nickel price 128.7 91.3 54.1

Copper price 112.2 91.3 70.6

$/€ Exchange rate 35.5 91.3 159.7

Total revenue 150.2 91.3 32.6

A 2% increase or decrease in discount rate changes the NPV to €62.8M and €129.3M respectively.

Table 14-4 Kevitsa IRR Project Sensitivity (%)

+10% Base Case 4.5Mtpa

-10%

Operating cost 12.7 17.5 22.3

Capital cost 16.2 17.5 19.0

Nickel price 21.5 17.5 13.4

Copper price 19.6 17.5 15.3

$/€ Exchange rate 11.3 17.5 24.6

Total revenue 23.7 17.5 10.9

The average cost of nickel nett of by product copper credits and payable elements in the nickel concentrate is €1.06/lb.

A 5% relative drop in nickel recovery would reduce the project NPV7% to €71.6M and the IRR to 15.3%.

The sensitivity results will form the basis on which the project is further optimised in the Full Feasibility stage.


15 Project Implementation and Schedule The project schedule is presented in tabular form below. The critical path is expected to be:

Environmental impact assessment (EIA) and environmental permitting; Financing and contract award; Engineering and construction and start-up.

The application for planning permission for the proposed road and power upgrades is anticipated to run in parallel with the environmental permitting.

The total project duration is expected to be 3.5 years with the engineering and construction phase taking 1.5 years, including two spring periods. Further drilling will be carried out during a two year period to finalised mine design.

Prior to the full feasibility study, pilot plant tests will be done during a 5 month period on samples representative of the average material to be mined over the life of the Project. The full feasibility study will take 12 months.

The construction of the mine and associated infrastructure, concentrator and tailings ponds will be a turnkey project. Suitably reputable and experienced companies will be approached to provide quotations for the complete engineering, procurement, construction, and start-up. The actual construction of buildings, tailings ponds, roads, power line, etc. will be done by local contractors.

Ramp up to full production is planned over a 6 month period starting in September 2009. In the first 4 months of the ramp up 500,000t are considered to be processed, based on an average throughput of 33% of design. 4.37Mt is then treated in the first full year, 2010.

The environmental impact assessment will be carried out by a respected local company, probably Lapin Vesitutkimus Oy. LVT is currently carrying out the base line studies.


Table 15-1 Project Schedule

MONTHS 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

‘07

M J J A S O N D J F M A M J J A S O N D

Pre-Feasibility Completion 0

Exploration

Drilling x x x x x x x x x x x x x x x x x x x x

Trial Mining &

Pilot Plant Testing x x x x x

Financing,

Contracts x x x x x x

Feasibility

Study x x x x x x x x x x x x

EIA,

Environment

Permit x x x x x x x x x x x x x x x x x x x x

Engineering &

Procurement

Construction

Testing and

Start-Up

St

Barb

ara

Consu

ltancy

Serv

ices

97

E705 –

Kevitsa

Pre

-Feasi

bili

ty S

tudy

MO

NT

HS

2

1

22

2

3

24

2

5

26

2

7

28

2

9

30

3

1

32

3

3

34

3

5

36

3

7

38

3

9

40

4

1

42

‘0

8

‘0

9

J

F

M

A

M

J

J

A

S

O

N

D

J

F

M

A

M

J

J

A

S

O

Pre

-Feasib

ilit

y C

om

ple

tio

n

Ex

plo

rati

on

Dri

llin

g

x

x

x

x

Tri

al

Min

ing

&

Pil

ot

Pla

nt

Te

sti

ng

Fin

an

cin

g,

Co

ntr

ac

ts

x

x

x

x

Fe

as

ibil

ity

Stu

dy

EIA

,

En

vir

on

me

nt

Pe

rmit

En

gin

ee

rin

g &

Pro

cu

rem

en

t

x

x

x

x

x

x

x

x

x

x

x

x

Co

ns

tru

cti

on

x

x

x

x

x

x

x

x

x

x

x

x

x

x

x

Te

sti

ng

an

d

Sta

rt-U

p

x

x


16 Conclusions, Risks and Opportunities The Kevitsa Project at the current prefeasibility stage shows an IRR of 17.5% and an NPV7% of €91.3M using appropriately conservative medium term metal prices and smelter terms based on those expected to be reached following initial discussions with smelters.

Although the metallurgical performance of the project is based on pilot plant testwork the performance of the differential flotation process is not considered to be fully optimised, particularly with regard to minimisation of reagent use and control of pulp Redox potential at the head of the copper circuit. This represents both risk and the opportunity to improve flotation performance.

The following “SWOT” analysis summarises SBCS’s perceived status of the Project.

Strengths

Project 100% owned by SML;

Large resource with 33% of the >0.2%Ni material in the measured and indicated categories;

66.8Mt open pit reserve outcropping at surface with 84% in the Proven category;

Located in a politically stable country in a region favouring mine development;

Project located close to existing infrastructure;

Good markets for products;

Currently excellent metal prices facilitating cash raising for a Full Feasibility Study.

Weaknesses

Moderate ore grades;

Complex metallurgy leading to low grade nickel concentrate at modest recovery.

Opportunities

Current drilling programme should enable an increase in measured and indicated resources permitting an increase in mining rate and/or mine life;

Optimisation of project schedule to bring slightly earlier cashflow;

Optimisation of the process at pilot scale to reduce reagent costs;

Finnish Government grants and European Union structural funding.

Threats and Threat Mitigation

Strengthening of the Euro against the dollar – possibility for hedging;

Non sulphide nickel content is higher than expected – repeat non sulphide nickel assays, correlate with Sulphur Index and incorporate in geological block model as part of the Full Feasibility Study;

Project owner is an experienced explorer but without project management experience – hire in expertise.


17 References SRK Consulting, Cardiff. The Kevitsa Ni-Cu-Co-Au-PGE Deposit, Finland – Pre-feasibility study. January 2004

GTK (http://www.gsf.fi/explor/eco_legis_frame.htm)

Gervilla, F.; Kojonen, K.; Merkle, R. K. W.; 2004. Platinum group Minerals in the Proterozoic Kevitsa mafic-ultramafic intrusion. Sodylankylä, Northern Finland. 4 p.

Gervilla, F.; Kojonen, K.; Parkkinen, J.; Välimaa, J.; Platinum Group Element Mineralogy, Geochemistry and 3-D Modelling of the Kevitsa Ni-Cu-PGE Sulphide Deposit, Northern Finland.

Heiskanen, K., Kirjavainen, V., Laapas, H., Lyyra, M. and Vesanto, A. 1994 (?). Preliminary flotation study with a low grade ore sample from Kevitsa deposit. Helsinki University of Technology, Laboratory of Mineral and Particle Technology. Espoo.

Heiskanen, K., Kirjavainen, V., Laapas, H., Lyyra, M. and Vesanto, A. July 1994. Preliminary study on concentration of precious Metals from Kevitsa ores. Helsinki University of Technology, Laboratory of Mineral and Particle Technology. Espoo.

Kalapudas, R.; Kojonen, K.; Laukkanen, J; Luukanen, S.; Applied Mineralogy and Mineral Processing of the Disseminated Cu-Ni-PGE Ore of the Kevitsa Deposit, Sodankylä, Northern Finland. Geological Survey of Finland, Espoo & Outokumpu. 5p

Kortman, C.; Nurmi, P. A & Vuotevsi, T. (comps) 1996, Introduction to Mineral Legislation in Finland. 2nd Revised Edition. Espoo: Geological Survey of Finland. 8 p.

Lamberg; et al 2005. Structural, Geochemical and Magmatic Modelling of the Early Proterozoic Kevitsa Ni-Cu-PGE Deposit in Sodankylä, Northern Finland. 4 p.

Lamberg, P.; 2005. Nickel and Cobalt Numbers – Novel Methodology for Tracing Critical Processes of Ni-Cu-PGE Ore Formation. 4 p.

Mutanen, T. (1997) Geology and Ore Petrology of the Akanvaara and Koitelainen Mafic Layered Intrusions and the Kevitsa-Satovaara Layered Complex, Northern Finland. Bull. Geol. Sur. Finland, V 395, pp 1-233.

Young, R.S.; 1971. Chemical Phase Analysis. Barnes & Noble Inc, New York. Pages 81-86

Appendix 3A –

Kevitsa 43-101 Geological Report – Refer to SEDAR filing 10 March 2006

Appendix 4A –

Whittle Optimisation Report

KEIVITSA DEPOSIT SCANDINAVIAN GOLD

DRAFT REPORT 1

The Keivitsa Ni, Cu, Co, Pt, Pd & Au deposit. Finland

Mining Report ByCSA International (in conjunction with St Barbara Consultancy Services)


DRAFT REPORT 2

Executive Summary

This document serves as a record of the optimisation, design, and scheduling parameters used for the Keivitsa Ni, Cu, Co, Pt, Pd & Au Deposit carried out for Scandinavian Gold. A detailed geological model was provided and adapted and with the use of a strategic planning software package an open pit design was created. This design was used to assess the mineable reserves of the Keivitsa project and also to provide an annual life of mine schedule.

The geological model was adapted to enable its use in Whittle, the strategic planning software used for the optimisation of the open pit. Waste blocks were added and a density of 3.15 t/m

3

applied to both ore and waste blocks. The respective ore tonnages were calculated along with the contained element tonnages. As no mining activity has occurred at Keivitsa before, quotations from mining contractors were used to estimate the mining cost. To adjust the mining cost by depth the Mining Cost Adjustment Factors for a similar project were used. Processing recoveries were provided by St. Barbara Consultancy Services. These recoveries were applied within the geological model as the equations were too complex for Whittle.

The Whittle optimisation was run on a wall slope set of 45° for the first 30m below surface. The wall angle was then increased to 55° for the remainder of the pit. The metal prices used in the pit optimisation are listed below.

Table 1-1: Metal Prices Element Current Price Forecast ($) Units Ni 10 319 $/t Cu 2 668 $/t Co 33 069.35 $/t Pt 16.0750 $/g Pd 6.4301 $/g Au 14.4680 $/g

A mining dilution of 3% at zero grade and a mining recovery fraction of 0.97 were applied. A constant discount rate of 7% was used due to the low level of risk associated with Finland. A maximum processing plant capacity was set to 4.5 million tonnes per annum with a processing cost of $7.15 per tonne. A selling cost of $429.10 per saleable Ni tonne was ascribed to the Nickel and $208.62 per saleable Cu tonne was ascribed to Copper.

From the applied parameters and Revenue Factors of 0.3 to 1.5, Whittle generated 101 pit shells. The revenue factor used in the Whittle optimisation is an adjustment in the price received for the particular commodity, Whittle uses the user specified range of adjusted “Revenue Factors” to determine a pit value and size based on a particular expected revenue factor. From the results of Whittle determining the optimal size and value for each of the revenue factors a pit shell with the highest pit value or alternatively size may be selected. On analysis of these an optimal pit was chosen based on the specified case scenario. Whittle optimises using three scenarios: best, worst and specified case. The best case is defined as mining each successive incremental pit shell exactly to the shell before the next successive shell is to be mined. The worst case scenario is defined as mining the ultimate pit shell from surface down to the deepest point in the shell in incremental benches, a complete bench cut must be mined before mining can proceed to the next bench. The specified case is defined by the user since both the afore-mentioned scenarios are unrealistic in terms of practical mining application. For this reason the user specifies the pushbacks or major shells and the required bench lead before a successive pushback may begin. The specified case’s results are chosen since it would be the closest to a practical mining plan and hence real-world results. The other two scenarios will then define the variability or risk in terms of mining poorly or within accepted standards and the potential gains to be had if better mining practises are employed.

A design was generated from the optimal pit. The generated design is roughly oval. The pits’ longest axis which runs roughly Northwest Southeast has a length of 900m. On its Northeast Southwest axis the design produced a length of 700m. At its deepest point the pit is 400m


DRAFT REPORT 3

below surface. A single ramp system was incorporated into the design. This was done to ensure efficient ore extraction. At its steepest point the pit wall has an angle of 53.8°.

Figure 1-1 overleaf shows a schematic of the pit layout.

Figure 1-1: Final Pit design

An analysis of the designed pit was conducted against the Whittle pit shell in Datamine (see Table 1-2).

Table 1-2: Design vs. Whittle results

Optimised Ore selected for Processing

Base Case Total Mineral Reserve

Whittle Design Variance

(%)

Total Tonnage (Mt) 223.1 223.9 0.36%

Ore Tonnage (Mt) 104.5 104.3 -0.21%

Payable Recovered Metal

Ni (t) 113,859 110,978 -2.53%

Cu (t) 259,054 253,509 -2.14%

Co (t) 2,595 2,545 -1.93%

Pt (g) 4,089,684 4,004,977 -2.07%

Pd (g) 2,480,565 2,432,052 -1.96%

Au (g) 2,387,794 2,365,595 -0.93%

Feed Grade (Diluted In-Situ)

Ni (%) 0.244% 0.242% -0.96%

Cu (%) 0.350% 0.346% -1.09%

Co (%) 0.013% 0.013% -0.42%

Pt (g/t) 0.246 0.244 -0.92%

Pd (g/t) 0.157 0.155 -1.02%

Au (g/t) 0.115 0.114 -0.65%

Metal Contents (Diluted In-Situ)


DRAFT REPORT 4

Ni (t) 254,979 251,996 -1.17%

Cu (t) 365,799 361,046 -1.30%

Co (t) 13,488 13,402 -0.64%

Pt (g) 25,720,229 25,429,981 -1.13%

Pd (g) 16,375,763 16,174,658 -1.23%

Au (g) 11,964,382 11,861,754 -0.86%

These variations are considered acceptable for the Feasibility Study since they are within 5% of the Whittle Optimisation..

To increase the NPV of the project and attain the required feed grade a cut-off optimisation was performed. A cut-off of 0.18% Ni was selected since it provided the required feed grade of 0.295% Ni and life of mine schedule of 16 years that satisfied the strategic business targets. The Scheduling proved that the NPV curve at the various cut-offs has a relatively flat apex around the maximum that will provide for flexibility in the mining operation. This produced the following Proven and Probable Reserves.

Table 1-3: Mineral Reserve

Mineral Reserve


Proved 0.19% 45.1 Mt 0.295 0.398 0.014 1.628 0.145 0.213 0.328

Probable 0.19% 5.9 Mt 0.279 0.453 0.014 1.793 0.137 0.172 0.267

Total 0.19% 51.0 Mt 0.293 0.404 0.014 1.647 0.144 0.208 0.321

Mineral Reserve


Proven 0.18% 56.2 Mt 0.295 0.415 0.014 1.683 0.141 0.201 0.310

Probable 0.18% 10.6 Mt 0.295 0.492 0.015 1.896 0.142 0.171 0.267

Total 0.18% 66.8 Mt 0.295 0.427 0.014 1.717 0.141 0.196 0.303

Measured Mineral Resource material was converted to Proven Reserves and Indicated Resources to Probable Reserves. Inferred resources were assumed to have zero value for the pit optimisation.

At a throughput of 4.5 million tonnes per annum the reserve will be mined over 16 years with one additional year included for pre-stripping. The open pit contains 223 million rock tonnes of which 66.8 million is ore. This results in an average stripping ratio of 2.34t waste: 1t ore. Error! Reference source not found. shows a production profile. Figure 1-3 shows the expected Ni and Cu feed grades and payable outputs.


DRAFT REPORT 5

Production Profile

-

5,000,000

10,000,000

15,000,000

20,000,000

25,000,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

ag

e

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00


Figure 1-2 Production profile


-

2,000

4,000

6,000

8,000

10,000

12,000

14,000

16,000

18,000

20,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

es

0.000%

0.100%

0.200%

0.300%

0.400%

0.500%

0.600%

CU NI NI% CU%

Figure 1-3 Expected Nickel and Copper Payable Output.


DRAFT REPORT 6

1. INTRODUCTION .............................................................................................................. 8

2. INPUT PARAMETERS..................................................................................................... 9

2.1 MODEL GENERATION .................................................................................................. 92.2 WASTE BLOCKS.......................................................................................................... 92.3 ELEMENT GRADE........................................................................................................ 92.4 DENSITY .................................................................................................................... 92.5 TONNAGES................................................................................................................. 92.6 MINING COST AND MINING COST ADJUSTMENT FACTOR (MCAF) ............................... 102.7 RECOVERIES ............................................................................................................ 112.8 RECOVERED TONNES ............................................................................................... 12

3. WHITTLE PARAMETERS.............................................................................................. 13

3.1 SLOPE SET .............................................................................................................. 133.2 PIT SHELLS .............................................................................................................. 13

Mining.............................................................................................................................. 13Processing ...................................................................................................................... 13Selling Costs ................................................................................................................... 13Metal Prices .................................................................................................................... 14

3.3 TIME COSTS............................................................................................................. 14Discount Rate ................................................................................................................. 14Mining Limits ................................................................................................................... 14

3.4 OPERATIONAL SCENARIOS ........................................................................................ 14

4. WHITTLE RESULTS ...................................................................................................... 16

5. DESIGN CRITERIA ........................................................................................................ 17

5.1 OPEN PIT DIMENSIONS ............................................................................................. 175.2 PIT WALL PARAMETERS ............................................................................................ 175.3 RAMP SYSTEM ......................................................................................................... 175.4 DESIGN ANALYSIS .................................................................................................... 18

6. MINEABLE RESERVE ESTIMATION............................................................................ 21

7. MINE SCHEDULING ...................................................................................................... 22

7.1 PIT SHELL ILLUSTRATIONS ........................................................................................ 297.1.1 Final Year 19 Pit Shell Illustration.......................................................................... 29

8. RECOMMENDATIONS .................................................................................................. 31

List of figures

FIGURE 1-1: FINAL PIT DESIGN ..................................................................................................... 3FIGURE 1-2 PRODUCTION PROFILE................................................................................................ 5FIGURE 2-1: MCAF IN RELATION TO AN INCREASE IN MINING DEPTH. ........................................... 10FIGURE 4-1: WHITTLE OPTIMISATION RESULTS – PIT BY PIT GRAPH............................................. 16FIGURE 5-1: PIT DESIGN ............................................................................................................ 18FIGURE 5-2 EAST TO WEST CROSS SECTION OF THE PIT DESIGN VS. THE WHITTLE PIT SHELL 48. 19FIGURE 5-3 NORTH TO SOUTH CROSS SECTION OF THE DESIGN VS. THE WHITTLE SHELL 48. ....... 20FIGURE 5-4 3D VIEW OF THE PIT DESIGN (BROWN) AND THE WHITTLE SHELL 48 (BLUE).............. 20FIGURE 7-1: TONNAGE PIT PROFILE AT THE VARIOUS CUT-OFFS................................................... 23FIGURE 7-2: PIT VALUE VERSUS KEY ELEMENT GRADES.............................................................. 23FIGURE 7-3 NICKEL AND COPPER OUPUT AND FEED GRADES ...................................................... 24FIGURE 7-4 PGM OUTPUTS AND FEED GRADES.......................................................................... 24FIGURE 7-5: PRODUCTION PROFILE AT OPTIMAL CUTOFF. ............................................................ 25FIGURE 7-6: NICKEL AND COPPER PLANT OUTPUT AND FEED GRADES......................................... 26FIGURE 7-7: PGE PLANT OUTPUT AND FEED GRADES................................................................. 26FIGURE 7-8: EXPECTED CASHFLOW............................................................................................ 27FIGURE 7-18: PLAN VIEW FOR PIT SHELL IN FINAL YEAR 16. ........................................................ 29FIGURE 7-19: PLAN VIEW FOR PIT SHELLS IN FINAL YEAR 16........................................................ 30


DRAFT REPORT 7

FIGURE 7-20: BOTTOM 3D VIEW FOR PIT SHELLS IN FINAL YEAR 16. ............................................ 30

List of Tables

TABLE 1-1: METAL PRICES ........................................................................................................... 2TABLE 1-2: DESIGN VS. WHITTLE RESULTS.................................................................................... 3TABLE 1-3: MINERAL RESERVE..................................................................................................... 4TABLE 2-1: MODEL FIELDS ........................................................................................................... 9TABLE 3-1: ELEMENT SELLING COSTS ........................................................................................ 13TABLE 3-2: METAL PRICE FORECASTS ........................................................................................ 14TABLE 5-1: COMPARISON OF THE WHITTLE GENERATED PIT SHELL TO THE DESIGN ........................ 18TABLE 6-1: PROVEN MINERAL RESERVE – ALL METALS AT NI CUT-OFF ........................................ 21TABLE 6-2: PROBABLE MINERAL RESERVE– ALL METALS AT NI CUT-OFF ..................................... 21TABLE 6-3: TOTAL MINERAL RESERVE– ALL METALS AT NI CUT-OFF............................................ 21TABLE 7-1: KEY OUTPUTS COMPARISON AT VARIOUS NI% CUT-OFFS. ......................................... 22TABLE 7-2: SCHEDULE RESULTS AT A CUT-OFF OF 0.18% NICKEL (IN-SITU OR UNDILUTED) .......... 25TABLE 7-3 - LIFE OF MINE SCHEDULE AT 0.18% CUT-OFF ............................................................. 28

DRAFT REPORT 8

1. Introduction

This document serves as a record of the optimisation, design, and scheduling parameters used for the Keivitsa Ni, Cu, Co, Pt, Pd & Au Deposit for Scandinavian Gold. A detailed geological model was provided and adapted and with the use of a strategic planning software package an open pit design was created. This design was used to assess the mineable reserves of the project and also to provide a detailed life of mine schedule.

DRAFT REPORT 9

2. Input parameters

The following input parameters were used and adapted to enable the use of the strategic mine planning software.

2.1 Model Generation

An ore model constructed by Dexter Ferreira (Final.Model) was used in this exercise. This model contained the following fields.

Table 2-1: Model Fields Element Quantity Co Percent Cu Percent Ni Percent S Percent Au Grams/TonnePd Grams/TonnePt Grams/Tonne

In order to enable this model to be used in whittle the following steps were taken.

2.2 Waste Blocks

Since the model received contained ore blocks only, provisions were made for the waste that must be mined to access the ore. The waste blocks were created using the (Final.Model) as a prototype model. The ore model and waste models were then combined before it was imported into the Whittle Pit optimiser.

2.3 Element Grade

The different element grades were read from the respective fields. The respective grade fields within the Datamine model were in the appropriate format i.e., Elements Nickel, Copper, Cobalt & Sulphur were represented in percentages and the PGM & Au elements were represented in Grams per Tonne. Additional grade fields were developed in order to take into account the dilution factor of 3% for the appropriate feed grade when used in Scheduling.

2.4 Density

A waste and ore density of 3.15 t/m3 was used in the model to obtain the respective masses

in the ore-body model.

2.5 Tonnages

From the applied density the ore block tonnages were calculated from the following ore and waste block dimensions:

XINC = 10m YINC = 10m ZINC = 10m

ORETONS = (XINC*YINC*ZINC*DENSITY)

These blocks also formed the Smallest Mining Unit (SMU) for the optimisation and scheduling process. This was then used to calculate the following element tonnes and grams.

DRAFT REPORT 10

NI TONS = (ORETONS*(NI/100)) CU TONS = (ORETONS*(CU/100)) CO TONS = (ORETONS*(CO/100)) S TONS = (ORETONS*(S/100)) PT GRAM = (ORETONS*PT) PD GRAM = (ORETONS*PD) AU GRAM = (ORETONS*AU)

The in-situ metal masses were then calculated for the specified rock type codes. The specified rock type codes are:

Rock 1 = Waste Rock Rock 13 = Inferred Material Rock 14 = Indicated Material Rock 15 = Measured Material

This enabled the reporting and optimisation of the model using the relevant categories originally stated in the (Final.Model) ore-body model file.

2.6 Mining Cost and Mining Cost Adjustment Factor (MCAF)

A MCAF field (Mining Cost Adjustment Factor) was applied to the model. This is used to enable Whittle to calculate the mining cost accurately with increasing depth. A mining cost of $1.50/t (1.24 Euro/t) was used at a MCAF of 1, however as the pit progressively gets deeper the MCAF increases and thus the mining cost increases in relation to the MCAF. An equation used to determine the MCAF from a similar operation was applied. Because the contractor quotations reflected constant mining cost for the first 30m below surface an MCAF of 1 was applied for these blocks. An average surface elevation of 240m (ZC) was estimated. This was then used to provide the following logical statement.

IF (ZC>210) MCAF=1 IF (ZC<=210) MCAF= ((-0.0015*ZC) + 1.3276)

MCAF vs Pit Depth

y = -0.015x + 1.3276

R2 = 1

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

-1200 -1050 -900 -750 -600 -450 -300 -150 0 150 300

mamsl (m)

MC

AF

($

/t)

Figure 2-1: MCAF in relation to an increase in Mining Depth.

DRAFT REPORT 11

2.7 Recoveries

The following recovery equations were derived from the Keivitsa financial spreadsheet provided by Saint Barbara Consultancy Services. These figures represent overall recovery to saleable metal in both concentrates combined.

Ni Recovery = (((79.92+1.49+11.766*LN(Ni%))/100)*0.67) Cu Recovery = (57.9-0.629*(79.92+1.49+11.766*LN(Ni%)))/100*0.76+ (60/100*(24.5-1)/24.5) Co Recovery = (((0.98*(79.92+1.49+11.766*LN(Ni%)))/100)*0.3) Au Recovery = (((0.693*(79.92+1.49+11.766*LN(Ni%))-27.86)/100)*0.36)+ (0.21*(AUGT*0.21*(0.245*100)/(Cu%*0.6)-1)/(AUGT*0.21* (0.245*100)/(Cu%*0.6))) Pt Recovery = (((1.325*(79.92+1.49+11.766*LN(Ni%))-52.402)/100)*0.45) Pd Recovery = (((1.023*(79.92+1.49+11.766*LN(Ni%))-34.31)/100)*0.45)

Because the grades in the model are very low in some areas the equations given produced negative values for the Platinum, Palladium and Gold recoveries in certain situations. Due to this an evaluation was conducted to determine the point at which the recoveries turned negative. A logical statement was then used within Datamine to eliminate the negative recoveries. These logical statements can be seen below.

if (NI%<0.0022) NIREC = 0.0634 elseNIREC = (((81.41+(11.766*LN(NI%)))/100)*0.67)

if (NI%<0.0022) CUREC = 0.9706 elseCUREC = (((57.9-(0.629*(81.41+11.766*LN(NI%))))/100)*0.76)+(0.5755)

if (NI%<0.0022) COREC = 0.0277 elseCOREC = (((0.98*(81.41+(11.766*LN(NI%))))/100)*0.3)

if (NI%<0.0286) PTREC = 0.0002 elsePTREC = (((1.325*(81.41+11.766*LN(NI%))-52.402)/100)*0.45)

if (NI%<0.0176) PDREC = 0.0016 elsePDREC = (((1.023*(81.41+11.766*LN(NI%))-34.31)/100)*0.45)

if (NI%<0.0022) AUREC = 0.0597 elseA = (0.693*(81.41+11.766*Loge(NI%))-27.86) B = (0.21)*(AUGT*0.21*(24.5)/(CU%*0.6)-1) / (AUGT*0.21*(24.5)/(CU%*0.6)) AUREC = A/100*0.36+B

Due to the number of variables contained in the equation the AUREC still provided negative results. On analysis of the model less than 1% of the Au recoveries were negative. It was decided to make all of the negative recoveries zero.

DRAFT REPORT 12

2.8 Recovered Tonnes

Due to the complex nature of the recovery equations the recoveries had to be calculated in Datamine. Since the recoveries were calculated in Datamine and the new block model imported into Datamine contained the recovered values, 100% recovery was applied in Whittle.

The recovered tonnes were calculated by the following equations.

Recovered Ni Metal = NITONS*NIREC Recovered CU Metal = CUTONS*CUREC Recovered CO Metal = COTONS*COREC Recovered PT Metal = PTGRAM*PTREC Recovered PD Metal = PDGRAM*PDREC Recovered AU Metal = AUGRAM*AUREC

DRAFT REPORT 13

3. Whittle Parameters

3.1 Slope Set

A wall angle of 45° was applied to the top 30m of the open pit. This slope angle is affected from the +210m elevation upward to the highest elevation of +260m. The average surface elevation was determined to be approximately 240m.

The maximum allowable slope angle for the unweathered material is 55°. SRK identified a maximum inter-ramp angle of 65° in the preliminary geo-technical report. A single 25m roadway including a 3m tyre berm will be adequate for bi-directional traffic for mining trucks up to CAT 785 size, this is deemed adequate for the expected pit tonnages.

Benches of 10m and 5m berm widths produced an inter-ramp angle of 63.4°. The overall wall angle for the un-weathered material below the +210m level is 55°. From the -90m elevation to the pit floor (-170m) a 15m width roadway was used in order to access the ore at these lowest elevations. Practically this is achievable since the final pushback has been selected to push the final pit from these elevations down and out to the final pit shell.

The overall pit wall angle is 53.8°. Batter angles on the bench faces were designed at 90°, practically the batter angle will be approximately 85°.

3.2 Pit Shells

Mining

Within the Whittle optimisation a mining cost of $1.50/tonne was used for the reference block. As previously stated a MCAF was applied into the model. A Mining Recovery Fraction of 97% and a Mining Dilution Factor of 3% were also applied.

Processing

As the model used was a multi-element model, no specific cut-off was applied. Rather, the cash flow ore selection method was used in Whittle. Whittle calculates a block value for each individual block. If the block value is positive is is selected for processing. In this way an economic cut-off can be determined. The economic cut-off from the Whittle optimisation was computed using back-analysis and was determined to be approximately 0.113% Nickel Feed Grade.

