Form 6-K
 
SECURITIES AND EXCHANGE COMMISSION
 
Report of Foreign Issuer
Pursuant to Rule 13a-16 or 15d-16 of
The Securities Exchange Act of 1934

For the month of October, 2007	 	Commission File Number: 1-15142	
 
NORTH AMERICAN PALLADIUM LTD.
(Name of Registrant)
 
130 Adelaide Street West
Suite 2116
Toronto, Ontario
Canada  M5H 3P5
(Address of Principal Executive Offices)
 
Indicate by checkmark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F:

 	Form 20-F    o	 	Form 40-F    x	 	
 
Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1):    o
 
Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7):    o
 
Indicate by checkmark whether the registrant, by furnishing the information contained in this Form is also thereby furnishing the information to the SEC pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934:

 	Yes    o    Assigned File No.                                      	No    x	 	
 
If “Yes” is marked, indicate the file number assigned to the Registrant in connection with Rule 12g3-2(b).

 
This report on Form 6-K is specifically incorporated by reference into North American Palladium’s registration statement on Form S-8 (File No. 333-13766), which was originally filed with the Securities and Exchange Commission in July 2001.

SIGNATURES
 
Pursuant to the requirements of the Securities Exchange Act of 1934, the Registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

 	 	NORTH AMERICAN PALLADIUM LTD.	

Date:	  October 31, 2007	 	By:	  /s/ Trent Mell	
 	 	 	 	  Trent Mell
  VP, General Counsel and Corporate
  Secretary	
 
2

 
EXHIBIT INDEX

Exhibit	 	Description of Exhibit	
 	 	 	
1	 	43-101 Technical Report – “Technical Report on the Suhanko Project, Northern Finland”	
23.1	 	Qualified Person Consent – F. Brown	
23.2	 	Qualified Person Consent – E. Puritch	
23.3	 	Qualified Person Consent – W. Ewert	
23.4	 	Qualified Person Consent – J. Rickard	
23.5	 	Qualified Person Consent – D. King	
 
3

Exhibit 1
 
TECHNICAL REPORT,
MINERAL RESOURCE ESTIMATE,
AND
PRELIMINARY ECONOMIC ASSESSMENT
(SCOPING STUDY)
OF THE
SUHANKO PROJECT
NORTHERN FINLAND
 
FOR
NORTH AMERICAN PALLADIUM LTD
 
BY
 
AKER KVAERNER,
P&E MINING CONSULTANTS INC.,
AND
F.H. BROWN, CPG, PR. SCI. NAT
 
NI-43-101 & 43-101F1
Technical Report and Scoping Study
Report No 135
 
Mr. Eugene Puritch P.Eng
Dr. Wayne Ewert P.Geo
Mr. F.H Brown CPG, Pr. Sci. Nat 
 Mr. Jason Rickard P.Geo
Mr David King P.Eng
 
Effective Date: September 01, 2006
Report Date/Signing Date: October 29, 2007

 
TABLE  OF  CONTENTS

EXECUTIVE SUMMARY	i	
 	 	 	 	 	
1.0	INTRODUCTION AND TERMS OF REFERENCE	1	
 	1.1	INTRODUCTION	1	
 	1.2	SOURCES OF INFORMATION	2	
 	1.3	UNITS AND CURRENCY	2	
 	1.4	USE OF REPORT	2	
 	1.5	GLOSSARY OF TERMS	3	
 	 	 	 	 	
2.0	RELIANCE ON OTHER EXPERTS	6	
 	 	 	 	 	
3.0	PROPERTY DESCRIPTION AND LOCATION	7	
 	3.1	LOCATION	7	
 	3.2	TENURE	7	
 	 	3.2.1	PREVIOUS AND CURRENT TENEMENT OWNERSHIP	13	
 	3.3	ENVIRONMENTAL PERMITING	14	
 	 	3.3.1	SUHANKO MINING LEASE AND ENVIRONMENTAL PERMIT CURRENT STATUS	14	
 	3.4	ENVIRONMENTAL OBLIGATIONS	16	
 	 	 	 	 	
4.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY	18	
 	4.1	LOCATION, AND ACCESS	18	
 	4.2	CLIMATE	18	
 	4.3	INFRASTRUCTURE	18	
 	4.4	PHYSIOGRAPHY	18	
 	 	 	 	 	
5.0	HISTORY	 	20	
 	5.1	PAST EXPLORATION KONTTIJÄRVI	20	
 	5.2	PAST EXPLORATION AHMAVAARA	21	
 	5.3	APP - GFAP ACTIVITY	21	
 	5.4	HISTORICAL DRILLING	21	
 	 	5.4.1	DRILL CORE MANAGEMENT	22	
 	5.5	PAST PRODUCTION	22	
 	5.6	MINERAL RESOURCE ESTIMATES	23	
 	 	5.6.1	KONTTIJÄRVI	24	
 	 	5.6.2	AHMAVAARA	25	
 	 	 	 	 	
6.0	GEOLOGY	 	26	
 	6.1	GENERAL GEOLOGICAL SETTING	26	
 	6.2	REGIONAL GEOLOGICAL SETTING	26	
 	6.3	LOCAL GEOLOGICAL SETTING	27	
 	6.4	KONTTIJÄRVI PROPERTY	27	
 	 	6.4.1	KONTTIJÄRVI DEPOSIT	28	
 	6.5	AHMAVAARA PROPERTY	29	
 	 	6.5.1	AHMAVAARA DEPOSIT	29	
 	 	 	 	 	
7.0	DEPOSIT TYPES	32	
 	7.1	KONTTIJÄRVI AND AHMAVAARA DEPOSIT TYPE AND MODEL	32	
 	7.2	GENETIC MODELS	32	
 	 	 	 	 	
8.0	MINERALIZATION	35	
 	8.1	OVERVIEW AND BASE METAL SULPHIDES	35	
 	8.2	PLATINUM GROUP MINERALS	36	

 

9.0	EXPLORATION	 	37	
 	9.1	GEOLOGICAL DATA COMPILATION	37	
 	9.2	MINERALOGICAL AND GEOCHEMISTRY STUDY	38	
 	9.3	SAMPLING AND METHODS	38	
 	 	 	 	 	
10.0	DRILLING	 	40	
 	 	 	 	 	
11.0	SAMPLING METHODS AND APPROACH	41	
 	 	 	 	 	
12.0	SAMPLE PREPARATION, ANALYSIS AND SECURITY	42	
 	12.1	SAMPLE PREPARATION AND ANALYSIS	42	
 	12.2	SAMPLE SECURITY	42	
 	 	 	 	 	
13.0	DATA VERIFICATION	43	
 	 	 	 	 	
14.0	ADJACENT PROPERTIES	46	
 	14.1	REGIONAL METALLOGENIC SETTING	46	
 	 	14.1.1	MINERALIZATION OF THE PORTIMO COMPLEX	46	
 	 	 	 	 	
15.0	MINERAL PROCESSING AND METALLURGICAL TESTING	51	
 	15.1	INTRODUCTION	51	
 	15.2	SUMMARY	52	
 	 	15.2.1	SAMPLE SELECTION AND PREPARATION	52	
 	 	15.2.2	MINERALOGY	53	
 	 	15.2.3	COMMINUTION	54	
 	 	15.2.4	PHASE 4 FLOTATION TESTWORK	55	
 	 	15.2.5	PHASE 5 FLOTATION TESTWORK	57	
 	 	15.2.6	GTK PILOT PLANT TESTWORK	58	
 	 	15.2.7	COMPARISON WITH PREVIOUS TEST PHASES	59	
 	 	15.2.8	ANCILLARY TESTWORK	60	
 	 	15.2.9	RECOMMENDED ADDITIONAL TESTWORK AND PROCESS REVIEW	60	
 	 	15.2.10	ALTERNATIVE PROCESS REVIEW	61	
 	 	 	 	 	