A processing cost of $7.15/tonne of ore processed (including fixed costs).

Selling Costs

The selling costs used in the Whittle optimisation can be seen below.

Table 3-1: Element Selling Costs Element Sell Cost ($) Units Ni 429.10 TonneCu 208.62 TonneCo 0 TonnePt 0 Gram

DRAFT REPORT 14

Pd 0 Gram Au 0 Gram

The selling costs were attributed to the Nickel and the Copper concentrate stream respectively. A portion of Copper concentrate attributable selling costs was discounted from the Copper selling cost since approximately 20% of the total copper produced is in the Nickel concentrate stream.

Metal Prices

Metal prices were provided in the financial spreadsheet supplied to CSA. Table 3.2 below outlines the metal prices that were used.

Table 3-2: Metal Price Forecasts Element Price Forecast ($) Units

Ni 10 319.00 $/Tonne Cu 2 668.00 $/Tonne Co 33 069.35 $/Tonne Pt 16.0750 $/Gram Pd 6.4301 $/Gram Au 14.4680 $/Gram

3.3 Time Costs

No time costs were allowed for. All fixed costs were converted to a processing cost per tonne milled. No capital costs were applied during the Whittle optimisation. This will not affect the size of the optimal pit. During the Feasibility Study the effect of varying ongoing capital and time costs should be analysed.

Discount Rate

It was agreed that a Discount Rate per Period of 7% was to be used in the Whittle optimisation. This is due to the low level of risk associated with Finland. This rate was constant for the life of the project.

Mining Limits

A maximum Plant Processing capacity was set at 4.5 million tonnes of ore per annum for the life of the project. The ore processing limit was not adjusted to accommodate for ramp-up, instead a waste pre-strip of approximately 10 million tonnes was included in order to account for possible delivery and ramp-up of the mining fleet capacity. The tail end of the production profile is extremely short and only covers one period. The practicality of this must be investigated during the feasibility study.

The mining capacity requirements decrease drastically in year 8 from approximately 21 million tonnes per annum to approximately 9 million tonnes per annum. Thus, a reduced mining capacity would need to be effected in this period. This is because of the reduced stripping ratio at the lower elevations of the pit.

3.4 Operational Scenarios

For the Whittle optimisation 6 push-backs including the ultimate optimal pit were set at intervals of approximately 10.0 million tonnes of ore per push-back. A fixed lead of 10

DRAFT REPORT 15

benches was used for the schedule, this was determined to be the minimum mining width required for the trucks and shovels to operate effectively. In order for the specified case to represent a practical mining scenario a bench lead of 10 was used. The lead forces the current pushback that is being mined to mine 10 benches (in this case 100m vertical depth and 75m horizontal distance) before the following pushback may begin mining. This ensures that equipment and personal have sufficient space and reduced constraints to mine the two pushbacks simultaneously without constraining production.

DRAFT REPORT 16

4. Whittle Results

Figure 4.1 illustrates the Pit by Pit graph output from the Whittle optimisation based on the above-mentioned input parameters.

Figure 4-1: Whittle Optimisation Results – Pit by Pit Graph.

From the input parameters and Revenue Factors of 0.3 to 1.5, Whittle was used to generate 101 pit shells. On analysis of these results Pit 48 was selected as the optimal, giving the maximum project NPV at the minimum total tonnage moved for the specified case scenario and the applied pushbacks.

Pit 48 represents a flex point in the scheduled case NPV graph. Beyond this point the NPV curve is relatively flat. This means that the pit size increases without adding NPV to the overall project. Choosing a pit bigger than pit 48 would represent increased risk for the project. If any of the economic parameters changed for the worse while the final pushback is being stripped it could put the mine into a loss making position.

DRAFT REPORT 17

5. Design Criteria

5.1 Open Pit Dimensions

A design was generated from the chosen Whittle optimal pit which is roughly oval. The pit’s longest axis runs roughly Northwest/Southeast and has a length of 900m. On its Northeast/SouthWest axis the design produced a length of 700m. At its deepest point the pit is 330m below surface.

5.2 Pit Wall parameters

The pit was designed with 10m berms for the first 30m reducing to 5m berms for the remainder of the pit. For the first 3 benches a batter angle of 90° was used. Each bench was set to a height of 10m in line with the block model dimensions. This resulted in an overall wall angle of 45° for the top 30m. The remainder of the pit was designed with a batter angle of 90° and an inter ramp wall angle of 63.4°. With the inclusion of the ramp width the pit wall at its steepest point has an angle of 53.8°. The pit wall dimensions were chosen in line with the 2003 SRK report as no further geotechnical information is available.

Further geotechnical investigations will be carried out as part of the Feasibility Study to enable optimisation of the overall pit angles.

5.3 Ramp System

A single 25m wide ramp system was incorporated into the top 320m of the design to enable efficient ore extraction and provide flexibility. At 320m below the surface elevation (-90 mamsl pit elevation) the ramp access was designed to include a single ramp 15m wide to the full depth of the pit. This was done primarily as a method to increase the inclusion of ore at depth thus increasing the slope angle for the deeper parts of the pit and to reduce the waste mining required at the shallower depths of the pit. The single ramp width of 25m allows for a greater wall angle using the same stack heights, catch berm widths, batter angles and step offs. This has the effect of increasing the inter-ramp angle allowing the pit design to “hug” the optimal Whittle design more effectively.

DRAFT REPORT 18

Figure 5-1: Pit Design

5.4 Design Analysis

A comparison of the designed pit was conducted against the Whittle pit shell in Datamine. The results are shown in Table 5-1.

Table 5-1: Comparison of the Whittle generated pit shell to the design

Optimised Base Case Ore selected for Processing

Proved Probable Total Mineral Reserve

Whittle Design Variance

(%) Whittle Design Variance

(%) Whittle DesignVariance

(%)

Total Tonnage (Mt) 223.1 223.9 0.36% 223.1 223.9 0.36% 223.1 223.9 0.36%

Ore Tonnage (Mt) 87.4 87.5 0.04% 17.0 16.8 -1.53% 104.5 104.3 -0.21%

Payable Recovered Metal Payable Recovered Metal Payable Recovered Metal

Ni (t) 94,099 93,144 -1.01% 19,760 17,834 -9.75% 113,859 110,978 -2.53%

Cu (t) 208,379 207,088 -0.62% 50,675 46,421 -8.39% 259,054 253,509 -2.14%

Co (t) 2,129 2,117 -0.56% 466 428 -8.15% 2,595 2,545 -1.93%

Pt (g) 3,457,541 3,434,581 -0.66% 632,143 570,396 -9.77% 4,089,684 4,004,977 -2.07%

Pd (g) 2,102,927 2,088,607 -0.68% 377,638 343,445 -9.05% 2,480,565 2,432,052 -1.96%

Au (g) 1,991,105 2,000,430 0.47% 396,689 365,165 -7.95% 2,387,794 2,365,595 -0.93%

Feed Grade (Diluted In-Situ) Feed Grade (Diluted In-Situ) Feed Grade (Diluted In-Situ)

Ni (%) 0.243% 0.241% -0.78% 0.245% 0.241% -1.55% 0.244% 0.242% -0.96%

Cu (%) 0.335% 0.332% -0.92% 0.400% 0.395% -1.38% 0.350% 0.346% -1.09%

Co (%) 0.013% 0.013% -0.78% 0.014% 0.013% -0.74% 0.013% 0.013% -0.42%

Pt (g/t) 0.255 0.253 -0.78% 0.222 0.218 -1.53% 0.246 0.244 -0.92%

Pd (g/t) 0.162 0.160 -0.93% 0.140 0.138 -1.36% 0.157 0.155 -1.02%

Au (g/t) 0.113 0.113 -0.53% 0.116 0.115 -0.95% 0.115 0.114 -0.65%

Metal Contents (Diluted In-Situ) Metal Contents (Diluted In-Situ) Metal Contents (Diluted In-Situ)

DRAFT REPORT 19

Ni (t) 213,186 211,487 -0.80% 41,793 40,509 -3.07% 254,979 251,996 -1.17%

Cu (t) 297,590 294,802 -0.94% 68,209 66,244 -2.88% 365,799 361,046 -1.30%

Co (t) 11,192 11,154 -0.34% 2,296 2,248 -2.09% 13,488 13,402 -0.64%

Pt (g) 21,942,010 21,765,966 -0.80% 3,778,219 3,664,015 -3.02% 25,720,229 25,429,981 -1.13%

Pd (g) 13,985,754 13,854,075 -0.94% 2,390,009 2,320,583 -2.90% 16,375,763 16,174,658 -1.23%

Au (g) 9,991,696 9,937,960 -0.54% 1,972,686 1,923,794 -2.48% 11,964,382 11,861,754 -0.86%

Figures 5-2 to 5-4 below show cross sections of the conceptual pit design wireframe against the optimal pit shell from the Whittle optimisation (Pit 48). A percentage of the optimal Whittle pit may not be mined using a 25m wide roadway due to practical mining considerations such as minimum mining width and equipment size and the minimum space required for equipment to operate. For this reason the roadway below the -90 mamsl was adjusted to 15m wide in order to ensure optimal extraction of the Whittle generated pit shell. The design is deemed acceptable since the variances between the Whittle pit shell and the design are within 5%.

Figure 5-2 East to West Cross Section of the Pit Design vs. the Whittle Pit Shell 48.

DRAFT REPORT 20

Figure 5-3 North to South Cross Section of the Design vs. the Whittle Shell 48.

Figure 5-4 3D View of the Pit Design (Brown) and the Whittle Shell 48 (Blue).

DRAFT REPORT 21

6. Mineable Reserve Estimation

Evaluating the pit design against the geological model at various cut offs gave the following results. Table 6-1: Proven Mineral Reserve – All Metals at Ni Cut-off

Proved Mineral Reserve

Ni Cut-Off (%) Percent Feed Grade (g/t) Feed Grade

Feed Grade (Mt) Ni Cu Co S Pt Pd Au

0.05% 144.1 0.177% 0.225% 0.011% 1.023% 0.177 0.106 0.076

0.10% 95.8 0.229% 0.307% 0.012% 1.308% 0.235 0.148 0.104

0.15% 67.9 0.273% 0.379% 0.013% 1.561% 0.285 0.183 0.129

0.16% 63.8 0.280% 0.339% 0.014% 1.601% 0.294 0.189 0.133

0.17% 60 0.288% 0.403% 0.014% 1.642% 0.302 0.195 0.137

0.18% 56.2 0.295% 0.415% 0.014% 1.683% 0.310 0.201 0.141

0.19% 52.9 0.302% 0.425% 0.014% 1.717% 0.317 0.206 0.144

0.20% 49.7 0.309% 0.434% 0.014% 1.748% 0.325 0.211 0.148

0.25% 34.8 0.345% 0.482% 0.015% 1.907% 0.367 0.240 0.167

Table 6-2: Probable Mineral Reserve– All Metals at Ni Cut-off

Probable Mineral Reserve



0.05% 27.2 0.175% 0.266% 0.011% 1.189% 0.154 0.094 0.077

0.10% 17.4 0.232% 0.373% 0.013% 1.519% 0.209 0.132 0.109

0.15% 12.4 0.277% 0.459% 0.014% 1.785% 0.252 0.161 0.133

0.16% 11.8 0.283% 0.470% 0.015% 1.823% 0.257 0.164 0.136

0.17% 11.2 0.289% 0.480% 0.015% 1.858% 0.261 0.167 0.139

0.18% 10.6 0.295% 0.492% 0.015% 1.896% 0.267 0.171 0.142

0.19% 9.9 0.303% 0.506% 0.015% 1.935% 0.273 0.175 0.146

0.20% 9.2 0.311% 0.518% 0.015% 1.967% 0.28 0.179 0.15

0.25% 6.7 0.344% 0.575% 0.017% 2.128% 0.311 0.196 0.168

Table 6-3: Total Mineral Reserve– All Metals at Ni Cut-off

Total Mineral Reserve



0.05% 171.3 0.177% 0.232% 0.011% 1.049% 0.173 0.104 0.076

0.10% 113.2 0.229% 0.317% 0.012% 1.340% 0.231 0.146 0.105

0.15% 80.3 0.274% 0.391% 0.013% 1.595% 0.280 0.180 0.130

0.16% 75.6 0.280% 0.359% 0.014% 1.636% 0.288 0.185 0.133

0.17% 71.2 0.288% 0.415% 0.014% 1.676% 0.296 0.191 0.137

0.18% 66.8 0.295% 0.427% 0.014% 1.717% 0.303 0.196 0.141

0.19% 62.8 0.302% 0.438% 0.014% 1.751% 0.310 0.201 0.144

0.20% 58.9 0.309% 0.447% 0.014% 1.782% 0.318 0.206 0.148

0.25% 41.5 0.345% 0.497% 0.015% 1.943% 0.358 0.233 0.167

DRAFT REPORT 22

7. Mine Scheduling

Mine scheduling was done using LQS scheduling software. In order to improve the project NPV a cut-off optimisation was performed. Several Ni% cut-off scenarios were scheduled to determine which cut-off would match the project goals. Table 7-1 below outlines the key outputs from these schedules.

Table 7-1: Key Outputs Comparison at Various Ni% Cut-offs.

Schedule Ni Cut-off 0.10% 0.15% 0.16% 0.17% 0.18% 0.19% 0.20% 0.25%

Ore Tonnes 112,079,151 79,424,356 74,867,014 70,578,209 66,127,336 62,218,894 58,456,217 41,264,143

Waste Tonnes 110,656,304 143,311,101 147,868,442 152,157,246 156,608,119 160,516,561 164,279,240 181,471,313

Strip ratio 0.99 1.80 1.98 2.16 2.37 2.58 2.81 4.40

Ni % 0.230% 0.274% 0.281% 0.288% 0.296% 0.302% 0.309% 0.345%

Cu % 0.316% 0.391% 0.403% 0.414% 0.426% 0.437% 0.447% 0.496%

Co % 0.012% 0.014% 0.014% 0.014% 0.014% 0.014% 0.015% 0.016%

Pt g/t 0.233 0.282 0.290 0.297 0.305 0.313 0.320 0.361

Pd g/t 0.146 0.181 0.187 0.192 0.197 0.202 0.207 0.235

Au g/t 0.105 0.130 0.134 0.138 0.142 0.145 0.149 0.167

Payable Nickel (t) 112,243 97,072 94,290 91,452 88,285 85,312 82,287 65,884

Payable Copper (t) 248,687 216,059 209,549 202,889 195,225 188,044 180,410 140,380

Payable Cobalt (t) 2,592 2,095 2,014 1,935 1,850 1,772 1,694 1,296

Payable Platinum (g) 4,021,928 3,647,643 3,565,389 3,478,026 3,375,934 3,280,262 3,182,152 2,631,480

Payable Palladium (g) 2,433,507 2,217,112 2,168,644 2,117,190 2,055,597 1,996,455 1,935,153 1,599,669

Payable Gold (g) 2,354,732 2,097,241 2,039,799 1,980,314 1,911,508 1,849,367 1,785,659 1,444,357

Life of Mine 25 18 17 16 15 14 13 10

NPV 298,872,640 379,535,104 385,792,960 387,305,312 385,640,000 382,237,568 375,975,360 309,877,344

After a few iterations a cut-off of 0.18% Ni was selected for the final schedule as it provided the required feed grade and life of mine required by the strategic business targets.

Figures 7-1 and 7-4 below illustrate how the various output parameters vary with increasing cut-off grade.

DRAFT REPORT 23

Cutoff Analysis - Production Volumes

-

50,000,000

100,000,000

150,000,000

200,000,000

250,000,000

0.10

%

0.15

%

0.16

%

0.17

%

0.18

%

0.19

%

0.20

%

0.25

%

Nickel Cut-off

Tonnes

0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

5.0

Str

ip R

atio

Ore Tonnes Waste Tonnes Strip ratio

Figure 7-1: Tonnage Pit Profile at the various cut-offs.

One can see how the increasing stripping ratio and consequent increase in costs outweigh the benefit of higher feed grades above a 0.18%Ni cut-off. Above a 0.20%Ni cut-off the size of the pit decreases considerably.

P i t Va l ue s - NP V

295,000,000

315,000,000

335,000,000

355,000,000

375,000,000

395,000,000

0.10% 0.15% 0.16% 0.17% 0.18% 0.19% 0.20% 0.25%

Ni ckel Cut - of f

60,000

70,000

80,000

90,000

100,000

110,000

120,000

NPV Payable Nickel (t) Poly. (NPV)

Figure 7-2: Pit Value versus Key Element Grades.

DRAFT REPORT 24

Nickel and Copper Output & Grade

-

50,000

100,000

150,000

200,000

250,000

0.10

%

0.15

%

0.16

%

0.17

%

0.18

%

0.19

%

0.20

%

0.25

%

Nickel Cut-off

Tonnes

0.200%

0.250%

0.300%

0.350%

0.400%

0.450%

0.500%

0.550%

%

Payable Copper (t) Payable Nickel (t)

Ni % Cu %

Figure 7-3 Nickel and Copper Ouput and Feed Grades

P GE Out put a nd Gr a de s

-

500,000

1,000,000

1,500,000

2,000,000

2,500,000

3,000,000

3,500,000

4,000,000

0.10% 0.15% 0.16% 0.17% 0.18% 0.19% 0.20% 0.25%

Ni c k e l Cut - of f

0.0

0.1

0.1

0.2

0.2

0.3

0.3

0.4

0.4

Payable Platinum (g) Payable Pal ladium (g) Payable Gold (g)

Pt g/ t Pd g/ t Au g/ t

Figure 7-4 PGM Outputs and Feed Grades

DRAFT REPORT 25

The result from the schedule can be seen in the table below. The grade shown is the feed grade to the plant.

Table 7-2: Schedule Results at a Cut-off of 0.18% Nickel (In-Situ or undiluted)

Schedule Ni Cut-off 0.18%

Ore Tonnes 66,127,336

Waste Tonnes 156,608,119

Strip ratio 2.37

Ni % 0.296%

Cu % 0.426%

Co % 0.014%

Pt g/t 0.305

Pd g/t 0.197

Au g/t 0.142

Payable Nickel (t) 88,285

Payable Copper (t) 195,225

Payable Cobalt (t) 1,850

Payable Platinum (g) 3,375,934

Payable Palladium (g) 2,055,597

Payable Gold (g) 1,911,508

Life of Mine 15

NPV 385,640,000

Figures 7-5 to 7-8 show the results from the Schedule ran at a Nickel Cut off of 0.18%

Production Profile

-

5,000,000

10,000,000

15,000,000

20,000,000

25,000,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

ag

e

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00


Figure 7-5: Production Profile at optimal cut-off.

.

DRAFT REPORT 26


-

2,000

4,000

6,000

8,000

10,000

12,000

14,000

16,000

18,000

20,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

To

nn

es

0.000%

0.100%

0.200%

0.300%

0.400%

0.500%

0.600%

CU NI NI% CU%

Figure 7-6: Nickel and Copper Plant Output and Feed Grades

Payable PGE Output

0

50000

100000

150000

200000

250000

300000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

Pt Pd Au PTGT PDGT AUGT

``

Figure 7-7: PGE Plant Output and Feed Grades

The production profile illustrated in Figure 7-5 above shows the ore and waste tonnages that are to be expected from the ultimate pit design. Waste stripping of 10 million tonnes was brought forward to begin in period 0. This was done to expose enough ore to allow a consistent feed to the plant. It also allows for blending. Period 8 shows a decrease in total tonnages required to be mined in order to achieve the plant feed tonnages.

DRAFT REPORT 27

Open Pit Cash-flow

-

10,000,000

20,000,000

30,000,000

40,000,000

50,000,000

60,000,000

70,000,000

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

Period

CashFlow

``

Figure 7-8: Expected Cashflow.

The expected open pit cashflow in Figure 7-8. The cashflow is increased in the later years due to the fact that the stripping ratio is drastically reduced for the final pushbacks.

For purposes of clarity the graph has been set to zero on the y-axis, however, due to the forced waste stripping in period zero, the cash-flow is negative to the order of $15 million.

The table below shows the life of mine schedule for the Keivitsa open pit at a 0.18% Ni cut-off.

DR

AF

T R

EP

OR

T

2

8

Tab

le 7

-3 -

Lif

e o

f m

ine s

ch

ed

ule

at

0.1

8%

cu

t-o

ff

Nic

ke

l F

ee

d G

rad

e

Cu

toff

:

0.1

8%

Pa

ya

ble

Me

tal

Re

co

ve

red

F

ee

d G

rad

e (

In-S

itu

Dil

ute

d)

Op

tim

ised

Ba

se

Ca

se

Pit S

he

ll -

4.5

Mtp

a P

roce

ssin

g L

imit

To

nn

es

Gra

ms

%G

ram

s/t

on

ne

Pe

rio

d C

as

hF

low

R

oc

k

Ore

Wa

ste

S

trip

R

ati

o

NI

CU

C

O

Pt

Pd

A

u

NI%

C

U%

C

O%

P

TG

T

PD

GT

A

UG

T

0

10

,00

0,0

00

10

,00

0,0

00

-

- -

- -

- -

19

,413

,04

8

18

,54

5,1

40

4

,500

,00

0

14

,04

5,1

40

3

.12

5

,151

9

,646

1

05

2

26

,846

1

43

,895

1

17

,778

0

.285

%

0.3

42

%

0.0

13

%

0.3

49

0

.231

0

.135

25

2,8

94

,145

1

3,4

67

,628

4

,500

,00

0

8,9

67

,62

8

1.9

9

6,6

50

1

1,2

17

1

25

3

01

,089

2

03

,817

1

57

,599

0

.320

%

0.3

59

%

0.0

14

%

0.3

71

0

.268

0

.156

32

5,5

43

,982

2

0,8

84

,570

4

,500

,00

0

16

,38

4,5

70

3

.64

5

,515

1

0,3

69

1

18

2

46

,012

1

66

,375

1

36

,736

0

.272

%

0.3

27

%

0.0

13

%

0.3

25

0

.231

0

.138

43

4,1

62

,106

2

0,9

16

,724

4

,500

,00

0

16

,41

6,7

24

3

.65

6

,063

1

1,8

73

1

24

2

33

,031

1

43

,580

1

35

,860

0

.296

%

0.3

79

%

0.0

14

%

0.3

09

0

.201

0

.142

53

1,5

57

,027

2

0,9

42

,502

4

,500

,00

0

16

,44

2,5

02

3

.65

5

,919

1

1,8

73

1

22

2

37

,281

1

44

,034

1

33

,401

0

.290

%

0.3

78

%

0.0

14

%

0.3

14

0

.203

0

.140

63

7,2

61

,024

2

1,0

12

,441

4

,500

,00

0

16

,51

2,4

41

3

.67

5

,741

1

4,9

22

1

26

2

42

,200

1

40

,414

1

55

,905

0

.283

%

0.4

75

%

0.0

14

%

0.3

26

0

.200

0

.167

73

3,3

68

,186

2

0,7

67

,327

4

,500

,00

0

16

,26

7,3

27

3

.61

5

,956

1

2,6

99

1

24

2

58

,000

1

49

,296

1

37

,967

0

.291

%

0.4

05

%

0.0

14

%

0.3

42

0

.210

0

.145

82

7,1

05

,822

2

1,0

28

,409

4

,500

,00

0

16

,52

8,4

09

3

.67

6

,028

1

1,0

82

1

24

2

30

,488

1

30

,397

1

21

,021

0

.294

%

0.3

52

%

0.0

14

%

0.3

03

0

.182

0

.127

95

8,8

19

,836

9

,144

,48

8

4,5

00

,00

0

4,6

44

,48

8

1.0

3

6,1

48

1

4,5

65

1

30

2

50

,004

1

39

,535

1

50

,412

0

.300

%

0.4

66

%

0.0

15

%

0.3

30

0

.196

0

.160

10

59

,68

3,8

69

9

,269

,15

6

4,5

00

,00

0

4,7

69

,15

6

1.0

6

6,1

66

1

5,0

22

1

27

2

40

,920

1

38

,629

1

45

,337

0

.302

%

0.4

82

%

0.0

14

%

0.3

19

0

.196

0

.158

11

55

,39

1,4

16

9

,169

,77

9

4,5

00

,00

0

4,6

69

,77

9

1.0

4

6,0

31

1

4,3

70

1

24

2

17

,513

1

31

,918

1

30

,127

0

.294

%

0.4

55

%

0.0

14

%

0.2

88

0

.186

0

.143

12

61

,76

7,9

60

9

,082

,63

8

4,5

00

,00

0

4,5

82

,63

8

1.0

2

6,3

83

1

5,5

93

1

30

2

26

,072

1

39

,129

1

38

,499

0

.306

%

0.4

94

%

0.0

14

%

0.2

92

0

.192

0

.151

13

50

,87

1,7

07

9

,062

,19

5

4,5

00

,00

0

4,5

62

,19

5

1.0

1

5,7

74

1

4,4

49

1

24

1

71

,026

1

06

,165

9

8,7

03

0

.285

%

0.4

62

%

0.0

14

%

0.2

32

0

.153

0

.120

14

61

,84

5,1

17

5

,926

,50

2

4,5

00

,00

0

1,4

26

,50

2

0.3

2

5,8

82

1

6,0

39

1

39

1

63

,726

1

02

,624

8

8,3

53

0

.290

%

0.5

13

%

0.0

16

%

0.2

20

0

.147

0

.116

15

53

,35

6,3

79

3

,515

,95

6

3,1

27

,33

6

38

8,6

20

0

.12

4

,878

1

1,5

06

1

08

1

31

,726

7

5,7

89

6

3,8

10

0

.338

%

0.5

37

%

0.0

17

%

0.2

38

0

.148

0

.119

16

17

To

tal

65

3,0

41

,62

4

22

2,7

35

,45

5

66

,12

7,3

36

1

56

,608

,11

9

2.3

7

88

,28

5

19

5,2

25

1

,850

3

,375

,93

4

2,0

55

,59

7

1,9

11

,50

8

0.2

96

%

0.4

26

%

0.0

14

%

0.3

05

0

.197

0

.142

DRAFT REPORT 29

7.1 Pit Shell Illustrations

Pit shell illustrations are given for years ended 15. Bench details were extracted from the results of the schedule at the optimum cut-off of 0.18%Ni. The pit shells exported from the Whittle optimisations were then imported into Datamine and converted into a wireframe.

7.1.1 Final Year 19 Pit Shell Illustration

Figures 7-18 through 7-20 illustrate the pit shell at the end of period 16.

Figure 7-18: Plan View for Pit Shell in final year 16.

DRAFT REPORT 30

Figure 7-19: Plan View for Pit Shells in final year 16.

Figure 7-20: Bottom 3D View for Pit Shells in final year 16.

DRAFT REPORT 31

8. Recommendations

The following recommendations are made for the next phase of the project:

Updating the geological model with new geological information Optimise throughput Detailed equipment selection techniques for the selected production profile. Pit re-optimised based on additional geotechnical, grade, cost or revenue parameters

based on any changes to these parameters. Sensitivity Analysis on key variables to examine sensitivity of the ultimate optimal pit. Optimise ultimate pit design based on new information sought in point 3 above, with

specific attention paid to geo-technical parameters and the possibility of steepening pit walls.

Detailed manual design of the selected push-backs. Detailed cut-off optimisation Detailed mining schedule

Appendix 5A –

Ore Types at Kevitsa

Ore types at Keivitsa

The question about ore types at Kevitsa have been addressed by Mutanen and Lemberg. Based on

various element ratios, different ore types can be recognised, and Dexter Ferreira used the

segregation proposed by Lemberg to discriminate between ore types.

Economic grades originate from two ore types, respectively named main ore type and nickel-PGE

ore type. The distinct difference is that nickel-PGE type has a higher Ni/Co ration than the Main

type, and in general nickel-PGE type has lower sulphur and copper contents than the main type. The

main ore type make up 93 % of all ore blocks with more than 0,2 % nickel to a depth of 330 meter

(see Figure 1). Mineralogical the two ore types have the same gangue minerals and the same

sulphide minerals – pyrrhotite, pentlandite and chalcopyrite. With lowering sulphur content the

pyrrhotite content diminishes, pentlandite obtains a higher nickel content and nickel sulphides

without iron (millerite and heazlewoodite) start forming.

In samples with deficient sulphur (there is not enough sulphur available to form pentlandite and

chalcopyrite), nickel in addition to forming pure nickel sulphides also starts building into silicates.

Based on the chemical composition of pentlandite and chalcopyrite deficient sulphur can be

calculated as weight per cent sulphur less the sum of weight per cent nickel plus copper.

The relation has been investigated by comparing analyses after total extraction and partial

extraction (only nickel and copper hosted in sulphides are taken into solution). These tests show that

– uncorrelated with sulphur content - 150 to 300 ppm nickel is hosted in silicate minerals, and for

samples with deficient sulphur the amount goes up. Based on nickel, copper and sulphur content, it

is possible in sulphur deficient samples to calculate a nickel number, which can be treated in the

same way as analytical values from sulphur surplus areas with respect to attributing a recovery

factor.

In figure 1 all measured, indicated and inferred ore blocks with a content of more than 0,2 % nickel

and to a depth of 330 meter are plotted. The plot is made to illustrate ore type and sulphur

deficiency with the nickel/cobalt ratio plotted as a function of sulphur less copper plus nickel. The

distribution of the two mentioned ore types is shown. There is a gradual transition between the two

ore types.