16.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES	63	
 	16.1	INTRODUCTION	63	
 	16.2	2006 NI 43-101 MINERAL RESOURCE ESTIMATE	63	
 	16.3	KONTTIJÄRVI PROPERTY	64	
 	 	16.3.1	DATA SUPPLIED	64	
 	 	16.3.2	CONFIRMATION SAMPLES	64	
 	 	16.3.3	GEOLOGICAL MODEL	64	
 	 	16.3.4	MINERALIZATION MODEL	66	
 	 	16.3.5	COMPOSITING	66	
 	 	16.3.6	GRADE ELEMENTS	66	
 	 	16.3.7	SPECIFIC GRAVITY	66	
 	 	16.3.8	SUMMARY STATISTICS	67	
 	 	16.3.9	VARIOGRAPHY	70	
 	 	16.3.10	BLOCK MODEL	72	
 	 	16.3.11	CUT-OFF GRADES	73	
 	 	16.3.12	ESTIMATION	73	
 	 	16.3.13	CLASSIFICATION	75	
 	 	16.3.14	MODEL VALIDATION	76	
 	 	16.3.15	KONTTIJÄRVI MINERAL RESOURCE ESTIMATE	78	
 	 	16.3.16	KONTTIJÄRVI CONCLUSIONS	79	
 	16.4	AHMAVAARA PROPERTY	80	

 

 	 	16.4.1	DATA SUPPLIED	80	
 	 	16.4.2	GEOLOGICAL MODEL	80	
 	 	16.4.3	MINERALIZATION MODEL	81	
 	 	16.4.4	GRADE ELEMENTS	81	
 	 	16.4.5	COMPOSITING	82	
 	 	16.4.6	SPECIFIC GRAVITY	82	
 	 	16.4.7	SUMMARY STATISTICS	82	
 	 	16.4.8	VARIOGRAPHY	87	
 	 	16.4.9	BLOCK MODEL	91	
 	 	16.4.10	CUT-OFF GRADES	92	
 	 	16.4.11	ESTIMATION	92	
 	 	16.4.12	CLASSIFICATION	94	
 	 	16.4.13	MODEL VALIDATION	94	
 	 	16.4.14	AHMAVAARA MINERAL RESERVES ESTIMATES	96	
 	 	16.4.15	CONCLUSIONS	97	
 	 	 	 	 	
17.0	OTHER RELEVANT DATA AND INFORMATION	99	
 	 	 	 	 	
18.0	INTERPRETATION AND CONCLUSIONS	100	
 	 	 	 	 	
19.0	RECOMMENDATIONS	101	
 	19.1	INTRODUCTION	101	
 	19.2	RECOMMENDATIONS	101	
 	19.3	PROPOSED BUDGET	102	
 	 	19.3.1	PROPOSED 2007 EXPLORATION PROGRAM	102	
 	 	 	 	 	
20.0	REFERENCES	 	103	
 	 	 	 	 	
21.0	CERTIFICATES OF AUTHORS	106	
 	 	 	 	 	
SECTION 2:	ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT AND PRODUCTION PROPERTIES - PRELIMINARY ECONOMIC ASSESSMENT OF THE SUHANKO PROJECT	111	
 	 	 	 	 	
22.0	MINING OPERATIONS	112	
 	22.1	INTRODUCTION AND DESIGN PARAMETERS	112	
 	22.2	CUT-OFF GRADE	115	
 	22.3	PIT DESIGN	115	
 	22.4	DILUTION	121	
 	22.5	MINEABLE PORTION OF THE MINERAL RESOURCES	122	
 	22.6	ADDITIONAL MINEABLE RESOURCES	122	
 	22.7	MINE PLANNING BASIS	122	
 	22.8	MINE PRODUCTION SCHEDULE	123	
 	22.9	PROPOSED MINE OPERATIONS	125	
 	22.10	MINE EQUIPMENT SELECTION	127	
 	22.11	MINE LABOUR REQUIREMENTS PER DAY	128	
 	22.12	OTHER ON-SITE SUPPLIERS AND CONTRACTORS	128	
 	 	 	 	 	