Metallurgical testing on nickel-PGE ore has demonstrated that high grade nickel concentrate can be

produced directly. In sulphur deficient blocks the recovery is lower, but these only make up 15 % of

the nickel-PGE type and 1,3 % of all ore blocks. Based on the proportion of individual drill holes

intersections (block values are smoothed and tend to down grade the significance of sulphur

deficiency) it can be estimated that total amount of recovered nickel will be 1 to 2 % less than

otherwise expected, as long as correction for nickel from sulphur deficient samples not has been

done.

The procedure for production of separate copper and nickel concentrate developed for the main ore

type has not yet been tested on the nickel-PGE type.

0

20

40

60

80

100

120

140

160

-0,5 0 0,5 1 1,5 2 2,5

All measured, indicated and inferred ore blocks

with >0,2 % Ni to a depth of 330 meter

S-(Ni+Cu)

Ni/Co

ti

Main ore type

Ni-PGE ore type

Figure 1: Nickel/cobalt ratio as a function of sulphur deficiency/surplus. The diagram illustrates

the two ore types: main ore and nickel-PGE ore.

Appendix 5B –

Metallurgy Samples, May 2005

Organisation Type of work Phase Objective Ore type HOLE-ID FROM TO ORIGGEOL

GTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 24 26 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 26 28 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 28 30 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 30 32 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 32 34 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 34 36 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 36 38 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 38 40 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 40 42 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 42 44 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 44 46 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 46 48 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 48 50 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 50 52 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 52 54 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 54 56 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 56 58 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 58 60 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 60 62 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 62 64 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 64 66 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 66 68 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 68 70 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 70 72 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 72 74 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 74 76 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 76 78 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 78 80 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-10 80 82 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-11 6 8 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-11 8 10 MetaperidotiteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-9 74 76 Olivine pyroxeniteGTK-Mintech Bench scale Phase 1 Bulk_con Ni-PGE KV-9 76 78 Olivine pyroxeniteGTK-Mintech Bench scale Phase 2 Bulk_con Cu-Ni K_vitsa 34 59.9GTK-Mintech Bench scale Phase 2 Bulk_con Cu-Ni K_vitsa 15 29.2GTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 31 33 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 33 35 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 35 37 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 37 39 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 39 41 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 41 43 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 43 45 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 45 47 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 47 49 MetaperidotiteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 49 51 MetaperidotiteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 51 53 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 53 55 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 55 57 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 57 59 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 59 61 Olivine pyroxenite

GTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 61 63 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 63 65 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 65 67 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 67 69 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 69 71 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 71 73 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 73 75 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 75 77 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 77 79 Olivine pyroxeniteGTK-Mintech Bench scale Phase 3 Sep_con Cu-Ni KV-15 79 81 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 7 9 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 9 11 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 11 13 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 13 15 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 15 17 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 17 19 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 19 21 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 21 23 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 23 25 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 25 27 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 27 29 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 29 31 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 81 83 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 83 85 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 85 87 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 87 89 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 89 91 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 91 93 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 93 95 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 95 97 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 97 99 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 99 101 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 101 103 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 103 105 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 105 107 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 107 109 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 109 111 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 111 113 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-15 113 115 Olivine pyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 128 130 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 130 132 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 132 134 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 134 136 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 136 138 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 138 140 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 140 142 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 142 144 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 144 146 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 146 148 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 148 150 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 150 152 Metaperidotite

GTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 152 154 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 154 156 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 156 158 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 158 160 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 160 162 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 162 164 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 164 166 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 166 168 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 168 170 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 170 172 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 172 174 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 174 176 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 176 178 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 178 180 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 180 182 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 182 184 MetaperidotiteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 184 186 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 186 188 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 188 190 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 190 192 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 192 194 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 194 196 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 196 198 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 198 200 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-16 200 202 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 290 292 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 292 294 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 294 296 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 296 298 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 298 300 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 300 302 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 302 304 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 304 306 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 306 308 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 308 310 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 310 312 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 312 314 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 314 316 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 316 318 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 318 320 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 320 322 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 322 324 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 324 326 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 326 328 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 328 330 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 330 332 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 332 334 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 334 336 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 336 338 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 338 340 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 340 342 Olivinepyroxenite

GTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 342 344 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 344 346 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 346 348 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 348 350 OlivinepyroxeniteGTK-Mintech Mini-Pilot Bulk_con Cu-Ni KV-20 350 353.1 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 44 46 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 46 48 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 48 50 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 50 52 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 52 54 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 54 56 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 56 58 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 58 60 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 60 62 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 62 64 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 64 66 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 66 68 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 68 70 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 70 72 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 72 74 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 74 76 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 76 78 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 78 80 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 80 82 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 82 84 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 84 86 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 86 88 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 88 90 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 90 92 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 92 94 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 94 96 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 96 98 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 98 100 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 100 102 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 102 104 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 104 106 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 106 108 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 108 110 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 110 112 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 112 114 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 114 116 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 116 118 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 118 120 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 120 122 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 122 124 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 124 126 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 126 128 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 128 130 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 130 132 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 132 134 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 134 136 Olivinepyroxenite

GTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 136 138 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 138 140 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 140 142 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 142 144 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 144 146 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-22 146 148 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 24 26 Metased xenolithGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 26 28 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 28 30 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 30 32 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 32 34 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 34 36 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 36 38 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 38 40 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 40 42 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 42 44 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 44 46 Mixed olpx dunite GTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 46 48 Mixed olpx dunite GTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 48 50 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 50 52 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 52 54 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 54 56 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 56 58 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 58 60 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 60 62 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 62 64 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 64 66 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 66 68 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 68 70 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 70 72 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 72 74 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 74 76 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 76 78 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 78 80 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 80 82 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 82 84 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 84 86 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 86 88 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 88 90 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 90 92 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 92 94 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 94 96 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 96 98 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 98 100 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 100 102 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 102 104 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 104 106 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 106 108 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 108 110 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 110 112 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 112 114 Olivinepyroxenite

GTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 114 116 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 116 118 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 118 120 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 120 122 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 122 124 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 124 126 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 126 128 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 128 130 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 130 132 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 132 134 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 134 136 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-23 136 138 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 88 90 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 90 92 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 92 94 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 94 96 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 96 98 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 98 100 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 100 102 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 102 104 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 104 106 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 106 108 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 108 110 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 110 112 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 112 114 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 114 116 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 116 118 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 118 120 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 120 122 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 122 124 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 124 126 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 126 128 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 128 130 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 130 132 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 132 134 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 134 136 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 136 138 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 138 140 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 140 142 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 142 144 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 144 146 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 146 148 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 148 150 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 150 152 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 152 154 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 154 156 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 156 158 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 158 160 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 160 162 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 162 164 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 164 166 Olivinepyroxenite

GTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 166 168 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 168 170 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 170 172 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 172 174 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 174 176 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 176 178 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 178 180 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 180 182 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 182 184 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 184 186 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 186 188 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 188 190 MetaperidotiteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 190 192 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 192 194 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 194 196 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 196 198 OlivinepyroxeniteGTK-Mintech Mini-Pilot Sep_con Cu-Ni KV-24 198 200 OlivinepyroxeniteGTK-Mintech Bench scale Phase 4 Sep_con Cu-Ni KV-22 102 104 MetaperidotiteGTK-Mintech Bench scale Phase 4 Sep_con Cu-Ni KV-25 57 59 MetaperidotiteGTK-Mintech Bench scale Phase 4 Sep_con Cu-Ni KV-25 103 105 Olivinepyroxenite

Code geo CU NI CO S PT-FAMS PD-FINAL AU-FINAL X-intersect Y-intersect

1 0.004 0.344 0.007 0.243 1.040 0.516 0.020 3498819.00 7512473.981 0.006 0.594 0.008 0.394 0.332 0.265 0.032 3498820.54 7512473.952 0.006 0.550 0.007 0.393 1.270 0.945 0.051 3498822.07 7512473.931 0.015 0.404 0.006 0.220 1.080 0.860 0.041 3498823.60 7512473.902 0.006 0.187 0.004 0.088 0.304 0.200 0.008 3498825.13 7512473.872 0.082 0.600 0.007 0.636 0.966 0.716 0.149 3498826.66 7512473.852 0.116 0.648 0.010 0.867 1.030 0.707 0.195 3498828.20 7512473.822 0.034 0.498 0.008 0.573 1.270 0.889 0.087 3498829.73 7512473.792 0.043 0.409 0.006 0.535 2.270 1.650 0.072 3498831.26 7512473.771 0.135 0.747 0.010 0.926 4.390 3.430 0.147 3498832.79 7512473.742 0.081 0.544 0.009 0.723 0.882 0.693 0.098 3498834.32 7512473.712 0.150 0.403 0.011 0.886 0.347 0.266 0.106 3498835.85 7512473.692 0.151 0.370 0.010 0.896 0.319 0.246 0.109 3498837.39 7512473.662 0.188 0.596 0.012 1.130 0.499 0.364 0.145 3498838.92 7512473.632 0.073 0.734 0.013 1.420 0.560 0.456 0.067 3498840.45 7512473.612 0.287 0.665 0.019 2.340 0.730 0.725 0.142 3498841.98 7512473.582 0.283 0.510 0.013 1.830 0.524 0.403 0.131 3498843.51 7512473.552 0.276 0.580 0.018 1.470 0.418 0.340 0.177 3498845.05 7512473.532 0.189 0.782 0.012 1.470 0.596 0.478 0.180 3498846.58 7512473.501 0.199 0.932 0.011 1.420 0.672 0.534 0.183 3498848.11 7512473.471 0.217 1.100 0.011 1.510 0.719 0.610 0.171 3498849.64 7512473.451 0.162 0.950 0.010 1.290 0.996 1.880 0.154 3498851.17 7512473.422 0.137 0.722 0.009 1.090 2.290 2.130 0.102 3498852.70 7512473.392 0.015 0.703 0.008 0.818 3.720 3.070 0.093 3498854.24 7512473.372 0.222 0.879 0.014 1.630 0.930 0.759 0.228 3498855.77 7512473.341 0.440 1.120 0.016 2.280 1.800 1.330 0.265 3498857.30 7512473.312 0.375 1.100 0.012 1.990 2.390 2.180 0.196 3498858.83 7512473.282 0.255 1.520 0.011 1.960 1.580 1.650 0.337 3498860.36 7512473.262 0.167 0.518 0.010 1.090 1.800 1.210 0.112 3498861.90 7512473.232 0.093 0.284 0.008 0.724 1.120 0.964 0.135 3498814.23 7512524.052 0.043 0.384 0.015 1.060 0.624 0.693 0.157 3498815.77 7512524.051 0.266 0.410 0.009 1.060 0.821 0.446 0.182 3498852.47 7512449.211 0.178 0.339 0.009 0.905 1.620 0.474 0.081 3498854.00 7512449.21

1 0.895 0.366 0.018 2.450 0.367 0.211 0.186 3498922.33 7512475.141 0.757 0.287 0.015 2.000 0.281 0.164 0.158 3498920.94 7512475.161 0.716 0.339 0.016 2.310 0.313 0.221 0.202 3498919.54 7512475.201 0.545 0.292 0.014 1.880 0.278 0.196 0.181 3498918.15 7512475.241 0.762 0.289 0.014 2.030 0.271 0.183 0.119 3498916.76 7512475.281 0.497 0.232 0.012 1.480 0.211 0.144 0.128 3498915.38 7512475.311 0.383 0.219 0.013 1.360 0.205 0.142 0.132 3498914.00 7512475.351 0.292 0.184 0.011 1.060 0.145 0.099 0.098 3498912.62 7512475.392 0.067 0.074 0.005 0.392 0.041 0.025 0.019 3498911.24 7512475.432 0.049 0.071 0.005 0.414 0.031 0.017 0.018 3498909.87 7512475.481 0.202 0.146 0.009 0.833 0.110 0.070 0.072 3498908.50 7512475.521 0.300 0.189 0.011 1.120 0.163 0.119 0.106 3498907.12 7512475.571 0.329 0.199 0.011 1.210 0.181 0.127 0.110 3498905.75 7512475.621 0.428 0.230 0.012 1.430 0.214 0.155 0.128 3498904.38 7512475.681 0.449 0.234 0.012 1.640 0.254 0.178 0.135 3498903.01 7512475.72

1 0.380 0.267 0.014 1.830 0.297 0.218 0.142 3498901.65 7512475.781 0.571 0.296 0.014 1.880 0.293 0.201 0.199 3498900.29 7512475.841 0.617 0.325 0.015 1.920 0.309 0.201 0.202 3498898.92 7512475.811 0.526 0.322 0.015 1.900 0.327 0.189 0.198 3498897.56 7512475.921 0.422 0.378 0.017 2.330 0.366 0.332 0.178 3498896.20 7512476.041 0.386 0.342 0.016 2.220 0.318 0.267 0.176 3498894.85 7512476.071 0.522 0.343 0.016 2.220 0.380 0.239 0.225 3498893.49 7512476.131 0.646 0.332 0.016 2.240 0.391 0.224 0.263 3498892.13 7512476.191 0.829 0.320 0.016 2.260 0.350 0.196 0.301 3498892.78 7512476.261 0.436 0.343 0.017 2.140 0.300 0.233 0.177 3498890.10 7512476.331 0.558 0.289 0.015 1.650 0.308 0.215 0.189 3498939.21 7512474.901 0.664 0.305 0.015 2.060 0.365 0.239 0.252 3498937.80 7512474.901 0.321 0.257 0.014 1.400 0.228 0.192 0.135 3498936.38 7512474.921 0.418 0.292 0.016 1.850 0.346 0.253 0.209 3498934.97 7512474.931 0.392 0.325 0.017 1.820 0.345 0.283 0.137 3498933.56 7512474.951 0.423 0.311 0.016 1.960 0.321 0.255 0.173 3498932.15 7512474.961 0.360 0.288 0.015 1.780 0.258 0.224 0.157 3498930.75 7512474.982 0.414 0.238 0.014 1.470 0.187 0.134 0.117 3498929.34 7512475.002 0.572 0.278 0.015 1.730 0.273 0.175 0.156 3498927.93 7512475.032 0.218 0.195 0.012 0.793 0.201 0.148 0.081 3498926.53 7512475.052 0.186 0.172 0.010 1.030 0.121 0.092 0.060 3498925.13 7512475.072 0.488 0.310 0.016 1.920 0.267 0.183 0.155 3498923.72 7512475.111 0.714 0.345 0.017 2.380 0.316 0.206 0.233 3498887.41 7512476.391 0.642 0.362 0.017 2.260 0.289 0.191 0.204 3498886.73 7512476.471 0.641 0.355 0.017 2.260 0.301 0.203 0.209 3498885.38 7512476.541 0.612 0.354 0.017 2.290 0.290 0.219 0.196 3498884.03 7512476.621 0.547 0.333 0.016 2.070 0.309 0.196 0.188 3498882.68 7512476.691 0.652 0.318 0.016 2.110 0.276 0.196 0.199 3498881.34 7512476.761 0.573 0.372 0.018 2.410 0.386 0.257 0.325 3498879.99 7512476.841 0.762 0.356 0.018 2.380 0.319 0.199 0.238 3498878.64 7512476.921 0.648 0.317 0.017 2.150 0.293 0.170 0.218 3498877.30 7512477.001 0.580 0.358 0.018 2.260 0.295 0.203 0.201 3498875.95 7512477.082 0.618 0.651 0.031 5.400 0.261 0.330 0.131 3498874.61 7512477.162 0.310 0.407 0.019 3.010 0.412 0.304 0.132 3498873.27 7512477.241 0.424 0.369 0.018 2.580 0.393 0.249 0.240 3498871.92 7512477.321 0.694 0.406 0.019 2.870 0.410 0.243 0.232 3498870.58 7512477.411 0.589 0.368 0.016 2.320 0.331 0.194 0.188 3498869.24 7512477.501 0.751 0.375 0.016 2.290 0.306 0.205 0.276 3498867.90 7512477.581 0.851 0.383 0.016 2.320 0.360 0.255 0.232 3498866.56 7512477.671 0.226 0.189 0.012 1.410 0.222 0.137 0.108 3498920.73 7512530.391 0.217 0.185 0.011 1.220 0.140 0.085 0.107 3498919.38 7512530.531 0.168 0.151 0.010 0.939 0.145 0.085 0.073 3498918.03 7512530.671 0.253 0.205 0.012 1.250 0.148 0.098 0.086 3498916.69 7512530.821 0.607 0.429 0.019 2.480 0.153 0.147 0.037 3498915.34 7512530.961 0.289 0.208 0.012 1.290 0.203 0.140 0.109 3498914.00 7512531.101 0.281 0.190 0.012 1.270 0.208 0.139 0.122 3498912.65 7512531.251 0.484 0.297 0.015 1.870 0.147 0.102 0.079 3498911.31 7512531.391 0.391 0.210 0.013 1.300 0.304 0.188 0.229 3498909.97 7512531.542 0.181 0.206 0.013 1.570 0.146 0.096 0.068 3498908.63 7512531.692 0.252 0.387 0.024 3.600 0.188 0.114 0.046 3498907.29 7512531.832 0.329 0.250 0.016 2.440 0.175 0.126 0.111 3498905.95 7512531.98

2 0.241 0.254 0.014 2.240 0.240 0.166 0.127 3498904.61 7512532.122 0.147 0.200 0.009 1.400 0.160 0.103 0.073 3498903.27 7512532.272 0.089 0.187 0.008 1.460 0.128 0.080 0.063 3498901.93 7512532.422 0.196 0.210 0.016 1.780 0.198 0.114 0.124 3498900.59 7512532.561 0.275 0.290 0.015 2.080 0.284 0.179 0.110 3498899.25 7512532.711 0.486 0.341 0.017 2.260 0.338 0.225 0.210 3498897.91 7512532.861 0.446 0.381 0.019 2.560 0.293 0.192 0.118 3498896.57 7512533.011 0.483 0.329 0.016 2.160 0.312 0.192 0.179 3498895.23 7512533.161 0.499 0.389 0.018 2.410 0.429 0.274 0.231 3498893.90 7512533.311 0.641 0.488 0.021 3.010 0.461 0.282 0.258 3498892.56 7512533.461 0.897 0.412 0.018 2.940 0.503 0.278 0.307 3498891.23 7512533.611 0.585 0.386 0.018 2.660 0.435 0.248 0.264 3498889.89 7512533.771 0.580 0.493 0.022 3.370 0.338 0.218 0.203 3498888.56 7512533.921 0.329 0.609 0.029 4.060 0.195 0.152 0.078 3498887.23 7512534.071 0.328 0.647 0.029 4.120 0.210 0.146 0.054 3498885.89 7512534.232 0.572 0.388 0.018 2.740 0.281 0.178 0.180 3498884.56 7512534.391 0.512 0.358 0.017 2.370 0.346 0.278 0.199 3498883.23 7512534.551 0.636 0.500 0.022 3.050 0.240 0.155 0.237 3498881.90 7512534.701 0.649 0.520 0.022 3.150 0.282 0.181 0.212 3498880.57 7512534.861 0.723 0.380 0.018 2.330 0.303 0.172 0.103 3498879.24 7512535.031 0.645 0.482 0.020 2.750 0.465 0.259 0.219 3498877.91 7512535.191 0.592 0.466 0.020 2.700 0.416 0.263 0.175 3498876.58 7512535.351 0.577 0.372 0.018 2.260 0.371 0.217 0.195 3498875.26 7512535.521 0.383 0.403 0.020 2.590 0.295 0.191 0.129 3498873.93 7512535.691 0.445 0.330 0.017 2.150 0.309 0.178 0.161 3498872.60 7512535.851 0.708 0.329 0.0157 2.11 0.212 0.148 0.141 3498805.7 7512300.91 0.629 0.327 0.0155 2.03 0.198 0.135 0.11 3498806.7 7512300.91 0.72 0.311 0.0156 2.05 0.184 0.131 0.12 3498807.7 7512300.91 0.721 0.363 0.0169 2.31 0.195 0.138 0.137 3498808.7 7512300.91 0.888 0.414 0.0185 2.78 0.272 0.164 0.161 3498809.7 7512300.91 0.586 0.408 0.0178 2.45 0.237 0.164 0.125 3498810.7 7512300.91 0.848 0.364 0.0177 2.68 0.223 0.0974 0.298 3498811.7 7512300.91 0.498 0.386 0.0176 2.31 0.243 0.124 0.136 3498812.7 7512300.91 0.378 0.328 0.0145 1.9 0.21 0.154 0.0896 3498813.7 7512300.91 0.373 0.286 0.0136 1.7 0.15 0.104 0.0776 3498814.7 7512300.91 0.511 0.366 0.0161 2.25 0.195 0.133 0.109 3498815.7 7512300.91 0.717 0.379 0.0175 2.46 0.224 0.158 0.147 3498816.7 7512300.91 0.773 0.38 0.0169 2.54 0.233 0.165 0.163 3498817.7 7512301.01 0.484 0.411 0.0168 2.57 0.197 0.131 0.108 3498818.7 7512301.01 0.524 0.389 0.017 2.33 0.198 0.163 0.128 3498819.7 7512301.01 0.673 0.41 0.0179 2.47 0.21 0.163 0.159 3498820.6 7512301.01 0.821 0.369 0.0172 2.6 0.247 0.189 0.221 3498821.6 7512301.01 0.957 0.46 0.0206 3.22 0.414 0.265 0.276 3498822.6 7512301.01 0.589 0.412 0.0195 2.58 0.288 0.195 0.182 3498823.6 7512301.01 0.562 0.433 0.0201 2.65 0.229 0.175 0.173 3498824.6 7512301.01 0.8 0.393 0.0174 2.5 0.187 0.15 0.153 3498825.6 7512301.01 0.704 0.328 0.0156 2.09 0.145 0.114 0.0922 3498826.6 7512301.01 0.743 0.373 0.0169 2.35 0.151 0.148 0.112 3498827.5 7512301.01 0.816 0.429 0.0182 2.55 0.186 0.171 0.156 3498828.5 7512301.11 0.842 0.431 0.0178 2.55 0.21 0.178 0.174 3498829.5 7512301.11 0.827 0.427 0.0185 2.55 0.222 0.156 0.18 3498830.5 7512301.1

1 0.941 0.423 0.0184 2.76 0.325 0.198 0.209 3498831.5 7512301.11 0.69 0.529 0.0216 3.12 0.442 0.313 0.195 3498832.5 7512301.11 0.93 0.452 0.0197 3.02 0.393 0.205 0.223 3498833.5 7512301.11 0.671 0.367 0.0185 2.72 0.233 0.148 0.148 3498834.4 7512301.12 0.51 0.35 0.0162 2.28 0.163 0.115 0.0933 3498835.4 7512301.11 0.683 0.254 0.015 2.170 0.208 0.111 0.203 3498875.66 7512351.071 1.020 0.204 0.014 2.270 0.161 0.085 0.253 3498876.78 7512351.091 0.854 0.199 0.013 2.080 0.177 0.085 0.229 3498877.91 7512351.111 0.318 0.312 0.014 2.430 0.209 0.188 0.091 3498879.04 7512351.131 0.549 0.293 0.014 1.850 0.185 0.131 0.141 3498880.16 7512351.151 0.763 0.256 0.013 1.880 0.178 0.097 0.186 3498881.28 7512351.171 0.668 0.250 0.013 1.750 0.198 0.103 0.190 3498882.41 7512351.191 0.930 0.304 0.015 2.370 0.254 0.119 0.189 3498883.53 7512351.211 0.799 0.320 0.014 2.180 0.221 0.117 0.192 3498884.65 7512351.241 0.777 0.353 0.015 2.380 0.251 0.131 0.187 3498885.77 7512351.261 0.708 0.325 0.015 2.200 0.190 0.110 0.184 3498886.90 7512351.281 0.326 0.297 0.014 1.690 0.167 0.111 0.108 3498888.02 7512351.312 0.130 0.208 0.010 1.580 0.121 0.073 0.048 3498889.14 7512351.332 0.140 0.282 0.014 2.140 0.144 0.120 0.086 3498890.25 7512351.361 0.394 0.299 0.015 1.950 0.173 0.112 0.146 3498891.37 7512351.381 0.620 0.318 0.016 2.340 0.214 0.113 0.215 3498892.49 7512351.411 0.478 0.324 0.016 2.190 0.241 0.127 0.191 3498893.61 7512351.431 0.356 0.210 0.012 1.460 0.145 0.072 0.122 3498894.72 7512351.461 0.121 0.100 0.009 0.574 0.050 0.026 0.037 3498895.84 7512351.481 0.137 0.107 0.009 0.720 0.053 0.025 0.042 3498896.95 7512351.511 0.076 0.075 0.007 0.343 0.032 0.009 0.022 3498898.07 7512351.531 0.101 0.090 0.008 0.442 0.032 0.016 0.024 3498899.18 7512351.561 0.121 0.107 0.008 0.583 0.050 0.026 0.036 3498900.29 7512351.581 0.113 0.091 0.008 0.553 0.045 0.019 0.033 3498901.40 7512351.612 0.047 0.086 0.008 0.519 0.048 0.019 0.024 3498902.52 7512351.632 0.029 0.030 0.003 0.110 0.012 0.006 0.008 3498903.63 7512351.662 0.043 0.017 0.002 0.073 0.009 0.004 0.006 3498904.74 7512351.682 0.010 0.043 0.006 0.101 0.018 0.008 0.007 3498905.85 7512351.702 0.020 0.056 0.006 0.209 0.024 0.010 0.012 3498906.96 7512351.722 0.169 0.116 0.009 0.586 0.067 0.036 0.051 3498908.07 7512351.751 0.459 0.189 0.012 1.300 0.134 0.067 0.124 3498909.17 7512351.771 0.246 0.144 0.010 0.960 0.104 0.053 0.085 3498910.28 7512351.792 0.357 0.307 0.016 2.020 0.208 0.145 0.138 3498911.38 7512351.811 0.542 0.350 0.017 2.260 0.248 0.182 0.192 3498912.49 7512351.831 0.784 0.333 0.016 2.330 0.244 0.156 0.274 3498913.59 7512351.851 0.640 0.362 0.017 2.330 0.286 0.153 0.220 3498914.69 7512351.871 0.680 0.346 0.017 2.380 0.275 0.162 0.238 3498915.79 7512351.901 0.476 0.343 0.017 2.260 0.299 0.161 0.199 3498916.89 7512351.921 0.458 0.387 0.017 2.350 0.275 0.208 0.163 3498917.98 7512351.941 0.688 0.381 0.017 2.450 0.346 0.199 0.211 3498919.08 7512351.971 0.963 0.347 0.017 2.600 0.280 0.120 0.209 3498920.18 7512352.001 0.624 0.396 0.018 2.440 0.300 0.165 0.240 3498921.27 7512352.021 0.660 0.370 0.018 2.530 0.329 0.184 0.234 3498922.36 7512352.052 0.497 0.368 0.018 2.420 0.335 0.209 0.138 3498923.45 7512352.082 0.555 0.315 0.016 2.310 0.297 0.294 0.200 3498924.54 7512352.111 0.805 0.351 0.017 2.390 0.490 0.217 0.214 3498925.64 7512352.14