23.0	MINERALOGY DEPORTMENT	129	
 	23.1	INTRODUCTION	129	
 	23.2	PGE MINERALOGY	129	
 	 	23.2.1	BACKGROUND	129	
 	 	23.2.2	SAMPLING AND METHODS	130	
 	 	23.2.3	PGE DISTRIBUTION AND GRAIN SIZE	130	
 	 	23.2.4	PGE TYPES	132	
 	 	23.2.5	PGE – MINERAL ASSOCIATIONS	132	

 

 	23.3	SULPHIDE MINERALOGY	133	
 	 	23.3.1	SULPHIDE MINERALS	133	
 	 	23.3.2	SULPHIDE LIBERATION	134	
 	 	23.3.3	SULPHIDE OXIDATION	135	
 	23.4	SGS LAKEFIELD - BULK MODAL ANALYSES (BMA)	136	
 	23.5	WHOLE ROCK GEOCHEMISTRY	138	
 	23.6	COMPARISON TO THE LAC DES ÎLES DEPOSIT	138	
 	 	23.6.1	SILICATES AND SULPHIDES	138	
 	 	23.6.2	PGE MINERALS	139	
 	 	23.6.3	GEOCHEMISTRY	140	
 	23.7	SUMMARY POINTS	140	
 	 	23.7.1	KONTTIJÄRVI AND AHMAVAARA (SUHANKO)	140	
 	 	23.7.2	COMPARISON TO THE LAC DES ÎLES DEPOSIT	142	
 	 	 	 	 	
24.0	METALLURGICAL TESTWORK	144	
 	24.1	PREVIOUS TESTWORK	144	
 	24.2	RECENT TESTWORK	144	
 	 	24.2.1	TRADE-OFF STUDY	144	
 	 	24.2.2	SCOPING STUDY	152	
 	 	 	 	 	
25.0	CAPACITY TRADE-OFF EVALUATION	158	
 	25.1	INTRODUCTION	158	
 	25.2	CAPITAL COST ESTIMATES	158	
 	25.3	MINE PRODUCTION SCHEDULES AND OPERATING COSTS	158	
 	25.4	PROCESS OPERATING COST ESTIMATES	161	
 	25.5	FINANCIAL ANALYSIS	161	
 	25.6	COMPARISON OF THE OPTIONS	162	
 	 	 	 	 	
26.0	PROCESS DESIGN	163	
 	26.1	BASE CASE – SAG MILL	163	
 	26.2	ALTERNATIVE CASE - HIGH-PRESSURE ROLLS CRUSHING	164	
 	 	 	 	 	
27.0	INFRASTRUCTURE	167	
 	 	 	 	 	
28.0	CAPITAL COST ESTIMATE	168	
 	28.1	CAPITAL COST SUMMARY	168	
 	28.2	BASIS OF ESTIMATE	170	
 	28.3	DIRECT COST ELEMENTS	171	
 	28.4	INDIRECT COST ELEMENTS	171	
 	28.5	CONTINGENCY	172	
 	28.6	QUALIFICATIONS AND EXCLUSIONS	172	
 	 	28.6.1	QUALIFICATIONS	172	
 	 	28.6.2	EXCLUSIONS	172	
 	28.7	PIT CAPITAL COST	172	
 	 	 	 	 	
29.0	OPERATING COST ESTIMATE	174	
 	29.1	MINING 174	 	
 	29.2	PROCESS PLANT	174	
 	 	29.2.1	LABOUR	175	
 	 	29.2.2	REAGENTS AND CONSUMABLES	175	
 	 	29.2.3	ELECTRICAL POWER	176	
 	 	29.2.4	MAINTENANCE SUPPLIES	177	
 	 	29.2.5	CONTINGENCY	177	
 	29.3	GENERAL AND ADMINISTRATION (G&A)	177	