1 0.582 0.257 0.014 1.610 0.244 0.125 0.203 3498926.72 7512352.171 0.312 0.177 0.011 0.967 0.185 0.109 0.110 3498927.81 7512352.201 0.200 0.140 0.010 0.646 0.126 0.081 0.071 3498928.90 7512352.231 0.232 0.155 0.010 0.707 0.167 0.103 0.085 3498929.99 7512352.271 0.440 0.283 0.014 1.560 0.367 0.227 0.180 3498931.07 7512352.302 0.435 0.258 0.014 2.140 0.324 0.170 0.159 3498932.16 7512352.343 0.419 0.202 0.010 1.470 0.190 0.139 0.138 3498868.98 7512400.041 0.544 0.329 0.014 1.950 0.310 0.238 0.171 3498870.10 7512400.051 0.545 0.428 0.017 2.270 0.332 0.267 0.150 3498871.23 7512400.061 0.628 0.284 0.014 1.830 0.278 0.175 0.165 3498872.36 7512400.071 0.802 0.241 0.014 1.820 0.279 0.158 0.159 3498873.48 7512400.081 0.741 0.294 0.015 2.030 0.277 0.190 0.236 3498874.60 7512400.101 0.764 0.379 0.017 2.550 0.365 0.267 0.246 3498875.73 7512400.111 0.676 0.296 0.015 2.010 0.443 0.218 0.233 3498876.85 7512400.131 0.691 0.302 0.015 2.040 0.371 0.224 0.260 3498877.97 7512400.141 0.549 0.357 0.017 2.130 0.399 0.256 0.228 3498879.10 7512400.163 0.430 0.290 0.016 1.660 0.252 0.184 0.228 3498880.22 7512400.183 0.184 0.157 0.010 0.729 0.117 0.094 0.059 3498881.34 7512400.211 0.266 0.311 0.016 1.610 0.268 0.230 0.111 3498882.46 7512400.231 0.537 0.323 0.015 1.900 0.367 0.244 0.138 3498883.59 7512400.251 0.723 0.238 0.013 1.770 0.323 0.188 0.142 3498884.71 7512400.281 0.474 0.239 0.011 1.460 0.223 0.158 0.233 3498885.83 7512400.301 0.535 0.288 0.013 1.740 0.233 0.162 0.151 3498886.95 7512400.331 0.554 0.305 0.013 1.790 0.251 0.173 0.161 3498888.07 7512400.351 0.341 0.279 0.013 1.570 0.239 0.170 0.119 3498889.19 7512400.381 0.130 0.258 0.012 1.390 0.188 0.208 0.077 3498890.31 7512400.411 0.445 0.253 0.013 1.600 0.308 0.195 0.136 3498891.43 7512400.441 0.376 0.294 0.015 1.800 0.235 0.188 0.144 3498892.54 7512400.472 0.120 0.150 0.010 0.739 0.082 0.086 0.063 3498893.66 7512400.502 0.446 0.275 0.014 1.870 0.161 0.118 0.120 3498894.77 7512400.532 0.296 0.124 0.007 0.935 0.103 0.064 0.087 3498895.89 7512400.552 0.373 0.178 0.010 1.420 0.132 0.086 0.123 3498897.00 7512400.582 0.467 0.262 0.014 2.090 0.157 0.111 0.126 3498898.12 7512400.612 0.449 0.199 0.011 1.740 0.125 0.075 0.123 3498899.23 7512400.642 0.448 0.220 0.014 2.120 0.112 0.072 0.106 3498900.34 7512400.672 0.203 0.113 0.008 1.060 0.045 0.029 0.041 3498901.46 7512400.702 0.271 0.149 0.009 1.250 0.058 0.039 0.058 3498902.57 7512400.742 0.614 0.265 0.012 2.030 0.164 0.104 0.154 3498903.68 7512400.772 0.396 0.203 0.010 1.430 0.142 0.081 0.131 3498904.79 7512400.802 0.613 0.328 0.014 2.350 0.258 0.185 0.177 3498905.90 7512400.841 0.303 0.369 0.015 2.050 0.244 0.175 0.134 3498907.01 7512400.871 0.577 0.373 0.016 2.370 0.255 0.162 0.232 3498908.12 7512400.911 0.585 0.350 0.017 2.460 0.214 0.155 0.207 3498909.22 7512400.951 0.554 0.354 0.017 2.410 0.197 0.120 0.121 3498910.32 7512400.991 0.527 0.363 0.017 2.330 0.239 0.206 0.206 3498911.42 7512401.031 0.542 0.295 0.014 1.840 0.209 0.104 0.155 3498912.52 7512401.081 0.442 0.241 0.012 1.450 0.163 0.091 0.104 3498913.62 7512401.121 0.520 0.313 0.014 2.020 0.198 0.135 0.185 3498914.71 7512401.171 0.587 0.307 0.015 2.250 0.248 0.149 0.190 3498915.80 7512401.221 0.658 0.316 0.015 2.220 0.229 0.160 0.178 3498916.89 7512401.271 0.771 0.351 0.016 2.360 0.288 0.165 0.272 3498917.97 7512401.32

1 0.729 0.351 0.015 2.330 0.279 0.151 0.206 3498919.06 7512401.371 0.696 0.369 0.016 2.340 0.274 0.146 0.231 3498920.14 7512401.421 0.629 0.400 0.017 2.380 0.273 0.174 0.231 3498921.22 7512401.481 0.596 0.419 0.017 2.460 0.392 0.244 0.210 3498922.30 7512401.531 0.523 0.413 0.017 2.310 0.248 0.219 0.184 3498923.38 7512401.591 0.767 0.379 0.017 2.440 0.250 0.155 0.189 3498924.46 7512401.641 0.795 0.388 0.017 2.450 0.255 0.150 0.218 3498925.53 7512401.701 0.718 0.305 0.015 2.010 0.165 0.099 0.119 3498926.61 7512401.751 0.601 0.254 0.013 1.670 0.163 0.099 0.138 3498927.68 7512401.811 0.570 0.253 0.013 1.620 0.163 0.091 0.091 3498928.76 7512401.871 0.760 0.365 0.018 2.460 0.295 0.118 0.218 3498929.83 7512401.921 0.607 0.407 0.019 2.620 0.271 0.203 0.214 3498930.90 7512401.982 0.237 0.235 0.012 1.630 0.237 0.155 0.109 3498907.76 7512450.252 0.337 0.249 0.012 1.750 0.210 0.155 0.160 3498908.71 7512450.272 0.391 0.300 0.015 1.970 0.323 0.205 0.160 3498909.66 7512450.282 0.287 0.371 0.017 2.190 0.379 0.269 0.137 3498910.61 7512450.292 0.371 0.366 0.017 2.330 0.356 0.266 0.155 3498911.55 7512450.312 0.151 0.434 0.020 2.550 0.456 0.349 0.107 3498912.50 7512450.322 0.468 0.338 0.016 2.480 0.610 0.311 0.157 3498913.45 7512450.332 0.175 0.336 0.016 2.310 0.337 0.233 0.081 3498914.40 7512450.342 0.124 0.307 0.015 2.000 0.287 0.229 0.067 3498915.35 7512450.352 0.291 0.326 0.017 2.470 0.294 0.303 0.234 3498916.29 7512450.362 0.391 0.289 0.017 2.600 0.259 0.223 0.214 3498917.24 7512450.372 0.238 0.330 0.019 2.860 0.263 0.273 0.180 3498918.19 7512450.381 0.300 0.362 0.019 2.870 0.282 0.240 0.184 3498919.14 7512450.401 0.491 0.359 0.018 2.570 0.211 0.205 0.218 3498920.09 7512450.421 0.329 0.394 0.019 2.610 0.357 0.317 0.189 3498921.04 7512450.441 0.542 0.382 0.019 2.690 0.293 0.314 0.193 3498921.98 7512450.471 0.366 0.403 0.019 2.460 0.252 0.250 0.200 3498922.92 7512450.501 0.326 0.427 0.019 2.650 0.269 0.304 0.212 3498923.87 7512450.532 0.195 0.412 0.017 2.440 0.245 0.158 0.082 3498924.81 7512450.562 0.105 0.283 0.012 1.640 0.166 0.102 0.031 3498925.75 7512450.602 0.105 0.287 0.014 1.630 0.375 0.189 0.082 3498926.69 7512450.642 0.410 0.400 0.017 2.030 0.235 0.101 0.053 3498927.62 7512450.682 0.235 0.240 0.011 1.370 0.145 0.088 0.093 3498928.55 7512450.722 0.253 0.306 0.015 1.810 0.100 0.065 0.040 3498929.48 7512450.762 0.665 0.273 0.014 2.170 0.257 0.129 0.149 3498930.41 7512450.802 0.825 0.334 0.015 2.690 0.237 0.163 0.124 3498931.34 7512450.852 0.442 0.431 0.018 2.520 0.343 0.197 0.161 3498932.26 7512450.891 0.680 0.359 0.019 2.450 0.264 0.126 0.223 3498933.19 7512450.931 0.773 0.323 0.018 2.560 0.310 0.119 0.340 3498934.12 7512450.981 0.744 0.387 0.019 2.620 0.395 0.218 0.180 3498935.04 7512451.021 0.831 0.424 0.019 2.720 0.272 0.205 0.268 3498935.97 7512451.071 0.894 0.403 0.018 2.630 0.317 0.153 0.312 3498936.89 7512451.111 0.894 0.407 0.018 2.650 0.342 0.152 0.189 3498937.81 7512451.161 0.847 0.414 0.019 2.640 0.328 0.231 0.251 3498938.73 7512451.201 0.786 0.443 0.019 2.650 0.375 0.205 0.248 3498939.64 7512451.261 0.923 0.408 0.017 2.560 0.332 0.168 0.196 3498940.56 7512451.311 0.926 0.432 0.018 2.640 0.393 0.203 0.278 3498941.47 7512451.361 0.890 0.488 0.020 2.880 0.450 0.220 0.259 3498942.39 7512451.421 0.580 0.500 0.019 2.570 0.388 0.246 0.293 3498943.30 7512451.47

1 0.576 0.369 0.016 2.140 0.294 0.207 0.213 3498944.21 7512451.531 0.350 0.251 0.013 1.420 0.175 0.101 0.126 3498945.11 7512451.591 0.478 0.413 0.017 2.360 0.342 0.276 0.222 3498946.02 7512451.651 0.986 0.360 0.017 2.460 0.413 0.198 0.318 3498946.93 7512451.701 0.664 0.206 0.012 1.450 0.193 0.095 0.130 3498947.83 7512451.761 1.050 0.343 0.016 2.410 0.339 0.161 0.240 3498948.74 7512451.821 0.837 0.413 0.017 2.510 0.357 0.234 0.222 3498949.64 7512451.881 0.671 0.430 0.017 2.460 0.416 0.221 0.215 3498950.54 7512451.941 0.643 0.419 0.018 2.600 0.417 0.272 0.283 3498951.44 7512452.001 0.736 0.406 0.018 2.620 0.404 0.223 0.228 3498952.33 7512452.072 0.466 0.451 0.019 2.570 0.449 0.330 0.245 3498953.23 7512452.132 0.366 0.601 0.025 2.670 0.401 0.292 0.181 3498954.13 7512452.191 0.685 0.399 0.018 2.540 0.448 0.279 0.228 3498955.03 7512452.251 0.800 0.379 0.017 2.440 0.381 0.240 0.491 3498955.93 7512452.311 0.865 0.316 0.015 2.140 0.380 0.170 0.193 3498956.83 7512452.371 0.667 0.285 0.014 1.870 0.295 0.149 0.199 3498957.73 7512452.441 0.689 0.345 0.016 2.200 0.407 0.191 0.247 3498958.62 7512452.502 0.169 0.116 0.009 0.586 0.067 0.036 0.051 3498908.07 7512351.752 0.125 0.268 0.016 2.880 0.172 0.161 0.122 3498666.52 7512450.671 0.645 0.625 0.026 3.580 0.718 0.507 0.414 3498694.70 7512451.54

Z-intersect

221.72220.43219.15217.86216.58215.29214.01212.72211.43210.15208.86207.58206.29205.01203.72202.44201.15199.86198.58197.29196.01194.72193.44192.15190.87189.58188.29187.01185.72232.27230.98189.68188.39

212.64211.22209.77208.34206.90205.45204.00202.56201.11199.66198.20196.75195.30193.84192.38

190.92189.46187.99186.52185.06183.59182.12180.66179.18177.71229.70228.28226.87225.46224.04222.62221.20219.78218.35216.92215.50214.07176.23174.76173.29171.81170.34168.86167.38165.91164.43162.95161.48160.00158.52157.03155.55154.07152.59139.76138.28136.81135.34133.87132.39130.92129.44127.97126.49125.01123.53

122.06120.58119.10117.63116.15114.67113.19111.71110.23108.75107.27105.79104.31102.82101.3499.8598.3796.8895.4093.9192.4390.9489.4687.9786.48

-14.3-16.0-17.7-19.5-21.2-22.9-24.7-26.4-28.1-29.8-31.6-33.3-35.1-36.8-38.5-40.3-42.0-43.7-45.5-47.2-49.0-50.7-52.5-54.2-55.9-57.7

-59.4-61.2-62.9-64.6-66.4

200.33198.67197.02195.37193.71192.06190.41188.75187.09185.44183.78182.13180.47178.81177.15175.50173.84172.18170.52168.86167.20165.54163.87162.21160.55158.89157.22155.56153.90152.23150.57148.90147.23145.56143.89142.22140.55138.88137.21135.54133.86132.19130.52128.84127.16125.49

123.81122.13120.46118.78117.10115.42216.68215.03213.37211.72210.07208.41206.76205.10203.45201.79200.14198.48196.83195.17193.52191.86190.21188.55186.89185.23183.58181.92180.26178.60176.94175.28173.61171.95170.29168.63166.97165.31163.64161.98160.32158.65156.98155.32153.65151.97150.30148.63146.95145.27143.59

141.91140.23138.55136.87135.19133.50131.82130.13128.45126.76125.07123.39158.44156.68154.92153.16151.40149.64147.87146.11144.35142.59140.83139.07137.31135.55133.79132.02130.26128.50126.73124.97123.20121.43119.67117.90116.12114.35112.58110.81109.04107.26105.49103.72101.94100.1798.3996.6194.8393.0591.28

89.5087.7185.9384.1582.3780.5878.8077.0175.2373.4471.6669.8768.0966.3064.5262.7360.95

152.23190.30153.95

Appendix 5C –

Plan Showing Position of Metallurgy Samples, May 2005

Appendix 5D –

Comparison of Analytical Methods on the core from Kevitsa

Comparison of analytical methods

on

core material from Keivitsa

John L. Pedersen. May 3rd

2005

Background

Historically all drill core samples from Keivitsa are analysed for copper, nickel, cobalt etc.

by ICP after extraction with hot aqua regia. In metallurgical testing, the applied analytical

method is AAS after extraction with nitric acid. Composition of feed material based on the

two sets of analyses did not compare well, although a Student t-test at 95 % confidence

confirmed that the samples could be from the same population.

It was decided to do a systematic comparison of values obtained after different extractions

procedures and at the same time address the question about host mineral for nickel. Only

nickel locked up in sulphides will be available for concentration by flotation.

Methods used for comparison

Four methods were tested against each other.

1) To_GTK: ICP-AES finish after extraction with hot aqua regia at 90 degree.

2) To_MinT: AAS analyses after total dissolution (HNO3 + HCl + HF + HClO4)

3) HNO_MinT: AAS after extraction with nitric acid.

4) Brom_MinT: AAS after extraction with bromine-methanol

Where GTK is the GTK laboratory at Rovaniemi and MinT is the GTK laboratory at

Outokumpu.

2

Copper

Copper analysed by GTK after hot aqua extraction is 7,5 % larger than copper analysed

after total dissolution at Mintek. There is good correlation between the two data sets

(Figure 1). The difference is not reasonable and must depend on differences in the

calibration curves.

y = 1,0747x - 48,037

R2 = 0,9991

0

1000

2000

3000

4000

5000

6000

0,0 1000,0 2000,0 3000,0 4000,0 5000,0 6000,0

Cu - ppm

Total_GTK

Cu - ppm

Total_MinT

Figure 1: Copper after hot aqua regia as a function of total copper (Mintek)

Copper analysed after Nitric acid extraction by Mintek is 2,7 % larger than total copper

analysed Mintek. There is very good correlation between the two data sets (Figure 2).

y = 1,0271x - 8,5939

R2 = 0,9998

0,0

1000,0

2000,0

3000,0

4000,0

5000,0

6000,0

0,0 1000,0 2000,0 3000,0 4000,0 5000,0 6000,0

Cu - ppm

HNO3_MinT

Cu - ppm

Total_MinT

Figure 2: Copper after Nitric acid as a function of total copper (both

Mintek)

3

Total copper analysed by GTK is 4,6 % larger than copper analysed by Mintek after Nitric

acid extraction. There is very good correlation between the two data sets (Figure 3). It will

require a larger data base to support this observation.

y = 1,0464x - 39,085

R2 = 0,9993

0

1000

2000

3000

4000

5000

6000

0,0 1000,0 2000,0 3000,0 4000,0 5000,0 6000,0

Cu - ppm

Total _GTK

Cu - ppm

HNO3_MinT

Figure 3: Copper after hot regia as a function of copper extracted with nitric

acid.

Copper analysed after Nitric acid extraction by Mintek is 3,0 % larger than copper

analysed after bromine methanol extraction (Mintek). There is very good correlation

between the two data sets (Figure 4).

y = 1,0297x + 13,428

R2 = 0,9998

0,0

1000,0

2000,0

3000,0

4000,0

5000,0

6000,0

0,0 1000,0 2000,0 3000,0 4000,0 5000,0 6000,0

Cu - ppm

HNO3_MinT

Cu - ppm

Brom_MinT

Figure 4: Copper after Nitric acid as a function of copper extracted with

bromine methanol (both Mintek).

4

Nickel

Nickel analysed after hot aqua regia extraction by GTK only identifies 92,1 % of the

amount of nickel, which is identified by Mintek in their analyses for nickel after total

dissolution. There is good correlation between the two data sets (Figure 5). The relation 1:1

is illustrated for comparison.

y = 0,9211x - 9,2477

R2 = 0,9936

0

2000

4000

6000

8000

10000

12000

0,0 2000,0 4000,0 6000,0 8000,0 10000,0 12000,0

Ni - ppm

Total_GTK

Ni - ppm

Total_MinT

Figure 5: Nickel (A.R) by GTK as a function of total nickel (Mintek)

Nickel after aqua regia extraction determined by GTK, Rovaniemi has a good correlation

with nickel determined by Mintek after nitric acid extraction. Within the relevant interval

of nickel content, the two methods deliver nearly identical results (Figure 6).

y = 0,9648x + 153,6

R2 = 0,9969

0

2000

4000

6000

8000

10000

12000

0,0 2000,0 4000,0 6000,0 8000,0 10000,0 12000,0

Ni - ppm

Total_GTK

Ni - ppm

HNO3_MinT

Figure 6: Nickel after aqua regia extraction (GTK) as a function of nickel

after nitric acid extraction (Mintek).

5

Nickel determined after nitric acid extraction as a function of nickel determined after

bromine methanol extraction, which only respond to nickel in sulphides, displays a

scattered correlation (Figure 7). Some samples display a 1:1 relationship, but in the

remaining samples the fraction of the nickel hosted in sulphide minerals is only 60 to 95 %

of total nickel.

The most frequent sulphide minerals are pyrrhotite, pentlandite and chalcopyrite. The

average composition for these minerals is given in Table 1. Weight per cent nickel is

similar to the weight percent for sulphur in pentlandite and weight percent copper is similar

to sulphur in chalcopyrite. The amount of sulphur taken up in pentlandite and chalcopyrite

is thus equal to the sum of nickel and copper. The number: “Sulphur – (Sum of nickel and

copper)” is thus a measure for whether there is enough sulphur available to form

pentlandite and chalcopyrite. The relation is plotted in Figure 8.

Table 1: Composition of main sulphide minerals from Keivitsa

y = 1,0465x + 363,22

R2 = 0,9679

0,0

2000,0

4000,0

6000,0

8000,0

10000,0

12000,0

0,0 1000,0 2000,0 3000,0 4000,0 5000,0 6000,0 7000,0 8000,0 9000,0 10000,0

Ni - ppm

HNO_MinT

Ni-ppm

Brom_Min

T

Figure 7: Nickel determined after nitric acid extraction as a function of

nickel determined after bromine methanol extraction (both by Mintek)

Mineral Sulphur Iron Nickel Copper Cobalt Note

Pyrrhotite 38,3 61,4 0,17 0,01 - Analyse

Pentlandite 32,6 33,2 31,6 0,0 1,4 Analyse

Chalcopyrite 34,9 30,4 - 34,6 - Idealformular

6

0

0,2

0,4

0,6

0,8

1

1,2

-5000,0 0,0 5000,0 10000,0 15000,0 20000,0

Sulphur surplus

Ni(Brom)/Ni(Hot aqua)

Figure 8: Plot of ratio between nickel( extracted by bromine methanol) and

nickel (extracted by hot aqua regia) as a function of the variable S-(Ni+Cu).

When the surplus of sulphur is at least 0,1 % the proportion of nickel locked into sulphides

I fixed at c. 95 %. At lesser surplus more and more nickel goes into silicates.

Instead of plotting the ratio Ni(brom) divided with Ni (hot aqua regia) the number for

sulphur surplus is plotted against Ni(brom) divided with Ni (HNO) (Figure 9). For samples

with a sulphur surplus of more than c. 0,3 %, 98-100 % of the nickel is hosted in sulphide

minerals.

0

0,2

0,4

0,6

0,8

1

1,2

-5000,0 0,0 5000,0 10000,0 15000,0 20000,0

Ni(Brom)/Ni(HNO)

S - (Ni + Cu)

Figure 9: Plot of ratio between nickel( extracted by bromine methanol) and

nickel (extracted by nitric acid) as a function of the variable S-(Ni+Cu).

7

Iron

The various extraction methods have different capacity for extraction of iron. Total

extraction is achieved with a mixture of HNO3 + HCl + HF + HClO4. Hot aqua regia

extracts 64 % of total iron and with nitric acid the proportion drops to 42 %. Bromine

methanol only extracts iron in sulphides.

Table 2: Extraction of iron as a function of extraction media. The values are

the average of the 15 samples in the analytical test programme.

Laboratory Methods Iron content

________________________________________________________

Mintek Total dissolution (HF et al.) 6,18 %

GTK – Rovaniemi Hot aqua regia 3,93 %

Mintek Nitric acid 2,60 %

Mintek Bromine methanol 0,98 %

Sulphur

Sulphur is analyses by the same methods at both laboratories (Figure 10). The two data sets

are well correlated. The sulphur values from GTK, Rovaniemi are 3,1 % higher than the

values from Mintek.

y = 10532x - 116, 4

R 2 = 0 , 9979

0

5000

10000

15000

20000

25000

0,00 0,50 1,00 1,50 2,00 2,50

S - p p m

GT K - R ov

S - %

M int ech

8

Figure10. Analyses of sulphur. GTH, Rovaniemi values as a function of

Mintek values.

Conclusions

Summary on copper observations

Efficiency of copper extraction is only slightly dependant on extraction methods. All

methods used are capable of extracting all or near all copper. The small differences

observed most likely are due to calibration variation in the laboratory procedure. These

differences are however large enough to be of importance in ore reserve calculations.

For metallurgical testing analysing is done after extraction with nitric acid. When values

from nitric acid extraction are compared with values from hot aqua regia extraction, which

is the standard methods used in analysing the samples, the values determined by GTK,

Rovaniemi are 4 % higher than after nitric acid extraction. This implies that compared to

the core analyses, a recovery of 95 % is actually only 91,3 %. Comparison for additional

samples will have to be done in order to support this observation.

Copper analysed after nitric extraction is 2,2 % higher than after extraction with bromine

methanol. Bromine methanol extracts exclusively copper locked in sulphide-bearing

minerals. As the values for total copper and copper after bromine methanol extraction are

identical, all copper occur in sulphide minerals. The impression is that when sulphur is not

available, copper stays in solution. It can not escape into silicate minerals.

Summary on nickel observations

Extraction of nickel is dependent on extraction methods. The two standard methods used 1)

nickel after hot regia aqua extraction and 2) nickel after nitric acid extraction give very

similar results. For low nickel content extraction with Hot aqua regia gives the highest

values, as extraction with nitric acid gives the highest values for high nickel content (cross

over point 0,44 % Ni).

Both standard methods mentioned above give lower value than total nickel analysed after

total dissolution. As copper - analysed after total extraction - actually is lower than copper

analyses after the less aggressive extraction methods, it indicates that certain

inconsistencies may be influencing the data. The small differences observed most likely

are due to calibration variation in the laboratory procedure.

The proposal that if surplus sulphur is available, then nickel goes into sulphide minerals is

supported by the limited number of observations. It seems that if sulphur - not taken up by

nickel and copper - is larger than 0,1 %, then nickel analysed after bromine methanol

extraction has a high fixed ratio to nickel extracted by hot aqua regia (95 %) or by nitric

acid (98 %). Additional analyses are required to support this observation.

The implication is that a recovery factor, which has been determined for a mineralization

type with “Surplus Sulphur” can be directly applied to the nickel grade determined from

9

analysis after hot aqua regia extraction in all cases where the surplus sulphur is large than

0,1 %.

The relation ship between nickel in sulphides and nickel in silicates is not well understood,

when the sulphur surplus is less than 0,1 %. Additional data are required to reveal this

relationship - if any.

Summary on iron observations

The amount of iron extraction by the various methods is in line with the assumed

efficiency of the various extraction methods. So although hot aqua regia extractions and

nitric extraction give the identical of nearly identical values for copper and nickel, it is

evident that hot aqua brings more iron into solution.

Iron in solution after bromine methanol extraction corresponds closely with iron in

chalcopyrite, pentlandite and iron sulphides.

Appendix 6A –

Process Flowsheet, Crushing and Grinding

Legend1.

Receiving

hopper&

feeder2.

Jawcrusher

3.B

eltconveyor12

m4.

Beltconveyor

50m

5.C

overedsurge

orepile,live

capacity24

h6.

Beltconveyor

50m,capacity

541t/h

7.S

AG

milD

8.5m

,5500kW

8.V

ibratingscreen

9.B

eltconveyor7

m10.B

eltconveyor17

m11.B

eltconveyor7

m12.H

ydrocycloneclassifier,2x12

cyclones13.B

allmill,D

7.0m

,L11.0

m,6000

kW

Kevitsa

Process

Flow

sheetC

rushingand

grindingO

reF

eedC

apacity:541t/h,4.5

Mt/a

Appendix

6A

1.

2.3.

4.

5.

6.

7.

8.

9.

10.

11.

12.

13.

Flotation

Appendix 6B –

Process Flowsheet, Copper Flotation

Appendix

6B

Ore

Feed

Capacity:541

t/h,4.5Mt/a

Legend14.C

opperrougher

flotation,6x70m

3,pH11

15.Copper

firstcleaningflotation,3x30

m3,pH

1216.C

oppersecond

cleaningflotation,3x20

m3,pH

1217.H

ighR

atethickener,D

9m

18.Pressure

filter,52m

219.B

eltconveyor20

m20.B

eltconveyor30

m21.C

opperconcentrate

storage22.C

onventionalthickener,D40

m23.Lam

ellathickener,A

4000m

2

14

15

16

1718

19

2021

2223

Regrind

mill

Nickelflotation

Appendix 6C –

Process Flowsheet, Nickel Flotation

Kevitsa

Process

Flow

sheetN

ickelFlotation

Ore

Feed

Capacity:541

t/h,4.5M

t/a

Appendix

6C

Legend24.N

ickelrougherflotation,6x130

m3,pH

525.N

ickelroughertailings:H

ighR

atethickener,D

30m

26.Regrind

mill,D

2.7m

,L3.6

m,300

kW27.H

ydrocycloneclassifier,1x12

cyclones28.N

ickelfirstcleaningflotation:3x50

m3,pH

11.529.N

ickelcleaningtailing

30.Nickelsecond

cleaningflotation:3x30

m3,pH

11.531.H

ighR

atethickener,D

9m

32.Pressure

filter,62m

233.B

eltconveyor20

m34.B

eltconveyor30

m35.N

ickelconcentratestorage

24

25

26

27

28

30

3132

33

3435

29

Appendix 6D –

Materials Balance

Kevitsa 4.5Mtpa Material Balance

Design Parameters:

dry tonnes p.a 4,500,000 Ore sg 3.15Crushing

Operating time/shifts 1,095Utilisation % 95% SAG mill o'size in dischar ge 5Mill recovery % 100Flotation

Operating time/shifts 1,095Utilisation % 95

NOTE: All purple cells denote defined parametersAll other cells are calculated

Kevitsa 4.5Mtpa Material Balance

Description dry tph Wt % % solids m3/h water m3/h pulp Tph pulp

Crushed product 540.7 100.0 97.0 16.7 188.4 557.5SAG mill feed water 0.0 0.0 0.0 360.4 360.4 360.4SAG mill feed 567.8 105.0 96.9 18.1 198.4 585.9SAG mill discharge 567.8 105.0 60.0 378.5 558.8 946.3Screen undersize 540.7 100.0 58.9 377.1 548.8 917.8Screen oversize 27.0 5.0 95.0 1.4 10.0 28.5

Cyclone feed water 0.0 0.0 0.0 624.9 624.9 624.9Cyclone feed 2040.7 377.4 53.0 1809.7 2457.6 3850.4Cyclone u/f 1500.0 277.4 65.0 807.7 1283.9 2307.7Ball mill feed 1500.0 277.4 65.0 807.7 1283.9 2307.7Ball mill discharge 1500.0 277.4 65.0 807.7 1283.9 2307.7Cyclone o/f 540.7 100.0 35.0 1004.2 1175.9 1545.0

Rougher Cu feed 540.7 100.0 35.0 1004.2 1175.9 1545.0Rougher Cu Conc 11.8 2.18 50.0 11.8 15.5 23.6Rougher Cu tailings 529.0 97.82 34.0 1026.8 1194.7 1555.7Cleaner Cu conc 7.4 1.37 50.0 7.4 9.8 14.8Cleaner Cu tails 5.9 1.09 34.0 11.5 13.4 17.4Recleaner Cu conc 5.9 1.08 50.0 5.9 7.7 11.7Recleaner Cu tails 1.5 0.29 34.0 3.0 3.5 4.6

Rougher Ni feed 529.0 97.8 35.0 982.3 1150.3 1511.3Rougher Ni Conc 51.3 9.49 50.0 51.3 67.6 102.6RougherNi tailings 477.7 88.33 34.0 927.2 1078.8 1404.9

Regrind ball mill feed 157.2 29.07 58.0 113.8 163.8 271.1Regrind ball mill disch. 157.2 29.07 58.0 113.8 163.8 271.1Cyclone feed water 0.0 0.0 0.0 28.7 28.7 28.7Regrind cyclone feed 157.2 29.07 53.0 139.4 189.3 296.6Regrind cyclone o/f 57.2 10.58 35.0 106.3 124.4 163.5Regrind cyclone u/f 100.0 18.49 65.0 53.8 85.6 153.8

Cleaner Ni conc 15.3 2.82 50.0 15.3 20.1 30.5Cleaner Ni tails 47.9 8.87 34.0 93.1 108.3 141.0Reclean Ni conc 9.3 1.72 50.0 9.3 12.2 18.6Reclean Ni tails 6.1 1.12 34.0 6.0 7.9 12.0

Pulp sg (g/l) % Copper Rec. Cu % T Cu % Nickel Rec. Ni % T Ni

2.959 0.35 100 1.89 0.30 100 1.621.0002.9531.6941.6732.844

1.0001.5671.7971.7971.7971.314 0.35 100 1.89 0.30 100 1.62

1.314 0.35 100.0 1.89 0.30 100.0 1.621.52 13.80 85.9 1.63 1.09 10.9 0.181.30 0.08 14.1 0.27 0.31 89.1 1.451.52 20.50 80.2 1.52 0.65 2.9 0.051.30 3.23 10.1 0.19 1.60 8.0 0.131.52 24.50 75.8 1.43 0.57 2.8 0.051.30 6.70 4.4 0.08 1.60 0.1 0.00

1.31 0.08 14.1 0.27 0.31 89.1 1.451.52 0.34 6.6 0.12 2.34 74.0 1.201.30 0.05 7.5 0.14 0.06 15.1 0.24

1.661.661.001.571.311.80

1.52 1.72 11.2 0.21 7.90 74.3 1.211.30 0.32 6.00 0.11 0.45 13.30 0.221.52 2.91 10.8 0.20 12.00 68.7 1.111.52 0.65 0.4 0.01 1.50 5.6 0.09

Appendix 6E –

Process Plant Layout

Legend:1.C

ontrolroomand

analysis2.S

AG

milland

vibratingscreen

3.Ballm

ill4.R

egrindingm

ill5.C

opperrougher

flotation6.C

opperfirstcleaning

flotation

100000

45000

15000

7.C

oppersecond

cleaningflotation

8.N

ickelrougherflotation

9.N

ickelfirstcleaningflotation

10.Nickelsecond

cleaningflotation

11.Copper

pressurefilter

12.Nickelpressure

filter

1

2

3

46

7

910

12 11

5

Appendix

6E

8

Appendix 6F –

Metal Recoveries

KeivitsaCo Recovery to

Test Number Ni concentrate (%)24

5 (data included for purpose of plotting cobalt recovery) 57.68678 68.549

Reference Test 18Phase 4 1

23

Stage Recovery of Pt toTest Number Stage Ni conc (%)

2 37.474 29.716 14.417 16.288 20.909 21.32

Reference Test 18 41.99Phase 4 1 34.83

2 65.563 75.15

Note for non mets:Stage recovery is recovery across the nickel circuit

Expected recoveries of pr

Pt Recovery to Au Recovery toNi Conc (%) Cu conc (%)

32.3 14.627.1 23.3

12.9 16.614.8 11.716.8 23.318.4 19.635.9 19.434.2 3.965.1 461.7 30.9

Stage Recovery of Au to Stage Ni Conc (%) Stage Recovery of Ni to Stage Ni conc (%31.73 65.2522.16 56.819.11 54.058.15 58.489.78 69.68

22.89 63.6612.78 66.2310.09 49.3044.38 83.4222.29 79.59

recious metals

Cu Recovery toCu conc (%) Ni recovery to Ni conc(%)

62.5 64.637.4 56.3

70.142.1 53.444.1 57.674.7 67.863.2 6245.9 65.114.9 49.119.8 8366.5 77.6

%)

Au

Re

co

ve

ry a

ga

ins

t C

u R

ec

ov

ery

05

10

15

20

25

30

35

01

02

03

04

05

06

07

08

0

Cu

Re

co

ve

ry (

%)

Au Recovery (%)

Min

i p

ilot

lab

te

st

resu

lts

Te

st

18

Ph

ase

4 T

estw

ork

Au

Reco

very

ag

ain

st

Ni R

eco

very

05

10

15

20

25

30

35

40

45

01

02

03

04

05

06

07

08

09

0

Ni

Re

co

ve

ry (

%)

Au Recovery (%)

Min

i p

ilot

resu

lts

Re

fere

nce

Te

st

18

Ph

ase

4 t

estw

ork

re

su

lts

Pt

Reco

very

ag

ain

st

Cu

Reco

very

y =

0.2

61

6x -

1.5

44

7

R2 =

0.9

10

4

05

10

15

20

25

01

02

03

04

05

06

07

08

0

Cu

Re

co

ve

ry (

%)

Pt Recovery (%)

Min

i p

ilot

test

resu

lts

Re

fere

nce

Te

st

18

Ph

ase

4 t

estw

ork

re

su

lts

Lin

ea

r (M

ini p

ilot

test

resu

lts)

Pt

an

d C

o R

eco

very

ag

ain

st

Ni R

eco

very

0

10

20

30

40

50

60

70

80

01

02

03

04

05

06

07

08

09

0

Ni

Re

co

ve

ry (

%)

Pt and Co Recovery (%)

Min

i p

ilot

test

resu

lts

Re

fere

nce

Te

st

18

Ph

ase

4 T

estw

ork

Co

ba

lt (

da

ta f

rom

min

i p

ilot

tests

5 a

nd

8)

Pd

reco

very

ag

ain

st

Ni R

eco

very

0

10

20

30

40

50

60

010

20

30

40

50

60

70

80

90

Ni R

eco

very

(%

)

Pd Recovery (%)

Min

i pilo

t te

st re

sults

Refe

rence T

est 18

Phase 4

test re

sults

Co

pp

er

Reco

very

to

Nic

kel C

on

cen

trate

y =

-0

.62

93

x +

57

.9

R2 =

0.6

76

9

0

10

20

30

40

50

60

30

40

50

60

70

80

90

Ni

Re

co

ve

ry (

%)

Cu Recovery (%)

Min

i p

ilot

test

resu

lts

Re

fere

nce

Te

st

18

Ph

ase

4 T

estw

ork

Lin

ea

r (M

ini p

ilot

test

resu

lts)

Sta

ge R

eco

very

of

Au

ag

ain

st

Sta

ge R

eco

very

of

Ni

0.0

0

10

.00

20

.00

30

.00

40

.00

50

.00

60

.00

70

.00

80

.00 0

.00

10

.00

20

.00

30

.00

40

.00

50

.00

60

.00

70

.00

80

.00

90

.00

Sta

ge

Re

co

ve

ry o

f N

i

Stage Recovery of Au

Min

ipilo

t te

st re

sults

Refe

rence T

est 18

Phase 4

results

Sta

ge R

eco

very

of

Pt

ag

ain

st

Sta

ge R

eco

very

of

Ni

0.0

0

10

.00

20

.00

30

.00

40

.00

50

.00

60

.00

70

.00

80

.00 4

0.0

04

5.0

05

0.0

05

5.0

06

0.0

06

5.0

07

0.0

07

5.0

08

0.0

08

5.0

09

0.0

0

Sta

ge

Re

co

ve

ry N

i (%

)

Stage Recovery Pt (%)

Min

i p

ilot

resu

lts

Re

fere

nce

Te

st

18

Ph

ase

4 r

esu

lts

Ni R

eco

very

ag

ain

st

Feed

Gra

de

y =

11

.76

6L

n(x

) +

79

.91

6

R2 =

0.9

78

0

10

20

30

40

50

60

70

80

00

.10

.20

.30

.40

.50

.60

.7

Ni

Fe

ed

Gra

de

(%

)

Ni Recovery (%)

Ph

ase

4 t

est

resu

lts

Lo

g.

(Ph

ase

4 t

est

resu

lts)

Phase 4 Nickel Recovery

Recovery Feed grade

Zero 0.001 0.001

Test 1 47 0.114

2 68 0.255

3 76 0.613

Minor recoveries of precious metals

Cu to

Mini pilot Pt recovery to Cu conc Au recovery to Ni conc Ni to Ni conc Cu to Cu conc Ni

2 13.8 27.1 64.6 62.5 25.8

4 8.8 17 56.3 37.4 42.8

6 10.5 7.6 53.4 42.1 24.5

7 9.1 7.2 57.6 44.1 26

8 19.6 7.5 67.8 74.7 10.7

9 13.7 18.4 62 63.2 13.7

Test 18 14.5 10.3 65.1 45.9 26.1

Phase 4 1 1.8 9.7 49.1 14.9 42.3

2 0.7 42.6 83 19.8 55.7

3 17.9 15.4 77.6 66.5 18.5

All recoveries are stated as a percentage.

Cobalt data

Feed grade (%) Wt (%) Metal units

0.014 100 1.4

Conc. Grade (%) Wt (%) Metal units Co Recovery (%)

0.425 1.9 0.8075 57.67857143

0.505 1.9 0.9595 68.53571429

Average feed grade used

Test Pd recovery (%) Ni Recovery (%)

2 39.6 64.6

4 27.1 56.3

6 19.2 53.4

7 17.3 57.6

8 24.6 67.8

9 25.6 62

Ref. 18 31.9 65.1

Phase 4 1 20.8 49.1

2 48.9 83

3 52.7 77.6

Regression Analysis

m c Equation r2#VALUE! #VALUE! Au vs Cu y=0.333x + 1.066 0.594917

#VALUE! #VALUE! Pt vs Ni y=1.325x - 52.402 0.56607

0.69348036 -27.86 Au vs Ni y=0.693x - 27.86 0.425493

0.29907824 -3.04958 Pt vs Cu y = 0.299x - 3.05 0.910528

1.431692436 -56.7902 Stage Pt vs Ni y=1.431x - 56.79 0.570056

0.719102273 -27.1506 Stage Au vs Ni y=0.719x - 27.15 0.432081

1.022464265 -34.3099 Pd vs Ni y=1.022x - 34.31 0.756071

Appendix 7A –

General Site Layout, Including Tailings and Waste Dump

Appendix 7B –

Tailings Pond Cross Section

LowS

ulphurw

asteR

ock

Inclination1:1.3

Kevitsa,C

rossS

ectionofT

ailingsD

am(D

ownstream

Method)

+0.00

LowS

ulphurW

asteR

ock

Drainage

Layer,1m

Inclination1:2

91m

3m2m

33m

Starter

Dyke

TillM

aterial 5m

5m

Coarse

Tailings

TillM

aterialor

+26.00

+7.00

Appendix

7B

Appendix 7C –

Process Water Balance

Kevitsa/W

aterbalance

Ore

Feed

Capacity

4.5M

t/a

Appendix

7C

Legend:1.

Copper

flotation.Ore

feedcapasity

4.5M

t/a.S

olidscontent35

weight-%

.Water

8.36M

m3/a,pH

11.2.

Copper

concentrate49000

ton/a.M

oisture9

%.W

ater4800

m3/a.

3.T

hickeneroverflow

4.72M

m3/a,pH

11.4.

Thickener

underflow.S

olidscontent55

weight-%

.W

ater3.63

Mm

3/a,pH11.

5.S

ulfideflotation.S

olidscontent35

weight-%

,pH5.

6.T

hickeneroverflow

4.60M

m3/a,pH

5.7.

Thickener

underflow.Low

sulfurtailing

4.00M

t/a.S

olidscontent52

weight-%

.Water

3.70M

m3/a,pH

5.8.

Nickelconcentrate

77000ton/a.M

oisture9

%.

Water

7600m

3/a.9.

High

sulfurtailing

0.40M

t/a.Solids

content35w

eight-%.W

ater0.75

Mm

3/a,pH11.

10.High

sulfurtailings

pond.Area

0.22km2.C

apacity6.0

Mt

tailings/15yars.S

olidscontent70

weight-%

.Slurry

volyme

4.4M

m3/15

years.0.5.7M

m3/a

water

reclaimed

backfrom

tailings.Neteffectofrain

andevaporation

0.08M

m3/a.

Totalw

aterreclaim

ed0.65

Mm

3/a.11.O

penpit.R

ainand

snow0.4

Mm

3/a.Ground

water

1M

m3/a.

12.Waste

rock(totalarea

2.2km

2),neteffectofrain,snow

andevaporation

0.8M

m3/a.

13.Lowsulfur

tailingspond.A

rea2

km2.C

apacity60.0

Mt

tailings/15years.S

olidscontent70

weight-%

.Slurry

volyme

43.7M

m3/15

years.1.98M

m3/a

water

reclaimed

backfrom

tailings.N

eteffectofrainand

evaporation0.7

Mm

3/a.T

otalwater

reclaimed

4.88M

m3/a.(including

openpitand

waste

rockpiles)

14.Recycle

water

pond,receiving5.53

Mm

3/a.15.F

reshw

aterfor

pumps

andfiltration

fromriver

Kitinen.C

apacity0.8

Mm

3/a.16.R

ecyclew

aterto

flotationprocess

3.64M

m3/a,pH

7.17.S

urplusw

aterpum

pedto

riverK

itinen1.89

Mm

3/a.

River

Kitinen

1

24

5

67

8

91011 12

13

1415

1617

3

Appendix 8A –

Road and Power Lines Layout

Appendix 9A –

ABA Testing

Sa

mp

le i

nfo

rmati

on

Su

lph

ur

con

cen

trat

ion

(%)

Aci

d

po

ten

tia

l,

AP

Neu

tra

liza

tio

n

po

ten

tia

l, N

P

NP

/AP

Cla

ssif

ica

tio

n

LV

T

nu

mb

er

KE

IVIT

SA

MIN

IPIL

OT

2

(kg

Ca

CO

3/t

)

(kg

Ca

CO

3/t

)

75

45

TE

ST

2,

23.6

.2005

klo

:13.3

0-1

7.0

0

0,6

62

0,6

62

,23

,02

NA

F

75

46

TE

ST

4,

28.6

.2005

klo

:12.1

0-1

4.0

0

1,2

53

9,1

62

,81

,61

PA

F

75

47

TE

ST

5,

28.6

.2005

klo

: 18.1

0-1

9.1

0

0,9

22

8,8

61

,42

,13

NA

F

75

48

TE

ST

6,

29.6

.2005

klo

: 12.1

5-1

5.0

0

1,7

25

3,8

59

,81

,11

PA

F

75

49

TE

ST

7,

29.6

.2005

klo

: 15.2

0-1

8.2

0

1,4

04

3,8

63

,71

,46

PA

F

75

50

TE

ST

8,

1.7

.2005

klo

: 9.0

0-1

2.0

0

1,0

23

1,9

61

,31

,92

PA

F

75

51

TE

ST

9,

1.7

.2005

klo

: 12.0

0-1

5.1

5

1,4

24

4,4

61

,21

,38

PA

F

AB

A r

esu

lts

(Aci

d p

ote

nti

al c

alcu

late

d w

ith S

-SO

4 =

0)

1.

NAF n

on a

cid f

orm

ing,

PAF p

ote

ntially

aci

d f

orm

ing

Su

lph

ur

con

cen

trati

on

(%)

Aci

d

pote

nti

al,

AP

Neu

trali

zati

on

pote

nti

al,

NP

NP

/AP

Cla

ssif

icati

on

LV

T

nu

mb

er

Hole

nu

mb

er

Fro

m

To

Sam

ple

nu

mb

er

(kg

CaC

O3/t

)

(kg C

aC

O3/t

)

730

KV

-411

13

327164

0,0

64

1,9

927,0

13,6

NA

F

731

KV

-435

37

327176

0,1

91

5,9

728,0

4,6

9N

AF

732

KV

-449

51

327183

0,4

01

12,5

27,8

2,2

2N

AF

733

KV

-94

6327438

0,1

47

4,5

933,4

7,2

7N

AF

734

KV

-944

46

327458

0,3

00

9,3

854,6

5,8

3N

AF

735

KV

-946

48

327459

0,4

80

15,0

27,2

1,8

1P

AF

736

KV

-948

50

327460

0,1

99

6,2

228,7

4,6

1N

AF

737

KV

-962

64

327467

0,4

19

13,1

40,5

3,1

0N

AF

738

KV

-964

66

327468

1,0

332,2

34,7

1,0

8P

AF

739

KV

-968

70

327470

0,8

07

25,2

36,9

1,4

6P

AF

740

KV

-11

28

30

230672

2,8

087,5

37,5

0,4

3P

AF

741

KV

-11

42

44

230679

0,9

88

30,9

37,2

1,2

1P

AF

742

KV

-11

46

48

230681

2,2

670,6

12,8

0,1

8P

AF

743

KV

-11

58

60

230687

0,4

83

15,1

55,9

3,7

0N

AF

744

KV

-11

66

68

230691

0,3

00

9,3

859,9

6,3

9N

AF

1.

NAF n

on a

cid form

ing,

PAF p

ote

ntially

aci

d form

ing

AB

A r

esult

s (A

cid p

ote

nti

al c

alcu

late

d w

ith S

-SO

4 =

0)

Sam

ple

in

form

ati

on

Su

lph

ur

con

cen

tra

tio

n

(%)

Aci

d

po

ten

tia

l,

AP

Neu

tra

liza

tion

po

ten

tia

l, N

P

NP

/AP

Cla

ssif

icati

on

1

LV

T

nu

mb

er

Hole

nu

mb

er

Fro

m

To

Ori

gg

eol

(kg

Ca

CO

3/t

)

(kg

Ca

CO

3/t

)

81

49

KV

-12

41

43

Met

aper

iod

ite

0,4

21

3,1

51

,23,9

0N

AF

81

50

KV

-12

46

45

Oli

vin

e p

yro

xen

ite

0,9

93

0,9

41

,01,3

2P

AF

81

51

KV

-16

104

106

Met

aper

iod

ite

0,8

92

7,8

57

,82,0

8N

AF

81

52

KV

-16

106

108

Met

aper

iod

ite

0,3

61

1,3

52

,34,6

5N

AF

81

53

KV

-16

108

110

Met

aper

iod

ite

0,5

21

6,3

80

,84,9

8N

AF

81

54

KV

-16

110

112

Met

aper

iod

ite

0,7

32

2,8

18

78,2

2N

AF

81

55

KV

-20

18

20

Oli

vin

e p

yro

xen

ite

0,3

81

1,9

49

,44,1

6N

AF

81

56

KV

-20

20

22

Oli

vin

e p

yro

xen

ite

1,3

84

3,1

52

,01,2

1P

AF

81

57

KV

-20

30

32

Oli

vin

e p

yro

xen

ite

0,1

13

,44

58

,216,9

NA

F

81

58

KV

-20

232

234

Met

aper

iod

ite

0,6

92

1,6

2,9

66,9

8N

AF

81

59

KV

-20

234

236

Met

aper

iod

ite

0,9

32

9,1

5,2

43,5

6N

AF

Sa

mp

le i

nfo

rmati

on

1.

NAF n

on a

cid f

orm

ing,

PAF p

ote

ntially

aci

d f

orm

ing

Appendix 9B –

Acid Drainage Testwork

TEST REPORT

7670-7672

1/3

Orderer: Scandinavian Gold Ltd

Name of the research: Kevitsa tailing samples from GTK Outokumpu testwork

Samples taken by : Customer/GTK Reijo Kalapudas

LVT sample marks: 7670-7672

1 SAMPLE INFORMATION

LVT number Sample type Sample Analyses 7670 Tailing sample Test 18/J Leaching- and ABA-test

7671 Tailing sample Test 19/J Leaching- and ABA-test

7672 Tailing sample Test 21/J Leaching- and ABA-test

2 APPLIED METHODS

2.1 Leaching Test

Analysis performed by standard SFS-EN 12457-3. This standard could be used to test leachibility in

wastes and sludges. The sample particle size must be at least 95% (mass) less than 4mm, sample could

be crushed if needed.

The sample material was brought into contact with deionized water. In the first leaching step sample

was agitated 6 h ± 0,5 h with liquid/solid ratio L/S=2 (2 l/kg) and in the second leaching step 18 h ±

0,5 h with liquid/solid ratio L/S=8 (8 l/kg). The sample will be filtered and centrifuged if needed and

after that appropriate parameters will be chemically analysed. Every time the sample container is

opened, any gas development from the sample is observed and booked. Blank samples area treated

together with actual test samples. Sample capacity in the first shake is 0,44 l and in the second shake

1,50 l.

Test results interpretation includes comparison to the European Union Board decision about leachate

quality limits (directive 1999/31/EY, 16 article and enclosure II ‘basis and procedures to accept waste

into a landfill’).

2.2 ABA Test

2.2.1 Neutralization Potential (NP)

Testing is performed according to standard ABA testing procedure (Geochemical test procedures for

field reconnaissance programs-modified ABA, Robertson GeoConsultants Inc.). In the test, pulverized

(particle size in the sample must be >80 % <0,250 mm) sample is precisely weighted about 2 g and the

sample will be mixed in room temperature with hydrochloric acid for 24 hours. Mixture will be

titrated with NaOH-solution to pH 8,3 and the neutralization potential (kg CaCO3/t) will be counted

from the results. The concentration of hydrochloric acid needed in this test is solved by an advanced

test. In the advanced test few drops of 25% hydrochloric acid is dropped onto the sample material (1-

2 g) and the degree of reaction is evaluated.

TEST REPORT

7670-7672

2/3

2.2.2 Calculation of Acid Potential (AP)

Acid capacity is calculated from the ABA test results. Sulphide-sulphur content calculation:

Sulphide-sulphur = Stot – SSO4

Acid potential (kg CaCO3/t) is calculated by multiplying sulphide concentration (%) with 31,25.

2.2.3 Classification of Materials

Materials ability to form acid leachate can be identified with following criteria:

Non-Acid Forming (NAF) material:

- sulphur concentration S < 0,3 %

- if S > 0,3 %, NP/AP > 2

Potentially-Acid Forming (PAF):

- S > 0,3 %, NP/AP < 2

3 ANALYSATION AND RESULTS

3.1 Solubility Test

In the first leaching step sample was agitated 6 h ± 0,5 h with liquid/solid ratio L/S=2 (2 l/kg) and in

the second leaching step 18 h ± 0,5 h with liquid/solid ratio L/S=8 (8 l/kg). Cumulative release of

elements with liquid/solid ratio L/S=10 was calculated by methods from standard SFS-EN 12 457-3.

Immediately after the sample was filtered, the temperature of air and liquid pH, temperature and

conductivity (tables 1 – 2) was measured. During shaking, the sample container was not opened.

Soluble chemical elements analysed are on tables 3-5. Results are compared to leachability limit

values, which are applied in placing waste in landfills (tables 4 and 5).

Table 1. Blank test’s parameters that were analysed right after agitation and filtering..

Parameter Blind sample

1. Eluate

Blind sample

2. Eluate

Temperature (°C) 21,6 21,0

pH 6,22 6,45

Conductivity (mS/m) 0,4 0,2

Air temperature (°C) 24,2 22,9

Table 2. Measured parameters right after agitation and filtering.

Parameter 7670

1. Eluate

7670

2. Eluate

7671

1. Eluate

7671

2. Eluate

7672

1. Eluate

7672

2. Eluate

Temperature (°C) 20,5 19,0 20,7 20,4 20,5 19,0

pH 7,69 8,38 7,68 8,46 8,52 9,24

Conductivity (mS/m) 53,5 8,1 50,0 7,7 19,0 6,1

Air temperature (°C) 24,4 22,8 24,4 22,8 24,2 22,9

TEST REPORT

7670-7672

3/3

Table 3.Concentrations of blind samples.

As

(µg/l)

Cd

(µg/l)

Cr

(µg/l)

Cu

(µg/l)

Hg

(µg/l)

Ni

(µg/l)

Pb

(µg/l)

Sb

(µg/l)

Zn

(µg/l)

1. Eluate < 1,0 < 1,0 < 1,0 < 5,0 < 0,1 < 1,5 < 1,0 < 5,0 < 15

2. Eluate < 1,0 < 1,0 < 1,0 < 5,0 < 0,1 < 1,5 < 1,0 < 5,0 < 15

Table 4.Concentrations of dissolved materials at first leaching step (L/S = 2).

As

(µg/kg)

Cd

(µg/kg)

Cr

(µg/kg)

Cu

(µg/kg)

Hg

(µg/kg)

Ni

(µg/kg)

Pb

(µg/kg)

Sb

(µg/kg)

Zn

(µg/kg)

7670 < 2 < 2 5,60 < 10 < 0,2 50,0 < 2 < 10 < 30

7671 < 2 < 2 4,20 < 10 < 0,2 50,0 < 2 < 10 < 30

7672 6,00 < 2 12,6 < 10 < 0,2 54,0 < 2 < 10 < 30

Guide line (L/S=2) 100 30 200 900 3 200 200 20 2 000

Table 5.Concentrations of cumulative dissolved materials with solution solids relation L/S = 10.

As

(µg/kg)

Cd

(µg/kg)

Cr

(µg/kg)

Cu

(µg/kg)

Hg

(µg/kg)

Ni

(µg/kg)

Pb

(µg/kg)

Sb

(µg/kg)

Zn

(µg/kg)

7670 < 10 < 10 28,8 < 50 < 1 142 < 10 < 50 < 150

7671 < 10 < 10 27,7 < 50 < 1 142 < 10 < 50 < 150

7672 21,7 < 10 48,8 < 50 < 1 212 < 10 < 50 < 150

Guide line (L/S=10) 500 40 500 2 000 10 400 500 60 4 000

3.2 ABA-test

Table 6. ABA- test results.

Sulphur

concentration (%)

Acid potential, AP

(kg CaCO3/t)

Neutralization potential, NP

(kg CaCO3/t)

NP/AP Classification

7670 0,13 3,88 30,3 7,8 NAF

7671 0,14 4,20 29,5 7,0 NAF

7672 0,20 6,07 32,5 5,3 NAF

4 CONCLUSIONS

Kevitsa tailings samples acid-base accounting gave materials classification ‘non-acid forming’ (NAF).

Tailings leachability testing according to method SFS-EN 12457-3 gave low leachate concentrations

and hence classification to landfill waste class (directive 1999/31/EY, article 16 and enclosure II ‘basis

and procedures to accept waste into a landfill’).

Rovaniemi 14.2.2005

LAPIN VESITUTKIMUS OY

Tarja Olli

M.Sc. (chemistry)

Appendix 11A –

Specifications of Nickel and Copper Concentrates

Specifications of Nickel and Copper Concentrates Produced from Keivitsa Ore / Original Version / Based on Mini Pilot Test

Ni Concentrate:

Element Min Max TypicalNi (%) 9.0 13.0 12.0Cu (%) 2.0 6.0 2.9 Co (%) 0.4 0.5 0.5 S (%) 27.0 30.0 29.1Fe (%) 35.0 37.0 36.4Pd (g/t) 2.0 2.5 2.3 Pt (g/t) 1.7 2.4 1.9 Au (g/t) 0.5 1.0 0.6 MgO (%) 5.5 7.0 6.2

Cu Concentrate:

Element Min Max TypicalNi (%) 0.4 0.9 0.6 Cu (%) 22.5 26.5 24.5Co (%) 0.03 0.04 0.03S (%) 28.0 30.0 29.3Fe (%) 26.5 27.5 26.9Pd (g/t) 1.7 2.1 1.9 Pt (g/t) 2.5 3.0 2.6 Au (g/t) 2.0 3.5 2.3 MgO (%) 4.0 6.0 5.2

Appendix 11B –

Nickel and Copper Marketing Report

Prepared for:

Scandinavian Gold

Report E705-1

By

S•t B•a•r•b•a•r•a Consultancy Services 9 John Street London

WC1N 2ES United Kingdom

Keivitsa Nickel and Copper Concentrates

Marketing Study

3rd February 2006

St Barbara Consultancy Services E705 – Ni/Cu Pre-Fease Marketing Study

Table of Contents 1. INTRODUCTION ......................................................................................... 1

2. THE AUTHOR ............................................................................................. 2

3. METHODOLOGY.......................................................................................... 3

4. SMELTERS AND TRADERS CONTACTED.......................................................... 4

5. COPPER CONCENTRATES............................................................................. 5

5.1. The Material ........................................................................................... 5

5.2. Delivery of Concentrates to the Smelters .................................................... 6

5.3. Responses from the Smelters and Traders .................................................. 6

5.4. Terms and Conditions of Treatment ........................................................... 7

5.4.1. Delivery .......................................................................................... 7

5.4.2. Prices ............................................................................................. 7

5.4.3. Quotational Periods........................................................................... 7

5.4.4. Payment ......................................................................................... 7

5.5. Trader terms and conditions ..................................................................... 8

5.6. Comparison of the Treatment and Refining Charges at Various Prices ............. 9

6. NICKEL CONCENTRATES............................................................................ 10

6.1. Material ............................................................................................... 10

6.2. Terms and Conditions of Treatment ......................................................... 10

6.2.1. Delivery ........................................................................................ 10

6.2.2. Payments ...................................................................................... 10

6.2.3. Deductions .................................................................................... 11

6.2.4. Prices ........................................................................................... 11

6.2.5. Quotational Period .......................................................................... 11

6.2.6. Payment ....................................................................................... 11

6.2.7. Valuation of the contracts ................................................................ 12

7. CONCLUSIONS......................................................................................... 13


1. INTRODUCTION

The author, John Shaw, was requested to investigate the market for copper and nickel concentrates on behalf of Scandinavian Gold (the company) from their Keivitsa project in Finland.

The study was designed to determine the likely level of interest in the copper concentrates to be produced by Scandinavian Gold from 2008/9, when the mine will come on stream.

Scandinavian Gold had already spoken to New Boliden in Sweden and ascertained that they had an interest in both the copper and nickel concentrates and provided an indication of terms to the company.


2. THE AUTHOR

John Shaw is a graduate of Queen’s University, Belfast. He has over 30 years experience in the metals business, specialising in non-ferrous metals raw materials including, copper, nickel, lead and zinc concentrates and ores and blister copper trading.

His career has encompassed Amalgamated Metal Corp, Pechiney World trade’s Brandeis Division, where he was a director for 8 years and Head of Raw Materials and Metallgesellschaft where he was a member of the department handling over one million tons of concentrates per annum. He was responsible for the development of their business in Eastern Europe, African and certain Asian countries.

He has very extensive knowledge of international trading and the flow of materials from producers to consumers and the development of new and expanding markets especially those in the Far East and China. He has dealt with the major producers in all the leading producing countries, including those in North and South America.

He has worked closely with both producers and consumers to assist them in the development of new markets and increase market share. He has extensive knowledge of the establishment of long term and frame contracts and of risk management and long-term price management, as well as logistics.


3. METHODOLOGY

A number of potential purchasers were contacted and were provided with the indicative assays of the concentrates as determined by the testwork programme. Each company was asked to give an expression of interest to purchase the material, based on these assays, and their ability to smelt or trade them within their trading book.

Since production is some years in the future, the respondents were not asked to give firm indication of terms, but to provide where possible good indicative terms based on the current market terms and conditions. Their view of how these would change in the period leading up to the start of production was also sought.

Given the geographical location of the mine, it was agreed with the company that the smelter market in Europe (including Russia) for the copper concentrates would be the main focus, together with those trading companies which are known to have an active business in Europe.

It was not considered worthwhile to go outside Europe because of the inevitable high freight charges which would off-set any advantage that might accrue in the terms and conditions achieved. This proposition will be re-visited in detail at the feasibility stage, when a comprehensive study of the freight market will take place.

The approach to Russian smelters was directly through their representative offices in London and through trading companies known to have a significant business in copper and nickel in that country.

The response from the Russian smelters was poor and despite registering initial interest. Further contact is planned at the next major stage of discussions when it is hoped that they will more responsive.

A trading group with previous business relations and smelter investments in Russia has expressed and interest and discussions will be taking place with them in the near future to determine their interest for delivery to Russia and obtain an indication of terms.


4. SMELTERS AND TRADERS CONTACTED

The following contacts were selected for the study;

Smelters

Norddeutsche Affinerie, Hamburg, Germany Atlantic Copper, Huelva, Spain KGHM, Poland Norlisk, Russia Jinchaun Group Ltd. China OMG Finland Oy INCO Ltd

Traders

Sempra Metal and Commodity Corp, USA Trafigura Beheer B.V., Switzerland Marc Rich Investments, Switzerland Glencore A.G. Switzerland Traxys Belgium S.A.

Other traders were originally considered including Ocean Partners, Transamine and Louis Dreyfus, but it was felt that they would not offer at this stage anything different to those already on the list. If necessary we can include these companies in the future discussions.


5. COPPER CONCENTRATES

5.1. The Material

The copper concentrate specification as provided from the testwork by Jukka Karhunen of the Geological Survey of Finland is shown in the following table:

Metal %/gpt

Nickel 0.6

Copper 24.5

Sulphur 29.3

Iron 26.9

Palladium 1.9

Platinum 2.6

Gold 2.3

Magnesia 5.2

The copper concentrates are considered low to medium grade in terms of copper content, by the smelters, although this is not handicap in terms of selling them as all the European smelters consume a range of concentrates from the mid twenties up to 45%Cu. The quantity of very high grade concentrates available has fallen in recent years as the grade of Escondida (Chile) falls. The Keivitsa concentrates should provide a good blending material for the smelters.

The gold content, 2.3gpt, is well above the payment threshold for the smelters who would usually insist on a 1g deduction per ton of concentrates. At current prices, ($550/oz), a total payment of 1.3g would be worth $23/t of concentrates. However, applying a typical gold payment of 90% of the full content from a Far East smelter would make the content worth $40/mt of concentrates.

As a consequence, the benefit of the lower freight rates to European smelters can be off-set by the much higher gold payment in the Far East. A similar situation applies to the silver payment.

In Europe the total payment after deducting 30g and assuming a silver price of $8/oz, is $2/mt. In the Far East, where the payment is 90% of the full content, the value would be $10/mt.

The total difference in payment for these two elements is $40/mt of concentrates. This would be more than enough to cover the additional freight.

The platinum and palladium are assessed to be too low to be payable in today’s market. The smelters will no doubt recover them, but the standard deduction for each is likely to be around 5 g/mt, although with the prices of these metals at high levels, there is a case to be made for insisting on a lower deduction.


Impurity levels in general were acceptable for all the smelters, with the exception of the Nickel content. This is considered somewhat high in today’s markets and will likely be penalised in those smelters which will accept it. A penalty of $2.00 per 0.1% Ni above 0.4% is assessed.

Otherwise, the levels of impurities are not expected to be a drawback in marketing the material.

5.2. Delivery of Concentrates to the Smelters

The location of the mine in Finland gives rise to a number of delivery options. It would be expected that tonnage being delivered to the New Boliden smelters in Sweden and Finland would be delivered by rail or road transport. This facilitates deliveries in small and regular lots with a potential benefit to the cash flow and stock management. It is also likely to be cheaper as the handling is reduced with fewer losses. It would, of course, also be possible to deliver to their Swedish smelter by coaster.

Deliveries to other smelters in Europe would be by ocean going vessel. As the annual production is not large, it is anticipated that the shipment lots would be in the range of 3,000 to 5,000wmt.

Freight rates have in general fallen in recent months from the high levels of recent years and current indications of freight rates are to Hamburg $15-$20/wmt and to Huelva about $30/wmt.

Since the Russians have not so far indicated any interest, means or the cost of transportation to their smelters has not been investigated.

Deliveries to the Far East would be around the $60/wmt level. Currently shipments from Poland of concentrates in 2,500wmt lots are in the range of $52 - $55/wmt.

5.3. Responses from the Smelters and Traders

The overall response from the smelters was favourable, although this has to be seen in the context of the time frame of the project and their other long-term supply goals.

Norddeutsche Affinerie and Atlantic Copper both said that they would be interested in the material, but would likely take only a portion of the annual production. This is really a function of the copper content of the concentrates, which is low by current international standards. The smelters are looking for material closer to 30% copper content. Tonnage was not discussed with New Boliden as they had already been approached by Scandinavian Gold and had submitted an indication of terms.

KGHM Poland also expressed a concern about the copper content principally because their own concentrate is low grade, and they will be looking for higher grade concentrates when their new smelter is commissioned. Their precious metals recoveries are also not good.

Chinese and other Far East smelters can also take the material, and as indicated above, at current high prices, some of the additional freight costs in transporting the material to China would be off-set by the higher precious metals payments terms. It is proposed to investigate this market in greater depth during the next phase of the feasibility study.

Russian smelters have not responded so far, for either copper or nickel and have not been included as a potential outlet at this stage.


5.4. Terms and Conditions of Treatment

The term and conditions of treatment of the copper concentrates for a new mine coming on stream in 2009 cannot be forecast with any great degree of accuracy, in terms of the actual numbers for the important conditions such as the treatment and refining charges.

However, for a long-term contract, the so called Producer-priced contracts between the major smelters and the largest miners, Escondida, Los Pelambres and Collahausi have become well publicised and are becoming a “benchmark” for the trade. Applying these to the current market in Europe, the terms and conditions are confirmed to be similar to the following.

5.4.1. Delivery

CIF Free out, receiver’s port.

5.4.2. Prices

The total price would be the sum of the payable elements less the treatment and refining charges and any price participation.

Copper: Pay for 96.50% of the full copper content subject to a minimum deduction of 1.0 unit at the London Metal Exchange quotation for Grade A copper averaged over the quotational period.

Silver: Pay for 90% of the silver content, subject to a minimum deduction of 30 g/t of concentrates at the London Bullion Brokers fine silver dollar quotation averaged over the quotational period.

Gold: Pay for 90% of the gold content subject to a minimum deduction of 1.0 g/t of concentrates at the mean of the London morning and afternoon quotation averaged over the quotational period.

Deductions:-

Treatment Charge: US$ 95.00/t of concentrates Refining Charge: US$ 0.095/lb of copper paid for. US$ 0.35/oz of silver paid for.

US$ 6.00 per ounce of gold paid for.

Price Participation For each US$ 0.01 the price of copper exceeds US$0.9/lb, the buyer will participate in 10% of the difference between US$0.9 and the actual price established during the quotational period per pound of copper paid for.

5.4.3. Quotational Periods

Subject to negotiation, the quotational period is likely to be one of the following

1) The month after the month of arrival of vessel at the port of discharge 2) The second month after the month of arrival of the vessel at the port of discharge. 3) The third month after the month of arrival of the vessel at the port of discharge.

5.4.4. Payment


All payments in US Dollars, 90% provisionally, would be made approximately 30 days after the arrival of the vessel at the port of discharge.

The balance of 10% would be paid after all the cargo details, relating to the final weight, final assays, and final prices are established.

In China and other Far East destinations, we would expect to see a variation in these terms, principally to the precious metals payments. These would typically be:

Silver: If the silver content is over 30g/mt of concentrates, buyer will pay for 90% of the full content at the London Bullion Brokers fine silver dollar quotation averaged over the quotational period.

Gold: If the gold content is over 1.00 g/mt of concentrates buyer will pay for 90% of the full content at the mean of the London morning and afternoon quotation averaged over the quotational period.

The combined difference of these two payment conditions could be up to US$40/mt of concentrates if the gold and silver prices remain close to current levels.

This difference could bring the Far East market into consideration for future deliveries, as this difference would compensate for the higher freight charges.

5.5. Trader terms and conditions

All of the traders approached expressed and interest in the material, they all had in common existing business with the smelters in the region and the Far East and it was felt that they would be the most interested and competitive given the opportunities to trade the concentrates in Europe.

Their views on the terms and conditions were very similar to the smelters and they also would be looking towards the published “benchmark” terms for a long-term business.

However, Glencore, were somewhat more aggressive and indicated a treatment and refining charge for the “life of mine” of above US$80.00/mt of concentrates and US$0.8/lb of copper paid for. Although they did not put a figure on how much above this level, one can assume that it would be close to, or just below, the “benchmark” terms.

An alternative proposal to this was to pay for 17.5% of the prevailing copper price. No reference to made to silver and gold payment, but assume they are not paid for.

This percentage may look low but a comparison with the terms which would be achieved on the current producer contracts demonstrates that this is not necessarily the case. (See the table for the percentages which would be achievable by this method).

At a price of $5,000/mt the payment to Kevitsa would be about 19.2%, falling to 18.7% at $4,000 and 17.8% at $3,000. Therefore the Glencore proposal is very favourable when the price exceeds $3,000, but below this level it is actually in the miner’s favour. These percentages are before any precious metals credits, which would make only a very marginal difference in Europe.

While it may be tempting to have a single percentage payment for the concentrates, this would be a big gamble for a long-term contract. It is possible that if the copper price remains at close to current levels for an extended period, that the price participation threshold could be increased, thus giving the miner a larger percentage and monetary return.


5.6. Comparison of the Treatment and Refining Charges at Various Prices

The main parameters used are summarised as follows;

Treatment Charge (TC) 95.00

Refining Charge (RC) 0.095

Price Participation (PP) 10% of the price above $0.90/lb Cu

Copper content 24.5%

Copper payment 23.5% equivalent

Payments receivable at varying copper prices are shown in the following table:

$/mtCu

CuContent

TC RC TC+RC PP Total Deductions

Net $ per

mt Cu

%Payment

of Cu price

5,000 1,175 95 9.5 114.2 71 215.2 959.8 19.2

4,000 940 95 9.5 114.2 47 191.2 748.8 18.7

3,000 705 95 9.5 114.2 24 168.2 536.8 17.8

2,000 470 95 9.5 114.2 0.4 144.6 325.4 16.3

1,500 352 95 9.5 114.2 0.0 144.2 208.3 13.9


6. NICKEL CONCENTRATES

6.1. Material

The nickel concentrate grade is:

Metal %/gpt

Nickel 12.0

Copper 2.9

Sulphur 29.1

Iron 36.4

Palladium 2.3

Platinum 1.9

Gold 0.6

Magnesia 6.2

A number of major Nickel producers were contacted in Europe and the Far East, including OMG, Jinchuan, INCO and Norilsk. The first three have submitted an indication of terms.

The other three companies who have submitted terms are geographically widely separated and they represent an independent view of the global market at this time.

The material did not raise any problems for treatment in the smelters and the small but significant levels of precious metals were just above the thresholds for payment. The terms do not include any hidden penalties for impurities which are not specifically stated.

6.2. Terms and Conditions of Treatment

Unlike the copper concentrates, the terms for the nickel concentrates are more detailed and reflect the current levels obtainable.

A brief summary of the terms from OMG show the structure.

6.2.1. Delivery

Delivery by rail or road DDP Harjavalta Works, unloaded (Incoterms 2000). 6.2.2. Payments

Nickel 92% of the assayed nickel content in the Material, subject to a minimum deduction of 1.2 units.

Copper 92% of the assayed copper content in the Material, subject to a minimum deduction of 0.2 units.

Cobalt 45% of the assayed cobalt content in the Material, subject to a minimum deduction of 0.2 units.

Gold 50% of the assayed gold content in the Material, subject to a minimum deduction of 1 g/t.


Platinum 50% of the assayed platinum content in the Material, subject to a minimum deduction of 1.5 g/t.

Palladium 50% of the assayed palladium content in the Material, subject to a minimum deduction of 1.5 g/t.

6.2.3. Deductions

Treatment charges:

Nickel greater of 27% of nickel price or US$1.0/lb accountable nickel

Copper greater of 30% of copper price or US$0.30 accountable copper

Cobalt greater of 35% of cobalt price or US$3.5/lb accountable cobalt

Gold USD15/tr oz accountable gold.

Platinum USD15/oz accountable platinum.

Palladium USD15/tr oz accountable palladium.

6.2.4. Prices

Nickel Price of nickel shall be the lower of either (i) LME settlement or (ii) LME 3 month price quotation averaged over the quotational period, as published in Metal Bulletin.

Copper Price of copper shall be the lower of either (i) LME Grade Settlement or (ii) LME 3 month price quotation averaged over the quotational period, as published in Metal Bulletin.

Cobalt Price of cobalt shall be the Metal Bulletin Free Market (99.3 %); low, quotation averaged over the quotational period, as published in Metal Bulletin.

Gold Price of gold shall be the London AM Fix in USD averaged over the quotational period.

Platinum Price of platinum shall be the mean of London AM and PM Fixings less US$2/tr oz averaged over the quotational period.

Palladium Price of palladium shall be the mean of London AM and PM Fixings less US$4/tr oz averaged over the quotational period.

6.2.5. Quotational Period

The quotational period for nickel, copper and cobalt shall be the month following the month of arrival and Au, Pt and Pd shall be the 5th month after the month of arrival.

6.2.6. Payment

Payment of base metals shall be made in USD, 90% of provisional value less charges and penalties within 65 days after the arrival. PGM payments will be made after the quotational periods balance when all the facts are known.


6.2.7. Valuation of the contracts

Applying the following prices for each of the payable metals

Nickel $5.25/lb Copper $1.9725/lb Cobalt $13.00/lb Gold $478/tr oz Platinum $990/tr oz Palladium $258/tr oz

The payable metal contents come to a total of $1,449.03/mt of concentrate (compared with metal content valuations without any deductions of $1,746), and the deductions to $395.45; giving a net figure of $1,053.57. This net figure is approximately 72% of the payable metal contents, but only 62 % of the full metal contents.

By using the lowest of the published prices for the major metals, the percentage would in certain circumstance be reduced by approximately a further 1%.

Percentage payments by INCO appear to be up to 5% higher but after taking into consideration the higher freight costs the improvement is negated and the terms are little better than those paid for in Europe.

Chinese terms are significantly worse as they do not pay for any of the copper or the precious metals.

It is felt that although these terms are in line with current market conditions in the broadest sense, they reflect the fact that the Kevitsa is still a project and the timing of the production is uncertain. Therefore the smelters are providing terms which are very much their “wish list” and will probably not truly reflect the final terms. It is anticipated that the terms in the final contracts would be significantly better than those quoted.

The use of percentage deductions and treatment charges based on the quoted prices for each metal means that in the current climate with high prices for all the metals, the smelter is getting a significant additional benefit which is not a reflection of their true treatment costs.


7. CONCLUSIONS

The nickel concentrates is expected to find a ready market in Europe and Canada, although China at this stage looks to be uncompetitive due to lack of payments for the minor metals and the potentially higher freight costs.

Sales to Russia remain a possibility, although their lack of response is probably a reflection of their lack of experience in dealing with non-Russian materials.

The proposed copper concentrates production from Kevitsa would certainly be able to find a market. In Europe demand remain very strong for concentrates as a number of expansions are planned by Norddeutsche Affinerie, KGHM, and Atlantic Copper. The un-tapped Russian market is still a possibility.

However, the Far East market has become a possible option in recent months as the precious prices have strengthened and freight costs have fallen. From a position of not expecting their smelters to be competitive, there is a real possibility of them being in a position to purchase some of the material. Much will depend on the prices of precious metals remaining at high levels.

In summary the Kevitsa concentrates will find a ready market either at a smelter close to the mine or at a number of overseas smelters. Quality of either material is not a major issue, although the nickel in the copper concentrates should be kept at the lowest possible levels to avoid penalties.

Appendix 11C –

Cobalt Industry Review

St Barbara Consultancy Services – Cobalt Review

COBALT

AN INDUSTRY OVERVIEW

St Barbara Consultancy Services

February 2006

St Barbara Consultancy Services – Cobalt Review 2

1 INTRODUCTION

Cobalt was first found as the mineral cobaltite (a cobalt arsenide). The name cobalt is derived from the old German “Kobold”, a name given to a mythological creature that resembled a goblin and from the Greek “cobalos”, meaning mine. It was given due to the mischievous habits of goblins and the fact that early processing techniques when applied to the mineral could not produce any cobalt metal, but produced a strong garlic type smell from the arsenic compounds. Cobaltite was finally separated into its pure form by Georg Brandt in 1735.

The first recorded mining of cobalt took place in the 15th century with the working of the Schneeberg Saxony silver-cobalt deposit in Germany. However, use of cobalt has been dated back to production of blue glass jewellery in Persia in 2250BC.

The main use of cobalt was as a colouring agent until the early 20th century and it was generally only available in the oxide form. Cobalt metal was first produced at the start of the 20th century, and its first usage was in a steel alloy.

There are now several different uses for cobalt, the principal ones being:

Metallurgical Magnetic Alloys Chemicals Cemented Carbides Bonded DiamondsElectronicsCeramics & Enamels

The main properties of cobalt are shown in table 1.1.

Table 1.1 Main properties of cobalt Property Value Colour Grey Atomic Number 27 Atomic weight 58.933 Melting point 1493ºC (2719ºF) Curie point 1121ºC Modulus of elasticity 2.1 x 106 kg/cm2

Specific gravity 8.85g/cm3

Electrical resistivity 6.24µ /cm at 20ºC Valencies 2 + 3 Residual Magnetisation 4900 Gauss (0.49T)

Cobalt is not considered to be a rare element, as it is the 33rd most abundant metal in the earth’s crust. The majority of cobalt produced is as by-product of copper and nickel processing, although some is contained with arsenic or silver. The most common cobalt minerals are cobaltite (CoAs2), erythrite (Co3(AsO4)2.8H2O), glaucodot ((Co,Fe)AsS) and skutterudite ((Co,Fe,Ni)As2-3).


Annual cobalt production between 1998 and 2004 is shown in table 1.2.

Table 1.2 Annual Cobalt Production since 19981

1998 1999 2000 2001 2002 2003 2004 Total Tonnage

34,333 35,073 38,703 39,974 41,213 44,895 49,536

Shedd2 has reported worldwide production in 2003 as 48,400 tonnes, 52,400 tonnes for 2004 and predicts the same amount for 2005. A half yearly estimate by CoDI3 indicates that refined cobalt availability for 2005 will exceed 52,000 tonnes, in agreement with Shedd.

The high increase in cobalt production and usage has had particular impetus from recent developments in rechargeable battery technology and an increased demand in goods using this technology. It is estimated that the use of cobalt in batteries now accounts for a fifth of worldwide cobalt usage. Growth in demand has been predominantly in the chemicals sector which is expected to have accounted for about 52% of total world demand in 2005.

Most of the cobalt produced comes from copper cobalt sulphides, though increasing quantities are being extracted from nickel laterite deposits.

1 Anon, 2004. Cobalt Facts – Supply and Demand, The Cobalt Development Institute. Taken from http://www.thecdi.com/cobaltfacts/documents/COBALT_FACTS-Supply_and_demand.pdf2 Shedd, K, January 2005 and 2006. USGS Mineral Commodity Summaries. Taken fromhttp://minerals.er.usgs.gov/minerals/pubs/commodity/cobalt/cobalmcs05.pdf3 Anon, 2006, Outlook for the Cobalt Market 2005, Cobalt News 06/1, 6-10


2 COBALT PROCESSING TECHNOLOGY

As indicated previously, cobalt is not an easy metal to recover, being inevitably found in association with other metals. The continuing development of improved techniques has played a key role in increasing supply.

Cobalt is produced commercially as a by-product from the processing of nickel and copper/nickel sulphide ores and from the processing of copper-cobalt ores of the African copper belt in Zambia and in the Democratic Republic of Congo. Cobalt is also produced from lateritic nickel deposits where hydrometallurgical methods of treatment are used rather than smelting to ferronickel, in which process the cobalt is lost.

2.1 Cobalt Recovery in Pyrometallurgical Processes

Primary cobalt metal is not produced directly by pyrometallurgical means. In nickel or copper/nickel smelting it is concentrated into a matte. Thus it is subject to losses in the discard slag.

Unlike copper sulphide concentrate smelting, wherein approximately one tonne of slag is produced per tonne of copper, in copper/nickel/cobalt smelting about 5 tonnes of slag arises per tonne of Cu + Ni + Co produced. Therefore slag losses are much more important in nickel sulphide smelting than in copper smelting. Furthermore, while the product of copper smelting and converting is blister copper, in nickel smelting the product is a nickel matte. Because of these high slag volumes and because cobalt partition between matte and slag is at least a factor of ten less than nickel, cobalt recovery to the matte is at best 50%, so this route is not geared for high cobalt recoveries.

However Outokumpu Oy has recently developed the DON (Direct Outokumpu Nickel) flash smelting process in which cobalt recovery is now said to be >75%. In this process the flash furnace produces a matte (FSF matte) with 70 % Ni + Cu, compared with 53 % in the old flash smelting process. There is no converting step as in the old flash smelting process. This process was developed to allow treatment of high magnesia ore from Mt Keith in Western Australia. Slag from the flash furnace is treated in an electric furnace (EF) to produce a waste slag and a further matte (EF matte) product. Both these mattes and mattes produced by more traditional smelting processes for copper-nickel-cobalt concentrates are then treated by hydrometallurgy for recovery of copper nickel and cobalt.

In nickel laterite smelting to ferronickel, cobalt is not recovered. Where matte is the product in laterite smelting, such as in New Caledonia (in part), by PT INCO (Indonesia) and by Niquel Tocantins in Brazil, then cobalt partitions to the matte, but slag losses are high. The matte is then treated hydrometallurgically for both cobalt and nickel recovery.

2.2 Cobalt Recovery by Hydrometallurgy

Only one process has so far been used for direct hydrometallurgical processing of nickel-cobalt or copper-cobalt-nickel sulphide concentrates and that is the Sherritt Gordon process. This ammoniacal pressure leaching process is now used not only by Sherritt at Fort Saskatchewan, but also by Western Mining for matte refining at their Kwinana refinery in Western Australia. The process, while still operating at Fort Saskatchewan, is used nowadays only on mattes and particularly on nickel/cobalt mixed sulphide precipitates from the nickel laterite processing plant at


Moa Bay in Cuba. Ammoniacal leach liquors require little purification, apart from copper removal, if present.

Matte is also treated commercially by either acid chloride or acid sulphate leaching. Acidic chloride and sulphate leach liquors, in comparison to ammoniacal liquors, solubilise varying amounts of impurities, mainly iron, together with copper and zinc. Aluminium, chromium, manganese, magnesium, calcium etc. are not usually a problem in matte processing, even with mattes produced from laterite smelting, as these elements will report to the smelter slag.

Thus Falconbridge use their proprietary chlorine leach process at Kristiansand in Norway. Eramet at Sandouville in France and Sumitomo in Japan also use a chloride leach while Outokumpu Oy in Finland and Rustenburg Refiners in South Africa use sulphate based leaching processes. The treatment of the Jinchuan smelter high grade matte also uses a sulphate based leach process not unlike the Outokumpu process.

As laterites cannot be usefully concentrated by physical means, treatment of the ore directly is required. Two process routes exist, namely the Caron process (reduction roast followed by ammoniacal leaching) and acid pressure leaching (the Moa Bay process). A feature of acid pressure leaching of laterites is the production of leach liquors which contain nickel and particularly cobalt at low concentrations which can present processing problems.

2.2.1 Cobalt Recovery from Copper-Cobalt Concentrates

The copper-cobalt concentrates of interest here are those from the African copper belt. These concentrates are not smelted but roasted and in some instances combined with oxidic material. The roaster calcine is acid leached in spent copper tankhouse electrolyte after topping up with acid and, after solid/liquid separation, the copper is recovered by electrowinning. A bleed from the spent electrolyte is electrostripped and cobalt produced by electrowinning after a series of hydrolytic precipitative purification steps to remove iron, aluminium and copper. Finally cobalt is precipitated and re-dissolved in spent cobalt tankhouse electrolyte and the cobalt recovered by electrowinning. The final electrolytic cobalt product is often not particularly pure and cobalt losses in the various precipitation steps are significant.

2.2.2 Sulphide Concentrate Leaching, Matte Leaching and Solution Purification

Apart from the development of the Activox process for pressure leaching of Cu/Ni/Co sulphide concentrates, there have been no further advances on the chemical leaching processes discussed above. Commercial development of this process was originally anticipated for the Yakabindie project in Australia, but this has been shelved. The process is now being demonstrated in a plant operated by Tati Nickel in Botswana. Commercial development here will be followed by Nkomati in South Africa and in the Avalon process in Western Australia. For the Activox process, the Cu/Ni/Co concentrate will be subjected to ultrafine grinding to produce a feed material of particle size >80% passing 10 microns. This finely ground material will be treated in an autoclave at 1000 kpa O2 at a temperature of 105-110oC. Most of the soluble iron is rejected at this stage as goethite. Pyrite does not dissolve under autoclave conditions and chromium reports to the leach residue.

After 7 stages of CCD, the PLS is treated by solvent extraction with a suitable hydroxyoxime extractant to recover copper which is then electrowon. A 50% bleed from the Cu SX raffinate is sent to iron precipitation with limestone. This is done in two stages, the first to pH 3.5, provides a partial iron removal but essentially no loss


of Ni or Co. This precipitate is removed and passed to the CCD circuit. The second stage precipitation is carried out at pH 4.5. After thickening the thickener overflow is passed through a sand filter and finally polished in a cartridge filter prior to Co SX. The precipitate from the second stage iron precipitation does contain significant amounts of Co and Ni. This is re-dissolved and is passed back to either the autoclave or directly into the Cu SX circuit.

The Co SX circuit, which uses CYANEX 272 as the extractant, is run principally to ensure virtually complete removal of cobalt from the nickel liquor. The Co SX raffinate passes through an after-settler followed by a diluent wash. The treated raffinate will then be subjected to active carbon treatment to ensure that no CYANEX 272 leaks through to the Versatic 10 circuit. The Co strip is treated in a Spintek filter prior to CoCO3 precipitation.

Nickel is then subjected to solvent extraction with Versatic acid. The nickel strip is treated with active carbon prior to Ni EW. The Ni is 30 g/l and the anodes are bagged. The Ni tankhouse is similar in design to the one that operated at Cawse until its sale to OMG.

Bioleaching for nickel/cobalt sulphides has, however, reached the pilot plant stage in the development of the BioNIC process by Billiton with work still on-going. It is also being used in the Sherlock Bay project in Australia. In this process sulphide concentrates are bioleached with a mixed population of mesophilic bacteria at 30-45oC to give dissolutions of nickel and cobalt of >90%. Iron is also dissolved and is removed by hydrolytic precipitation. Separation of cobalt and nickel from impurities such as Mg and Ca is achieved by ion exchange with a chelating resin, although solvent extraction options are being considered. Cobalt separation from nickel will be by solvent extraction with CYANEX 272. Zinc will coextract with cobalt but it will be eliminated by a selective strip. Co-extracted copper will be removed by selective ion exchange.

2.2.3 Direct Leaching of Nickel Laterites

In the Caron process the laterite ore is reduced and leached with ammoniacal ammonium carbonate and iron is rejected at the leaching step. Other impurities that might report to the leach liquor were rejected in the original Nicaro process by precipitation of the nickel and cobalt as mixed sulphides, which effectively separates the nickel and cobalt from magnesium and calcium, but not zinc if present. Other impurities such as chromium, aluminium and manganese will remain in the leach residue. The use of solvent extraction to remove nickel from cobalt as practised at the Queensland Nickel Yabulu refinery in Australia means that any impurities in the leach liquor report to the raffinate which is subjected to sulphide precipitation for cobalt recovery. The leach efficiency of the Caron process for nickel and cobalt is not high, typically 75-80% recovery of nickel and only 40-50% recovery of cobalt. Much cobalt is lost by co-precipitation with ferric iron in the leaching stage.

In acid pressure leaching, i.e. the original Moa Bay process, while most of the impurities would be rejected in the oxidative conditions of the pressure leach, small amounts of iron etc will report to the leach liquor. This is removed by hydrolytic precipitation before bulk nickel/cobalt precipitation as mixed sulphides, which again effectively separates the nickel and cobalt from manganese, magnesium, calcium etc. At Moa Bay in Cuba the sulphide precipitation is performed under pressure. On the other hand precipitation can be carried out without resort to pressure at 90oC in the presence of high concentrations of magnesium as at Murrin Murrin in Western Australia with a high rejection of manganese.


2.2.4 Purification of Copper-Cobalt Calcine Leach Liquors

Purification of copper-cobalt calcine leach liquors in Zambia, after direct copper electrowinning, as currently practised, consists of electrostripping of copper to ~ 0.5 g/l followed by iron, aluminium and residual copper precipitation with limestone in two stages, the first at pH 3.5-4.5 followed by a "clean-up" precipitation at pH 6.9-7.1. This "clean-up" precipitate contains 40-70% of the cobalt and is recycled to the head of the purification circuit. This results in a considerable recycling load of cobalt and impurity elements, particularly zinc. This has led to the installation of a bleed of liquor between the iron precipitation and "clean-up" precipitation stages for zinc removal by solvent extraction with DEHPA, with the raffinate returning to the liquor flow to the "clean-up" precipitation stage. The filtrate from the "clean-up" precipitation stage is further neutralised to pH 8.5 to precipitate cobalt hydroxide which is redissolved in spent cobalt tankhouse electrolyte, treated with active carbon to remove sulphides and organics and nickel is removed by chelating ion exchange using the Dow resin XFS 4195 (now reclassified Dowex M4195) prior to cobalt recovery by electrowinning. This process is hardly current state-of-the-art and potential improvements will be discussed later.

2.2.5 Cobalt-Nickel Separation

Cobalt-nickel separation is not a major problem in the processing of the copper-cobalt ores of the African copper belt. It is, however, very important in the processing of matte leach liquors and in processing nickel laterite leach liquors.

Cobalt/nickel separation in matte leach liquors has traditionally been carried out by precipitative processes. These are not now considered to be state-of-the-art and solvent extraction is the preferred process nowadays. Where chloride-based leaching processes are used cobalt is readily separated from nickel using trialkyl amines. The extracted cobalt is recovered by water stripping. In sulphate-based matte leaches, separation is easily achieved by using solvent extraction with CYANEX 272, particularly when high Ni:Co ratios have to be treated. Alkyl phosphoric and/or phosphonic acids can also be used but the separation factors are much smaller and therefore these extractants can only be used where a rough precipitative separation has first taken place - i.e. as in the nickelic hydroxide or chlorine gas oxidation process as operated at, for example, Rustenburg in South Africa by Anglo Platinum (nickelic hydroxide) and Jinchuan Nickel (chlorine oxidation) in China. Solvent extraction, unlike the precipitation processes described earlier, does offer the opportunity of complete separation with high yields and purity of the separated metals.

From ammoniacal solutions however, provided cobalt is in the Co(III) state, nickel can be successfully separated from cobalt with hydroxyoximes as Co(III) is not extracted by these reagents. This has been successfully commercialised by Queensland Nickel at their Yabulu refinery in Queensland, Australia.

Outokumpu Oy has published details of its CYANEX 272 cobalt solvent extraction plant at Harjavalta. The plant treats a feed from leaching of the mattes produced in the DON process containing 130 g/l nickel, 0.8-1.0 g/l cobalt with very minor amounts of zinc, copper, lead, manganese, magnesium, calcium and iron. Cobalt is extracted in four countercurrent stages, the loaded organic is scrubbed with dilute sulphuric acid in five stages and cobalt is stripped with sulphuric acid in four stages to produce a raffinate containing 130 g/l Ni, 0.01 g/l Co and a cobalt strip liquor containing 110 g/l Co, 0.02 g/l Ni, together with co-extracted copper, lead, manganese and some calcium. Co-extracted zinc and iron are not significantly stripped with the cobalt and these metal ions are removed in a single stage with 200 g/l H2SO4. The continuous countercurrent operation is controlled using the


Outokumpu Courier X-ray system for on-line analysis of cobalt and nickel in both aqueous and organic phases. On-line pH measurements can also be used.

Solvent extraction of cobalt is now also of interest in the processing of copper-cobalt ores. Here reagent choice may be more flexible depending on the impurities in the leach liquor. The reagents of interest will again be the alkyl phosphorous acids. The ability to interface cobalt stripping directly with electrowinning when CYANEX 272 is used for cobalt solvent extraction will tend to favour this reagent.

2.3 Production of the Final Cobalt Product

Cobalt can be recovered from a solvent extraction strip liquor as a salt or as metal. While production of a salt such as cobalt carbonate or other salts is usually straightforward, the production of cobalt oxide hydroxide at QNI in Australia is worth a mention. Here a cobalt sulphide product is produced after separation of nickel from cobalt by solvent extraction from ammoniacal solution with LIX84QN. The sulphide product so produced contains 40% Co and 2% Ni. This is pressure leached with sulphuric acid and iron and zinc impurities removed by solvent extraction with Cyanex 272. This reagent was chosen because of its superior performance with regard to iron stripping over other organophosphorus acids such as DEHPA or PC88A in that iron can be stripped with 100-200 g/l H2SO4 rather than 6 N HCl. Cobalt is then transferred to an ammoniacal solution by solvent extraction with DEHPA (selected on the basis of price) and stripped with the ammonia/ammonium carbonate solution used in the main nickel plant. After cobalt oxidation by aeration and hydrogen peroxide addition, nickel is removed by solvent extraction with LIX84QN. Calcium and magnesium are now removed by chelating ion exchange and cobalt is then precipitated by steam stripping of the ammoniacal solution.

Recovery as metal is more complicated. Metal recovery can be by electrowinning or as cobalt powder by pressure hydrogen reduction.

2.3.1 Pressure Hydrogen Reduction

The production of cobalt by pressure hydrogen reduction was pioneered by Sherritt Gordon in Canada. The ammoniacal reduction feed solutions typically contain from 40-80 g/l Co and from 300-400 g/l ammonium sulphate with NH3:Co mole ratios in the range of 2.2-2.6. Cobalt powder is produced by pressure hydrogen reduction at Sherritt Gordon's plant at Fort Saskatchewan in Alberta, Canada, by Impala Platinum in South Africa and by Murrin Murrin in Western Australia. It has been proposed for the Ambatovy nickel laterite project in Madagascar where Impala Platinum, until they withdrew from the project, was to take the mixed sulphides produced in Madagascar, pressure leach them, separate Co from Ni by solvent extraction with CYANEX 272, with final recovery of both cobalt and nickel by pressure hydrogen reduction.

2.3.2 Cobalt Electrowinning

In the INCO cobalt electrowinning plant at Port Colborne the cobalt liquor, after dissolution of the cobalt-nickel carbonate feed with sulphuric acid is purified through a series of precipitative steps using copper cementation with nickel powder. Zinc removal by ion exchange follows using a resin impregnate OC 1026 (a DEHPA resin impregnate, Bayer AG). Cobalt is then oxidised to cobaltic hydroxide to separate cobalt from nickel. The precipitated cobalt cake is re-dissolved with spent electrolyte and SO2 with addition of barium carbonate to remove lead. The liquor so produced is treated by ion exchange with the Dow resin XFS 4195 for copper and


nickel removal. This is the same resin as used by ZCCM at Nkana and Chambishi. This final liquor containing 90 g/l Co is then passed to the tankhouse.

At the ZCCM cobalt plant at Nkana in Zambia, the cobalt advance electrolyte is produced as described earlier. The cathode cycle in the tankhouse is 4-5 days and 8 tpd cobalt is produced. The electrolyte flow is 80 m3/hr at a temperature of 58oCand the current density is 280-320 ASM. Electrolyte flow is set so as to maintain a cell pH of 2.2-1.5. Stainless steel cathode blanks are used with Luilu anodes.

In such tankhouses where plate cathode is produced rather than the cobalt rounds at INCO, the cathodes must be stripped by chipping the deposited cobalt off the stainless steel cathode blanks, crushing the cobalt chips and vacuum degassing the crushed product at 750-850oC to remove any absorbed hydrogen. The degassed cobalt product can be further refined in a vacuum induction furnace at 1600-1670oCto remove volatile impurities such as lead, zinc, thallium and cadmium to less than 1 ppm and to lower oxygen levels to <50 ppm with addition of carbon granules.

The adoption of solvent extraction rather than precipitation of cobalt hydroxide to provide the interface with purified leach liquors and the cobalt tankhouse is considered to be a significant improvement. Such a process offers the possibility of a reduction in operating costs compared to the lime precipitation route and a significant increase in cobalt recovery across the purification circuit. Feed materials containing higher levels of zinc, nickel, manganese and magnesium can be treated by this route while maintaining the quality of the finished cobalt. An opportunity to improve current efficiency by the production of a higher concentration advance cobalt electrolyte is also provided. Bagged anodes can also now be used as calcium will be absent from the electrolyte. This will also allow the use of higher current densities thus increasing the capacity of existing tankhouses.


3 PRODUCTION

3.1 Reserves

Cobalt minerals are usually found within stratabound copper, nickel-copper sulphides, nickel laterites and silver-cobalt sulpharsenides.

The stratabound copper deposits, such as those in the Congo and the Zambian copperbelt, were the main commercial source of cobalt until the 1990s. The cobalt in these deposits is present as cobalt minerals and as pyrite. Typically, grade varies between 0.1% and 0.4% Co.

The nickel-copper sulphides found in Russia, Canada and Australia, are the result of concentrations in mafic and ultramafic igneous rocks. The typical grade of these deposits is around 0.1% Co.

Nickel laterite deposits may contain cobalt as oxide and silicate minerals, brought about by the weathering of ultramafic rocks. The grade of the Co present in the laterites is generally between 0.05% and 0.15%. The lateritic deposits are currently mined in Cuba, New Caledonia, Indonesia and Australia.

The silver-cobalt sulpharsenides were mined in Canada and Morocco, but there is now only one mine working this type of deposit left. The Bou Azzer mine in Morocco mines a cobalt deposit that has been graded at 1.5% Co.

Secondary forms of cobalt are available, which are mostly derived from recycled super alloy or cemented carbide scrap and from spent catalysts.

It is thought that there are millions of tonnes of cobalt located on the sea floor in manganese nodules, although no successful mining of these deposits has taken place owing to the large costs involved and also the slow development of the technology for deep sea mining.

3.2 Production by Country

Generally, mine production in the major countries is increasing as the demand for cobalt is increasing. However, mine production has reduced in the D R Congo and in Zambia, and this has had the effect that the actual total mined cobalt is decreasing in this region.

New plants came into operation in Western Australia and also in Africa between 1999 and 2001. However, there is now only one of the Australian plants still producing cobalt metal (in powder form). Cawse produces a mixed Ni/Co hydroxide which is shipped to Harjavalta in Finland for refining while QNI produces cobaltic oxide hydroxide. The Kasese operation in Uganda was placed on care and maintenance in 2003, but with the rising cobalt price and the increased demand, this was brought back online in early 2004.

Ore reserves by country, as indicated by the USGS are shown in table 3.1.


Table 3.1 Reserves by country4

Country Reserves (tonnes)

Reserve base (tonnes)

United States NA 860,000 Australia 1,400,000 1,700,000 Brazil 140,000 350,000 Canada 35,000 40,000 D.R. Congo 3,400,000 4,700,000 Cuba 1,000,000 1,800,000 Morocco 20,000 NA New Caledonia 230,000 860,000 Russia 250,000 350,000 Zambia 270,000 680,000 Other countries 200,000 1,500,000 World total (rounded)

7,000,000 13,000,000

There are several new cobalt projects under development. A list of the major new project developments in cobalt is shown in table 3.2. The majority of these are due to commence full-scale production in 2007/2008, so their effect on the cobalt market will not occur for 1-2 years.

Table 3.2 New Cobalt Developments Project Name Company Location Expected

max. output (tpa)

Coral Bay Coral Bay Nickel Corp. Philippines 750 Voisey’s Bay INCO Canada 2,100 Ravensthorpe QNI Australia 1,600 Idaho Cobalt Formation USA 1,500 Çalda European Nickel Turkey 800 Goro INCO New Caledonia 4,500 Ambatovy Dynatec Corporation/

Sumitomo JV Madagascar 5,600

Vermelho Compania Vale do Rio Doce Brazil 3,000 Weda Bay PT Weda Bay/PT Aneka

Tambang Tbk JV Indonesia 4,500

Kolwezi Adastra DRC 6,000 TenkeFungurume

Tenke Mining Corporation DRC 6,000

Ramu Highlands Pacific Papua New Guinea

3,200

Agata Mindoro Nickel Philippines unknown Ruashi Metorex DRC 1,000 Malborough Preston Resources Australia 2,000 El Boleo Baja Mining Mexico 2,100 NICO Fortune Minerals Canada 1,500

TOTAL 33,050

Coral Bay ships mixed Ni/Co sulphides to Sumitomo's Niihama refinery in Japan.

4 Shedd, K, January 2005. USGS Mineral Commodity Summaries. Taken from http://minerals.er.usgs.gov/minerals/pubs/commodity/cobalt/cobalmcs05.pdf


Impala has withdrawn from the Ambatovy jv with Dynatec. A new partner for refining of the mixed sulphide product to be produced in Madagascar has yet to be found.

Weda Bay will also produce a mixed Ni/Co sulphide intermediate product for further refining possibly by OMG at Harjavalta in Finland.

Other projects of potential significance include IvanPlats Syerston project (4-500 t/y Co) in New South Wales and Geovic's Ni/Co/Mn project in Cameroon (4,700 t/y Co). Heron Resources in Western Australia aims to produce 3000 t/y Co in a mixed intermediate possibly for refining at INCO's Goro plant in New Caledonia. Other likely significant producers of cobalt are Equinox Minerals Ltd at Lumwana in Zambia, Metallica Minerals/BHP Billiton at the Nornico project in Australia and Sherlock Bay Nickel in Australia. One of the largest primary copper/cobalt oxide properties is the Nama deposit owned by Caledonia Mining Corporation in Northern Zambia. Development testwork is currently underway with an aim to produce a feed for smelting.

3.3 Production by Company

There are many companies worldwide who produce cobalt. The largest producer in 2004 was the Jinchuan Non-Ferrous Metals Company, based in China. The majority of the cobalt they refine is mined in the D.R. Congo and their production is currently around 8,000 tonnes a year. This is expected to increase in 2005.

In previous years, the largest producer was the OM Group Inc. of the USA. The OM Group do not mine cobalt, but process cobaltiferous materials obtained from smelting (Big Hill smelter, DRC; Harjavalta smelter, Finland and elsewhere - e.g. mixed hydroxides from Cawse, Western Australia) by hydrometallurgical means at Harjavalta and Kokkola in Finland. The cobalt production by company is listed in table 3.3. The table includes both mining and smelting companies.


Table 3.3 Cobalt production by company 1998-2004 1998 1999 2000 2001 2002 2003 2004CTT 241 470 1,200 1,200 1,354 1,431 1,593Eramet 172 180 204 199 176 181 199 Falconbridge 3,851 4,009 3,433 3,314 3,993 4,556 4,670ICCI 2,640 2,770 2,855 2,943 3,065 3,141 3,225Inco 1,740 1,420 1,470 1,450 1,480 1,000 1,562Norilsk 3,700 4,000 4,100 4,600 4,200 4,654 4,524OM Group 5,250 6,200 7,700 8,100 8,200 7,990 7,893QNI (BHP Billiton Group)

1,395 1,539 1,520 1,818 1,863 1,800 1,900

Sumitomo 329 221 311 350 354 379 429 Umicore5 1,200 950 1,110 1,090 1,135 1,704 2,947Chambishi Metals - - 2,316 2,789 4,344 4,570 3,769Bulong6 - 79 170 203 200 0 0 Gécamines 4,490 5,180 4,320 3,199 2,149 1,200 735 Kasese7 - 77 420 634 450 0 457 Mopani Copper - - 1,026 1,876 1,800 2,050 2,022Murrin Murrin - 83 925 1,452 1,838 2,039 1,979JNMC8 - - - - - - 8,000

5 Includes global refined production. 6 Closed in 2003 and subsequently sold to Western Minerals Technology who propose to process Ni/Co sulphide concentrates through the plant in their Avalon project using the Activox process.

7 Placed on care and maintenance 2003-2004. 8 Production data only available for 2004.


4 DEMAND

4.1 Applications

Cobalt usage is dominated by its usage in batteries and Superalloys. It is also used substantially in the production of catalysts and the electrowinning of copper. The major applications of cobalt include:

Metallurgical - Superalloys - Wear Resistant Coatings- High Speed Steels - Prosthetics- Low Expansion Alloys - Steels- Corrosion Resistant Alloys - Spring Alloys

Magnetic Alloys - Alnico’s- Rare Earths- Soft Magnetic Materials

Chemicals - Batteries - Catalysts - Adhesives – Cobalt Soaps - Specialist Chemicals: Driers, Pigments, Colours - Electroplating - Agriculture and Medicine - Electro-Magnetic Recording

Cemented Carbides - Tungsten hard metal powders

Bonded Diamonds

Electronics - Recording Material- Matched Expansion Alloys - Leads

Ceramics & Enamels - Colours in glass, enamels, pottery and china


4.2 Demand

The distribution of cobalt demand by application is shown in table 4.1

Table 4.1 Demand for cobalt by application Market % Cobalt Materials Possibility Superalloy (Ni/Co/Fe) 20.0 Metal, recycle

Hardfacing & Other Alloys 5.5 Metal & mesh powders, recycle Magnets All Types 7.0 Metal, powders, recycle Hard Materials – Carbides,

Diamond Tooling

10.5 Powders – Fine and very fine

Catalysts 11.0 Salts – Carbonate, sulphate, nitrate, Acetate, metal

Colours – Glass, Enamels, Plastics, Ceramics, Artists Colours, Fabrics

4.5 Oxide, + some sulphate, hydroxide, carbonate

Feestuffs, Anodising, Recording, Electrolysis, Cu Electrowinning

11.0 Mainly sulphate, but some carbonate and hydroxide

Batteries 21.0 Hydroxide, powder, LiCoO2

Tyre Adhesives, Soaps, Driers (paint/ink) 9.5 Soaps & complexes made from metal starting point

The major usage of cobalt is within batteries and superalloys. The cobalt market has been steadily growing since 2001 in Europe and America, but in China and Japan, the market for cobalt has been rising rapidly. Almost all the new demand for cobalt worldwide is for use in batteries. This is due to the increasing global demand for products such as mobile phones, PDAs, laptops, etc.

A new, significant growth area is for batteries for hybrid cars mainly in the USA, China and Japan. Currently nickel metal hydride batteries are used but Li ion or Li polymer batteries are preferred in the longer term which use more cobalt than NiMH batteries. Substitution of LiMnCoO2 or LiNiMnCoO2 will, however, reduce the rate of growth of cobalt demand as these compounds contain only 1/3rd of the amount of cobalt as LiCoO2.

4.3 Consumption

Cobalt consumption for 2004 was 51,400 tonnes, an increase of 8.29% on the figures for 2003. As previously stated, the growing demand and usage of cobalt is due mainly to the increased use of cobalt in the battery market. The increase has also been helped by the continued strength of the market for superalloys, especially with regards to the making of jet engines and aerospace parts.


5 PRICES

Cobalt is generally sold in one of two forms; as a 93% cobalt powder and as a 98% cobalt powder. Historical prices for these two forms of cobalt are shown in figure 5.1.

Figure 5.1 Average cobalt price from 2000 – 2005.

The fluctuations in the cobalt price have come as a result of the ever changing demand and supply of cobalt. Growth in the Chinese and Japanese markets has seen the amount of cobalt in demand increasing. Increasing supplies has driven down the price of cobalt to lows of about $13 per pound from around $19 at the beginning of 2004.

Some analysts see rising demand pushing it back up to the $18-to-$25 range through 2008, but the general consensus is that a medium-term range of $12-15 per pound is most likely.

By way of example, GFMS Metals Consultancy predicts a cobalt price averaging around $14.80/lb for the period 2001-2010. Inco predicts a price of $15/lb for 2009, when their Voisey’s Bay project is expected to be operating.


6. CONCLUSIONS

In spite of the relative abundance of cobalt and improved processing techniques, currently demand for cobalt is exceeding annual production. Reasons for this have been the strong increase in demand, the continuing low production from D R Congo and the legacy of problems associated with the first three nickel/cobalt laterite pressure leach plants in Australia, which have slowed further developments of this technology.

This deficit is likely to continue in the future; although perversely if all the proposed new cobalt developments, totalling an extra 33,050 tonnes per year, came to fruition, this could lead to the market being in significant surplus. With the increasing use of cobalt in batteries and chemical products, however, demand is set to grow further over the next few years. Accordingly, perceived wisdom for the medium-term cobalt market is based on the assumption that not all projects will proceed and the supply-demand balance will achieve equilibrium.

The caution on behalf of new projects reflects in particular continuing concerns about stability in D R Congo where many of the large projects are located.

The growth in demand for cobalt for batteries for hybrid electric vehicles could increase cobalt demand substantially in the next five years. Future demand will depend on not only economic growth in developed countries but also on the rapidly developing economies of China and India.

Demand for cobalt for superalloys is also likely to be strong, with aircraft orders at record levels. The demand for spare parts is also high as airplane capacities increase.

It can be seen that, prior to late 2003, the price of cobalt was drifting down to around $6/lb. The large surge which took cobalt to over $25/lb is now in the past, but the profile of the price decline since this point, coupled with buoyant demand predictions, indicates that a medium term forecast of around $14/lb is not unrealistic.

St Barbara recommends that a price range of $12 - $15/lb Co be considered for the medium term. It is felt that this price level allows both miners and producers to profit equally and represents a sensible level for a growing market in approximate supply/demand equilibrium.

Appendix 11D –

Platinum Group Metals Industry Review

PLATINUM GROUP METALS

MARKET REVIEW

St Barbara Consultancy Services Report No E705

March 2006

St Barbara Consultancy Services – PGM Market Review 2

1. INTRODUCTION

Although the presence of platinum has been detected in artefacts made by the Egyptians in around 700 BC, the first platinum product was jewellery, manufactured by the Incas around 500 AD. It took a further 12 centuries and contributions from scientists Lavoisier, Wollaston, Priestley and Franklin before the metal was properly identified around 1750. The other members of the group, palladium, rhodium, iridium and osmium were separated some 50 years later, with ruthenium identified belatedly in 1844.

Not only was platinum found to be attractive in ornamental uses and to have the potential for investment, but a number of extremely practical properties were noted, including hardness, resistance to corrosion, high melting point, reflectivity, conductivity and malleability. These prompted the likes of Faraday and Davy, who relied on platinum as a means of carrying, making and breaking a current. Platinum was used in the first electric telegraph, incandescent lamp and thermo-ionic valve and also in early internal combustion engines.

Platinum is the most widely known of the six platinum group metals and has been used in industry throughout the twentieth century, most notably for its catalytic properties. It is utilised as the main active component of catalytic converters, which are now fitted to all new automobiles so that they can meet low emissions legislation.

Platinum has a low concentration in the earth’s crust (0.0000028%), about half the level of silver, but 8 times higher than gold. It is found in alluvial deposits, but its exploitable reserves are in lode deposits in association with nickel and copper sulphides.

Unlike gold, the economic occurrence of platinum is dominated by one country, South Africa, which is responsible for over two-thirds of primary supply. Russia is the other major source, with around one eight of mined output. Consumption is dominated by Europe, North America and Japan, who together consume over three quarters of the world’s platinum. These countries are consequently the main sources of recyclable platinum.

Until fairly recently, the structure of the platinum industry was not very transparent even to specialist industrial analysts. This has changed, however, with developments in South Africa in particular, and it is now possible to gain a good insight into the industry and the major players, although reliable Russian data remains difficult to obtain.


2 SUPPLY

Platinum production has generally been increasing over the past 10 years. Chart 1 shows platinum production by country.

From this, it is easy to see that the supply from South Africa has been steadily increasing, but Russian supply is now in decline, although data for the Russian platinum industry is hard to come by, despite the recent passing of a bill to declassify platinum information. Of the other countries’ production, Zimbabwe is responsible for around 90%.

Figure 2.1 World Platinum Supply 1996 - 2005

World Platinum Supply 1996-2005

0

1,000

2,000

3,000

4,000

5,000

6,000

1996 1997 1998 1999 2000 2001 2002 2003 2004 2005

Year

'000s o

z South Africa

Russia

North America

Others

2.1 World Reserves

Given the relative abundance of platinum in the earth’s crust and the fact that it is concentrated into a few significant locations, it is not surprising that the mine owners are very positive about the size of their platinum reserves.

The Bushveld Complex supports most of the South African producers of platinum. In 2005, there were 17 mines in operation here – 12 of them from the Merensky Reef and the UG2 Chromite Layer.

Russian reserves of platinum occur mainly with the copper-nickel ores associated with the Norilsk and Pechanga groups of deposits. Further quantities are found with vanadium-iron-copper ores and as placer deposits.


World mine reserves and reserve base as estimated by USGS are shown below:

Table 2.1 Reserves of Platinum1

Country Reserves (kg) Base Reserves (kg) United States 900,000 2,000,000

Canada 310,000 390,000 Russia 6,200,000 6,600,000

South Africa 63,000,000 70,000,000 Others 800,000 850,000

TOTAL (rounded) 71,000,000 80,000,000

It can be seen that there are adequate reserves of platinum group metals to last for well over a hundred years.

2.2 Primary Producers

Primary producers are defined as those companies that mine the platinum bearing ore principally for the value of the platinum. In all cases the platinum occurs with other metals, such as nickel, copper, gold and cobalt, and the extraction processes will yield saleable quantities of these metals, thus adding to the income of the mining company.

The main primary producers of platinum are:

Lonmin Anglo Platinum Impala Platinum Aquarius Platinum ARM Platinum Stillwater Mining Zimplats

2.3 By-Product Suppliers

By-product suppliers are defined as those companies who recover platinum values as a by-product and for whom the base metals provide the greatest revenues.

The main producers of by-product platinum are:

Norilsk Nickel IncoFalconbridgeRio Tinto

2.4 Secondary Supply (Recycling)

The secondary supply comes from recycling and refining services. For obvious reasons, wherever possible all platinum products, with the exception of jewellery and investment products, are carefully collected by the customers after use and returned to the manufacturers for refining and credit. In many cases customers operate a “metal account” with their suppliers, with scrap adding to the value and products debited from it. Ongoing losses in the circuit are replaced with purchases of platinum to “top up” the account.

1 Michael W. George, US Geological Survey, Mineral Commodity Summaries, Platinum Group Metals, January 2006.


The secondary world supply of platinum is thus to be seen as a regular addition to the total quantity of platinum in circulation, providing for new and expanding markets, making up for processing losses and adding to the jewellery, investment and strategic stocks that are not recycled.

The total quantity of platinum in use by the industry is thus greater than the volume of the new supply provided by the producers. According to US government figures, during 2004/5 the US secondary refining industry processed over 0.3 million ounces of platinum.

The North American market is the most advanced with respect to recycling and has a relatively small investment and jewellery sector, thus the recycling rate of 30% is not typical for the world market. The Johnson Matthey reports on supply and demand only estimate the effect of recycling for the autocatalyst market. A reasonable global figure for recycling would be in the region of 15% of total production.

In the western world, most of the secondary platinum processors also operate in other business areas and each has developed in a different way, so that the proportion of their total business that is related solely to platinum is not readily apparent.

The major secondary producers of platinum are:

Johnson Matthey DegussaHeraeusEngelhardMultimetco

The largest single area of the platinum business, autocatalyst supply, has provided the greatest challenge to the recycling companies. The spent catalysts may be replaced with the exhaust system periodically during the life of a car, but a unit will still be fitted to the car when it reaches the breakers. A financial incentive is needed to have the unit removed and set-aside for collection. The collector will have to pay a price per unit without knowing the condition of the catalyst (it may have had most of the platinum eroded!). After a number of years addressing the market, a few specialist companies have solved the problems and operate a reasonably complete collection service in North America.

This activity is being extended into Europe and elsewhere, as the use of catalysts fon vehicles becomes “mature’. The collected catalysts are de-tinned and crushed to provide a refinery feedstock. The very low metal content makes standard refining by leaching or dissolution non-viable, and the only way to process it effectively is by a smelting technique. One specialist company in Alabama operates an arc furnace capable of processing spent autocatalyst and this facility processes a small portion of the North American business. Otherwise collectors ship their product for refining in Japan or in Europe. The Stillwater Company in Montana is hoping to attract some of this business to its upgraded smelter facility in Montana in the near future.

2.5 Stockpile Sales/Strategic Stocks

The influence of holding stocks is not as great as that experienced in the gold market, but platinum, palladium and iridium are held as a strategic stocks by the US Government. The current stockpile levels are approximately 650kg, 570kg and 500kg respectively.


Platinum stocks are held by producers and banks, notably in Russia, and are sold as required to balance the supplies with world demand. Other governments are also believed to hold certain strategic stocks of platinum, but data is not readily available.

2.6 Planned Expansions and New Projects

A list of new PGM projects and current expansion programmes can be found in the table below.

Table 2.2 A list of PGM developments Project Name Country Status Type CompanyImpala2 South Africa Operating,

expansion UG Implats

Sheba’s Ridge2 South Africa Prefeasibility OP Ridge Mining, Anglo Platinum

Twickenham2 South Africa Prefeasibility UG Anglo Platinum Pandora2 South Africa Feasibility UG Anglo Platinum, Lonmin,

Implats, Bapo Ba,Mvelaphanda

Rustenburg UG22 South Africa Operating, expansion

UG Anglo Platinum

Leeuwkop2 South Africa Prefeasibility UG Southern African Bafokeng Styldiri2 South Africa Prefeasibility UG Anglo Platinum Rustenburg Section2

South Africa Operating, expansion

UG Anglo Platinum

Mototolo2 South Africa Prefeasibility UG Anglo Platinum, Xstrata Two Rivers2 South Africa Construction UG Harmony, Implats Drenthe2 South Africa Conceptual UG Anglo Platinum, Anooraq Waterval2 South Africa Construction UG Anglo Platinum Marathon2 Canada Prefeasibility OP Marathon PGM Kola Russia Exploration n/a Eurasia West Kytlim Russia Exploration AM Eurasia, Anglo Platinum Baranchinski Russia Pitting n/a Eurasia Bakgaga South Africa Exploration OP Superior Mining Londokomanana Madagascar Exploration n/a Jubilee Platinum

(Abbreviations – UG = underground, OP = open pit, AM = alluvial mining, n/a = not available)

2 Ericsson, M., Olsson, A., 2006. E&MJ’s Project Survey 2006, Engineering and Mining Journal, Vol 207 (1), 55-59


3. DEMAND

Platinum’s end-uses can be grouped into two broad categories — those which are based on value/aesthetic appearance, namely investment and jewellery - and industrial applications which exploit platinum’s chemical and physical properties.

Platinum demand has risen overall by approximately 0.2 million oz between 2004 and 2005, from approximately 6.5 million oz to 6.7 million oz. It is apparent that the demand for platinum for investment purposes has dwindled to almost zero. Demand for industrial applications and jewellery have risen only very slightly. However, demand for platinum in catalysts has risen by approximately 0.17 million oz.

The table below shows the demand by application for platinum (2004 figures).

Table 3.1 Platinum Demand by Application Application ‘000 oz Autocatalyst 2,735 Chemical 350 Electrical 280 Glass 240 Investment 5 Jewellery 2,200 Petroleum 150 Other 510 Total 6,470

3.1 Automotive

Since the initial deliveries of autocatalysts to motor car manufacturers in 1974, over 400 million units have been fitted to passenger cars and small trucks. The growth in this market has been in line with the adoption of Clean Air programs in many countries, following the 1970 US Clean Air Act which set emission regulations for the 1975 automobile model year and the passing of legislation to enforce clean air standards.

The United States, particularly California, has set the lead both in legislation and in the subsequent tightening of emission standards. A National Low Emission Vehicle standard was applied across the United States from 2000. California has established tighter regulations (LEV II) that took effect in 2004, which have been applied across the States from that year.

The European Parliament has ratified new regulations which were implemented in 2000 and 2005 that reduce allowed emissions, leading to increased use of platinum for autocatalysts.

South Korea, Australia, Brazil, Chile and Argentina all have legislation requiring the use of autocatalysts. All cars sold into Malaysia, Singapore, Hong Kong, Taiwan and Thailand must be fitted with catalysts. Certain cities in India have pollution control legislation that require vehicles to be fitted with catalysts and this was supposed to be applied nationwide during 2000. China seems to be following the same pattern as India.

Legislation is now extending from cars to include light vans and motorcycles. This gives scope for continually increasing demand for platinum over the next 10 years.

The 2004 figures show that 2.74 million oz of platinum were used for autocatalysts worldwide


3.2 Industrial Applications

Platinum finds a host of uses within industry in general. The extreme corrosion resistance and the catalytic properties of platinum being the salient properties that differentiates it from non-noble metals. Applications are diverse, but include, by way of illustration:

Platinum and platinum/rhodium alloys are used to contain molten glass and to form the spinnerets for glass fibre production. Platinum, together with ruthenium and iridium, is used to form protective coatings for electrodes used in chlorine manufacture and cathodic protection devices. Platinum/rhodium catalysts are used for the oxidation of ammonia to produce fertilisers. Petroleum refining processes use platinum catalysts in a number of stages. The demand for cleaner fuels increasing their usage.

The demand for each of the industrial products that depend upon platinum technology fluctuates from year to year with the economic situation and regional market factors, but the spread of activity is so great that a fairly stable business base remains for the secondary platinum suppliers.

These various applications consumed around 0.5 million ounces of platinum during 2004. Supply was not a problem for these limited demand applications that were heavily involved in recycling platinum.

None of these applications had any effect on the prices for platinum; in effect they had to ride the price fluctuations caused elsewhere.

3.3 Jewellery

Platinum has a long history of use in jewellery manufacture and its appreciation has been most notable in Japan. Although some platinum items were imported before, the first platinum jewellery was produced there in 1908. Significantly, in 1952 the Japanese government liberalised platinum imports; this being 20 years before gold was liberalised. The growth in platinum jewellery production paralleled the rapid growth in the post-war economy of the country. Initially supplies were obtained from Russia then, at the end of 1969, South African platinum also began to be imported.

The Tokyo Gold Futures Exchange started a futures market for platinum in 1984. This was significant in that it supported a large home market for the platinum and also fuelled the domestic interest in platinum (and palladium) as both investment vehicles and as valued jewellery items.

3.4 Investment

Small scale investment into platinum stood at 25,000 oz in 2004, though this was offset by a reduction of 20,000 ounces in large scale investment. Overall investment in platinum was a mere 5,000 ounces compared with the 90,000 ounces overall in 2001. The demand for platinum as an investment product varies with the platinum price. 2004 saw the highest platinum prices recorded, and the price in 2005 was higher still.


3.5 Healthcare

In the medical market platinum is used as the metal (or alloy) in biomedical devices and, in salt form, as an anti-cancer treatment.

The use of platinum in medical devices and in cancer chemotherapy is of small volume, perhaps 20,000 ounces worldwide, but quite lucrative for the suppliers.

3.6 Substitution

Some motor vehicle manufacturers have substituted palladium for the more expensive platinum in catalytic converters. Until recently, only platinum could be used in diesel catalytic converters, although new technologies now allow palladium to be used. For most other end uses, PGMs can be substitute for other PGMs, with some losses in efficiency. In addition, electronic parts manufacturers are reducing the average palladium content of the conductive pastes used to form the electrodes of multilayer ceramic capacitors by substituting base metals or silver-palladium pastes that contain significantly less palladium.

3.7 Summary

Platinum demand is dominated by two main sectors:

Jewellery manufacture is the largest consumer of platinum metal and is thus the major driver of future demand. It accounted for around a third of platinum demand in 2004.

The autocatalyst sector has consumed ever increasing quantities of platinum over the past 5 years. Since 2000, the demand for platinum by the automotive industry has nearly doubled. The main drivers for off-take are platinum loadings in autocatalysts, vehicle production and catalyst recycling. In recent years, although vehicle production has been depressed, loadings have been on the increase, and this has help stabilise off-take. In the future, we believe that increased recycling of autocatalysts will play a significant role in platinum’s supply and demand balance. Since autocatalysts were first introduced, platinum loadings have steadily been increasing. Therefore, with each passing year, not only are more autocatalysts being recycled, but their platinum content is higher also.

Other applications consume much smaller quantities of metal and a normal growth of 1 - 2% per annum is likely for the next period. New technology, such as the supply of optical glass for high definition televisions, has demanded high levels of platinum to equip new factories, but this has now largely been satisfied. Such “local” surges in demand are a feature of sectors of the platinum industry where the metal is used as a “tool” to facilitate technology, as opposed to being used as an end product.

Total demand for platinum is set to increase by around 2-3% in 2006, with total consumption expected to reach over 7 million ounces by 2007.


4 PRICES

4.1 Platinum Price

Platinum prices have recently risen to over US$1000 an ounce, double the price of platinum in 2001. A graph showing the price trends over the last 30 years is shown below.

Pt Average Yearly Price 1975-2005

0

100

200

300

400

500

600

700

800

900

1000

1970 1975 1980 1985 1990 1995 2000 2005 2010

Year

US

$/o

z

Average Yearly Price

Figure 4.1 30 year price for platinum

Price predictions for platinum have been made by several companies.

Johnson Matthey predicts an average platinum price of US$890/oz in 2006. Stillwater Mining, a junior mining firm in America, predicts the platinum price to average US$830/oz between 2005 and 2007. Equapolar, a Canadian consultancy firm, have published a study entitled "Platinum Group Metals: World Resources Economic & the Future," authors Gary H.K. Pearse and Yana Amis predict that platinum prices could average $1,000 an ounce by 2009.

Currently demand for platinum is exceeding annual production, with this trend likely to continue. Projected strong demand for platinum from the automotive, electronics and jewellery sectors has enabled producers to raise capital for mining expansions in Southern Africa and Russia.

The price levels which should be sustained for the present will give producers and manufacturers the opportunity to operate profitably, but the high cost of platinum mining, processing and refining operations will limit success to those companies who have high quality resources and efficient management, plant and equipment. High prices


can cause shifts in demand from platinum to palladium (autocatalyst and jewellery) and this will result in eventual price stabilisation.

Comparison of Price and Pt Stocks, 1975-2005

-1000

-800

-600

-400

-200

0

200

400

600

800

1000

1975

1976

1977

1978

1979

1980

1981

1982

1983

1984

1985

1986

1987

1988

1989

1990

1991

1992

1993

1994

1995

1996

1997

1998

1999

2000

2001

2002

2003

2004

2005

Year

Difference in supply and demand ('000s oz) Pt Price (US$/oz)

This chart shows a comparison between the platinum price and the difference between the supply and demand. From the graph, it is apparent that the current high platinum price is due to a stock deficit, but this is likely to decrease once there is surplus stock. It is reasonable to assume from the data obtained, that the difference between the supply and demand is cyclical, and therefore, the deficit is likely to decrease sufficiently to drive the platinum price down as it has done in previous years.

The effect of investment funds has also undoubtedly been a factor in recent, buoyant PGM prices.


4.2 Palladium Price

Pt & Pd Price 1975-2005 (Yearly Average)

0

100

200

300

400

500

600

700

800

900

1000

1970 1975 1980 1985 1990 1995 2000 2005 2010

Year

US

$/o

z

Pt Price (US$/oz)

Pd Price (US$/oz)

The graph above shows the trend in the platinum price compared with the trend in the palladium price.

The palladium price is expected to continue to show the same trends, with the price moving with the change in palladium stocks. The market fluctuations are such that there can be huge differences between the supply and demand, both in surplus and shortfall.

The price is also dependent upon investment in platinum. High investment into platinum will increase the price and cause consumers of platinum to seek out the alternative, palladium. The palladium market cannot support this sudden change in demand, and there will be a huge shortfall leading to a price increase. This is turn will cause demand to fall, and lower the price.

Predictions for the palladium price could only be found from one source. This was written by Gary Pearse and Yana Amis of Equapolar Resources, and is entitled “Platinum Group Metals: World Resources Economic and The Future.” This gave a short term and longer term price prediction, which are US$200/oz until 2007, then increasing steadily to around US$350/oz by 2014.

Whilst predictions are useful, they should not be taken as hard evidence of the future trends in the palladium price, and caution should be taken when using predictions for the production of mining cost/revenue plans.


5 SUMMARY

The medium-term forecast for the PGM market is difficult to predict. New platinum developments are occurring in Southern Africa and Kazakhstan as outlined, but information regarding these developments is scarce and it is likely that there are more developments occurring for which there is no information. The industry is extremely secretive, and this has made efforts to investigate the industry difficult.

Currently, the demand for platinum is outweighing the supply. The difference between the two traditionally occurs as a cycle. The price of platinum is dependent upon this cycle, but it is also dependent on the investment side of the platinum market. High investment will cause the platinum price to rise, but this will have a negative effect on the consumers of platinum, who will try and switch to cheaper alternatives, generally palladium, in order to save costs. This will then lead to the price of platinum decreasing, especially as the decrease in price will lead to the selling of investment stocks.

A consensus price forecast for platinum the next five years is in the range $800 - $900/oz. For palladium, the forecast is in the range of $200-350/oz.

Appendix 12A –

Investment Costs

Appendix 12A

Kevitsa / Production of Ni and Cu Concentrates / Ore Feed Capacity 4.5 Mt/a

Investment Costs (All figures in EUR 1 000s) / Version 2

1. Receiving hopper & feeder 190

2. Jaw crusher 4003. Belt conveyor, 1.4 m/12 m 304. Crushing plant building (20 m x 20 m x 15 m), overhead crane 35 ton 8005. Belt conveyor, 1.4 m/50 m, 2 units 1606. Covered surge ore pile, live capacity 24 h = 13 000 ton (41 m x 41 m x 16 m) 2307. Apron feeder, 2 units 1508. SAG mill, D 8.5 m, L 4.3 m, 5500 kW 5,7009. Vibrating screen 10010. Belt conveyor, 1.0 m/7 m, 2 units 4011. Belt conveyor, 1.0 m/17 m 3012. Ball mill, D 7.0 m, L 11.0 m, 6000 kW 5,40013. Regrinding ball mill, D 2.7 m, L 3.6 m, 300 kW 48014. Hydro cyclone classifier, 3 x 12 cyclones, 360

15. Conditioner, 2 x 130 m3, 2 x 70 m3, 2 x 50 m3, 2 x 30 m3, 1 x 20 m3 700

16. Flotation cell 130 m3, 6 units 1,230





21. Flotation air blower, 2 units, 400 kW/unit 92022. High Rate thickener, 9 m diameter (2 units) 35023. High Rate thickener, 30 m diameter (1 unit) 59024. Conventional thickener, 40 m diameter (1 unit) 940

25. Lamella thickener, 4 000 m2 (1 unit) 940



28. Belt conveyor, 0.8 m/20 m, 2 units 5029. Belt conveyor, 0.8 m/30 m, 2 units 6030. Belt conveyor, 0.8 m/15 m, 2 units 4031. Concentrate storage, 2 units 20032. Front end loader 7033. Cyclone feed pump, 3 units duty + 2 units standby, 250 kW/pump 21034. Froth pump, 6 units duty + 6 units standby, 18.5 kW/pump 16035. Froth pump, 3 units duty + 3 units standby, 55 kW/pump 18036. Froth pump, 2 units duty + 2 units standby, 75 kW/pump 14037. Thickener underflow pump, 1 unit duty + 1 unit standby, 150 kW/pump 12038. Thickener underflow pump, 2 units duty + 2 units standby, 5 kW/pump 2039. Thickener overflow pump, 2 unit duty + 2 unit standby, 75 kW/pump 14040. Thickener overflow pump, 2 units duty + 2 units standby, 5 kW/pump 2041. Tailings pump, 2 units duty + 2 units standby, 250 kW/pump 31042. Tailings pump, 2 units duty + 2 units standby, 75 kW/pump 14043. Sump pump, 5 units duty, 18.5 kW/pump 8044. Waste rock heap and tailings pond ditch pump, 8 units duty + 8 units standby, 18.5

kW/pump230

45. Process control and management system 75046. On-stream XRF slurry analyzer system (Cu,Ni) 50047. On-stream particle size analyzer 8048. Field instruments (including reagent feeding instrument) 44049. Reagent storages and mixing 500

50. Laboratory analytical & mineral processing equipment 60051. Concentrator plant building (100 m x 45 m x 25 m), 2 overhead cranes 35 ton 6,70052. Dust control, heating and ventilation 1,00053. Maintenance shop (20 m x 15 m), overhead crane 35 ton 68054. First aid, change and warehouse (30 m x 15 m) 43055. Canteen (15 m x 15 m) 35056. Administration house (15 m x 12 m) 21057. Truck service shop (15 m x 15 m) 29058. Transportation of process equipment 70059. Steel support for process equipment 2,20060. Process piping inside process building 1,50061. Mechanical assembly of process equipment 1,40062. Electrical components and assembly (process equipment) 2,10063. Spare parts 3,00064. First fill 1,50065. Plant site, 2 ha 24066. Internal roads, including mine site, 1 km x 6 m, 1 km x 12 m, 0.5 km x 20 m 34067. Electric power line, 110 kV, 5 km 50068. Connection to Vajukoski power station 20069. Electric transformer at Keivitsa, 25 MVA, 6 connections for 20 kV 1,070

70. Low sulfur tailings pond (2 km2): Starter dams from 225 m to 232 m above sea

level, total length 3.7 km. Bottom improvement by adding 1 m till, area 0.4 km2

5,030

71. Low sulfur tailings pond: Raising of embankment to 239 m above sea level, total

dam length 4.3 km2,930





74. High sulfur tailings pond (0.22 km2): Starter dams from 223 m to 230 m above sea

level, total length 1.4 km. Bottom improvement by adding and tightening of peat, area

0.23 km2. Embankment improvement by adding bentonite cover

2,580

75. High sulfur tailings pond: Raising of embankment to 237 m above sea level, total

dam length 1.4 km980





78. Circulation water reservoir (0.19 km2): Dam length 1.8 km, height 7 m 1,260

79. Tailings pipe line, 3 km, D 0.35 m (steel) 54080. Reclaim water pump, (1 + 1), 55 kW/pump, pipe line 1.5 km, D 0.35 m (plastic) 180

81. Fresh water pipe line from Kitinen river + return water pipe line: 2 x 5 km, D 0.5 m

(plastic). Kitinen pump station: Duty 2 x 200 kW, standby 1 x 200 kW. Return water

pump station: Duty 1 x 55 kW, standby 1 x 55 kW

1,140

82. Ditches around waste rock and tailing ponds areas, 15 km 1083. Fence around open pit 2084. Acquisition of land area 25085. New bridge over Kitinen river 1,00086. Access road to Keivitsa, 6 km 1,20087. Mine electric power line, 1 km 7088. Mine surveying equipment 5089. Mine water removal pumps + pipe line to tailings pond, 2 km 20090. Mine planning system 10091. Auxiliary mine equipment 10092. Potable water supply + sewage treatment plant 200

93. Full feasibility study 2,00094. EPCM 4,000Contingency (20 %), excluding items 71, 72, 73, 75, 76 and 77, see table 12-1 14,320

Sum of investment costs 98,630

Appendix 13A –

Organisational Structure

Genera

l M

anager

Ad

min

Min

e M

anager

Secre

tary

Nurs

e

Secre

tary

A/C

Su

pL

og

istics S

up

A/C

Cle

rk

Geo (

2)

Surv

(3)

CA

D S

urv

ey

Ware

house (

3)

Cante

en (

4)

Cle

anin

g (

2)

Security

Guard

(5)

Pro

cess

Met

S/F

(5)

Chem

ist

Main

t

Pro

cess O

p (

10)

Lab T

ech (

2)

Ele

c(5

)In

st (5

)M

echanic

(8)

Appendix 14A –

Financial Model

Ke

vit

sa

P

roje

ct

Pro

du

cti

on

Su

mm

ary

an

d P

re-t

ax

DC

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qu

ity

Sta

rt o

f 2

00

6 M

on

ey

Te

rms

ITE

MY

ea

r20

06

20

07

20

08

20

09

20

10

20

11

20

12

20

13

20

14

20

15

20

16

20

17

20

18

20

19

20

20

20

21

20

22

20

23

20

24

TO

TA

L

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ar

No

.-3

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12

34

56

78

910

11

12

13

14

15

16

Min

ed

(t)

Ore

at

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f 0

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(g

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70

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50

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20

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0

Pt

(g/t

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50

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30

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10

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0.3

30

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67

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66

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67

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66

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66

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66

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66

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67

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67

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67

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67

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66

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66

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68

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Cu

15

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15

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15

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16

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15

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15

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16

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15

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15

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15

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15

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15

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15

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15

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15

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14

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65

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65

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66

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64

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65

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65

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65

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65

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65

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65

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65

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65

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66

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65

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65

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67

.00

Au

18

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19

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7.9

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8.5

18

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8.5

18

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8.7

18

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8.6

18

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8.3

18

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9.5

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35

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5.9

37

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5.1

36

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6.2

35

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6.2

36

.43

6.7

36

.83

6.4

37

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5.9

36

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Pd

33

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3.9

35

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3.3

34

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4.1

33

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4.1

34

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4.5

34

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93

.70

4.8

53

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84

.34

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24

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4.4

44

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50

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0.5

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0.5

50

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90

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0.5

50

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0.5

50

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0.5

50

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40

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(g/t

Au

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61

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61

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01

.94

1.6

61

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1.7

91

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21.6

61

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1.3

21

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(g/t

Pt)

7.9

7.9

7.7

7.6

6.8

7.0

7.4

7.6

6.7

7.2

6.9

6.4

6.3

5.3

4.7

4.3

(g/t

Pd

)4

.94

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5.1

4.2

4.3

4.3

4.4

3.8

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3.9

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60

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60

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0.0

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0.0

60

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0.0

60

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0.0

60

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Au

21

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1.0

21

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1.0

21

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1.0

21

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1.0

21

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1.0

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4.5

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4.5

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4.5

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17

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a d

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of

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0%

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nsitiv

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acto

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tes:

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ing

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09

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11

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12

20

13

20

14

20

15

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18

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19

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22

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41

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67

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79

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82

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89

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5,9

26

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3,4

26

0

Ni %

0.2

85

0.2

85

0.3

20

0.2

72

0.2

96

0.2

90

0.2

83

0.2

91

0.2

94

0.3

00

0.3

02

0.2

94

0.3

06

0.2

85

0.2

90

0.3

30

0.0

00

Cu

%0

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20

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20

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90

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70

.37

90

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80

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50

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50

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20

.46

60

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20

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50

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40

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20

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30

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50

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0

Co

%0

.01

30

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30

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40

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30

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40

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40

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40

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40

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40

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50

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40

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40

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40

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40

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60

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70

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0

Pt

g/t

0.3

49

0.3

49

0.3

71

0.3

25

0.3

09

0.3

14

0.3

26

0.3

42

0.3

03

0.3

30

0.3

19

0.2

88

0.2

92

0.2

32

0.2

10

0.2

10

0.0

00

Pd

g/t

0.2

31

0.2

31

0.2

68

0.2

31

0.2

01

0.2

03

0.2

00

0.2

10

0.1

82

0.1

96

0.1

96

0.1

86

0.1

92

0.1

53

0.1

40

0.1

40

0.0

00

Au

g/t

0.1

35

0.1

35

0.1

56

0.1

38

0.1

42

0.1

40

0.1

67

0.1

45

0.1

27

0.1

60

0.1

58

0.1

43

0.1

51

0.1

20

0.1

16

0.1

19

0.0

00

E705 Keivitsa Owner Mining Cost Study

Capital costs ($000's unless otherwise stated)Period 1 2 3 4 5Trucks 4800 1200 0 4800 0Loaders 1200 1200 0 1200 0Dozer 850 850 0 0 0Grader 675 675 0 0 0Drills 3300 1100 0 1100 0Total main fleet 10825 5025 0 7100 0Light vehicles 125 0 0 0 125Ancillary 500 0 0 0 0contingency 10% 1145 502.5 0 710 12.5Capex total ($000's) 12595 5527.5 0 7810 137.5 (Euro000's) 10391.91419 4560.644 0 6443.894 113.4488

Operating costs ($000's unless otherwise stated)Period 1 2 3 4 5Equipment maintenance 1168.5254 2081.07 1890.338 2456.816 2505.028Fuel 3699.059354 6430.834 5226.996 7783.871 8062.594Explosives 2625 4603.785 3366.907 5221.143 5229.181Drilling consumables 159.7222222 280.1245 204.8647 317.6886 318.1777Miscellaneous 60 60 60 60 60Sub total 7712.306976 13455.81 10749.11 15839.52 16174.98Salaries 5691.6 7066.5 7066.5 9367.5 9367.5Overall total 13403.90698 20522.31 17815.61 25207.02 25542.48Cost/t moved ($) 1.276562569 1.114426 1.322847 1.206969 1.221151 (Euro) 1.05 0.92 1.09 1.00 1.01

including capex element (Euro) 1.19 1.05 1.22 1.13 1.14

NPV of mining equipment (€ 000's) 29,604

( €/t moved) 0.13

Discount rate 7.00%

6 7 8 9 10 11 12 13 140 1200 1200 6000 1200 0 4800 0 00 0 0 1200 0 0 1200 0 00 0 850 850 0 0 0 0 00 0 675 675 0 0 0 0 00 0 0 0 0 1100 0 1100 00 1200 2725 8725 1200 1100 6000 1100 00 0 0 125 0 0 0 125 0

500 0 0 0 0 500 0 0 050 120 272.5 885 120 160 600 122.5 0

550 1320 2997.5 9735 1320 1760 6600 1347.5 0453.7954 1089.109 2473.185 8032.178 1089.109 1452.145 5445.545 1111.799 0

6 7 8 9 10 11 12 13 142553.07 2622.147 2674.201 2897.667 1723.435 1760.077 1790.421 1820.859 1607.645

8340.336 8739.683 9040.623 9921.345 4875.865 5087.701 5263.128 5439.097 5473.0785235.626 5253.11 5191.832 5257.102 2286.122 2317.289 2292.445 2270.66 2265.549318.5698 319.6337 315.9051 319.8766 139.1027 140.9991 139.4874 138.1618 137.8508

60 60 60 60 60 60 60 60 6016507.6 16994.57 17282.56 18455.99 9084.525 9366.066 9545.482 9728.778 9544.1229367.5 9751 9751 10518 6683 6683 6683 6683 6683

25875.1 26745.57 27033.56 28973.99 15767.52 16049.07 16228.48 16411.78 16227.121.235531 1.272845 1.301735 1.37785 1.724265 1.731448 1.769779 1.80694 1.790639

1.02 1.05 1.07 1.14 1.42 1.43 1.46 1.49 1.48

1.15 1.18 1.21 1.27 1.56 1.56 1.59 1.62 1.61

15 16 Total0 0 252000 0 60000 0 34000 0 27000 0 77000 0 450000 0 5000 0 15000 0 47000 0 517000 0 42656.77

15 16 Total1393.008 1055.865 32000.174232.214 2694.295 100310.71481.626 786.489 55683.8690.15182 47.85515 3388.172

60 60 9607257 4644.505 192342.9

6299.5 4765.5 122427.113556.5 9410.005 314770

2.287437 2.991143 1.4132011.89 2.47 1.17

2.02 2.60 1.30

Period

Tota

l-3

-2-1

14

711

16

Full

Feasi

bili

ty S

tudy

2.0

00

1.0

00

1.0

00

Land A

cquis

itio

n0.3

50

0.3

50

Min

ing E

quip

ment

0.4

50

0.1

00

0.3

50

Pro

cess

Pla

nt

42.6

20

18.9

50

23.6

70

Taili

ngs

Faci

lity

23.6

90

5.1

40

5.8

30

3.9

10

4.6

90

4.1

20

0.0

00

Infr

ast

ruct

ure

6.8

00

6.5

40

0.2

60

Indirect

s4.5

00

4.5

00

EPCM

4.0

00

2.0

00

2.0

00

Contingency

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Pro

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Days

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Pro

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Appendix 15A – Qualified Person’s Letters and Certificates

CSA Consulting International 9 John Street

London WC1N 2ES Tel: +44-20-7404-2221 Fax: +44-20-7404-2299

Email: [email protected]

17 July 2006

The Directors

Scandinavian Minerals

172 King St E.

3rd

Floor

Toronto

Canada

M5A 1J3

Dear Sirs,

(43-101/8.3) Consent of Qualified Persons in Respect of the Kevitsa Project, Finland

I, Pierre Fourie, in my capacity as qualified person consent to:

1. The filing of the technical report;

2. The written disclosure of the technical report;

3. Extracts from or a summary of the technical report in the written disclosure being

filed.

A certificate confirming that:

1. I have read the written disclosure;

2. I do not have any reason to believe that there are any misrepresentations in the

information derived from the technical report or that the written disclosure contains

any misrepresentation of the information contained in the technical report appended to

this letter.

Yours Faithfully,

CSA Consulting International 9 John Street

London WC1N 2ES Tel: +44-20-7404-2221 Fax: +44-20-7404-2299

Email: [email protected]

CERTIFICATE OF QUALIFIED PERSON (43-101 SECT 8.1)

KEVITSA POLYMETALLIC PROJECT, FINLAND

Qualified Person: Pierre Fourie

Address: 29 Graham Rd, Midrand, South Africa

Occupation: Consulting Mining Engineer

Qualifications: BEng Mining Engineering BCompt

Professional Institutions: SAIMM (701060) ECSA (965043)

Relevant experience: 11 years experience in Mineral Reserve estimation for both open pit and underground mines on various commodities (including PGM’s, Au, Ni, Cu, Co, Fe).

I, Pierre Fourie, state that by virtue of the qualifications, experience and membership of professional institutions listed above that I am “qualified person” for the purposes of Canadian National Instrument 43-101 “Standards of Disclosure for Mineral Projects.”

I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with this Instrument and Form 43-101F1.

I am responsible for the following sections of the Technical Report:

SummaryMining Effective Date and Signatures

I am not aware of any material fact or material change with respect to the subject matter of the Technical Report, which is not reflected in the Technical Report, the omission to disclose which makes the technical report misleading.

I am independent of the issuer applying the tests set out in section 1.5 of National Instrument 43-101.

I have had no prior involvement with the property that is the subject of the Technical Report other than the production of the Technical Report dated 10th March 2006.

17 July 2006

kevitsa pre-feasibility study · 7/1/2006 · pit mining operation at sml’s kevitsa property in...

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