 

30.0	FINANCIAL ANALYSIS	178	
 	30.1	INTRODUCTION	178	
 	30.2	METAL RECOVERIES AND MASS PULLS	178	
 	30.3	INFORMATION FROM OWNER	178	
 	30.4	INFORMATION FROM P&E	179	
 	30.5	INFORMATION FROM FEASIBILITY STUDY	180	
 	30.6	FINANCIAL RESULTS	180	
 	30.7	SENSITIVITY ANALYSIS	180	
 	 	 	 	 	
31.0	RISK ASSESSMENT	182	
 	31.1	POLITICAL/SOCIAL FACTORS	182	
 	31.2	PERMITTING/APPROVALS	182	
 	31.3	GEOLOGICAL COMPLEXITY	182	
 	31.4	COMMODITY PRICES/EXCHANGE RATES	183	
 	31.5	MINERAL RESOURCES	183	
 	31.6	MINING	183	
 	31.7	METALLURGY	183	
 	31.8	CAPITAL COST/OPERATING COST	184	
 	31.9	SALEABLE CONCENTRATE	184	
 	 	 	 	 	
32.0	RECOMMENDATIONS FOR FUTURE WORK	185	
 	32.1	OREBODY ISSUES	185	
 	32.2	MINERALOGY/PLANT DESIGN	185	
 	32.3	MARKETING	186	
 	 	 	 	 	
APPENDICES	 	 	187	
 	 	 	 	 	
APPENDIX 1A	KONTTIJÄRVI 2PGE + AU RESOURCE INVENTORY	188	
APPENDIX 1B	KONTTIJÄRVI PDEQ RESOURCE INVENTORY	190	
APPENDIX 1C	AHMAVAARA 2PGE + AU RESOURCE INVENTORY	192	
APPENDIX 1D	AHMAVAARA PDEQ RESOURCE INVENTORY	196	

 

List of Figures	 	
 	 	
3-1	Location of Suhanko Project	13	
3-2	GFAP Regional Claim Holdings	15	
3-3	GFAP Environmental Permit Areas	17	
4-1	Detailed Access Map	19	
5-1	Core Processing System	23	
6-1	Suhanko Project Regional Geology	27	
6-2	Cross Section of the Konttijärvi Body Showing Fault Blocks	29	
6-3	Schematic Cross Section - Ahmavaara Deposit Showing Mineralized Horizons	31	
11-1	Sample Preparation and Analysis Scheme	41	
14-1	Suhanko and Konttijärvi Intrusions showing PGE Mineralized Occurrences	48	
14-2	Metallogenic Zones of Finland	49	
14-3	N-E Fennoscandian Shield with Locations of Layered Mafic Intrusive Complexes	50	
16-1	Konttijärvi Geological Model	65	
16-2	Konttijärvi Fault blocks	66	
16-3	Konttijärvi Grade Probability Graphs	68	
16-4	Konttijärvi Palladium Correlation	70	
16-5	Konttijärvi Experimental Semi-Variograms	72	
16-6	Block Classification (Section 53802.5, looking east)	76	
16-7	Ahmavaara Schematic mineralization zones. Not to scale	81	
16-8	Ahmavaara Grade Probability Graphs	83	
16-9	Ahmavaara Experimental Semivariogram Models	89	
16-10	Ahmavaara Eastern and Western Resource Zones	93	
16-11	Ahmavaara Block Classification (Section 30855, looking east)	95	
22-1	Konttijärvi Deposit Ultimate Pit Plan	117	
22-2	Ahmavaara Deposit Ultimate Pit Plan	118	
22-3	Konttijärvi Deposit Section 54350 E (Looking Grid East)	119	
22-4	Konttijärvi Deposit Section 54000 E (Looking Grid East)	119	
22-5	Ahmavaara Deposit Section 30450 E (Looking Grid East)	120	
22-6	Ahmavaara Deposit Section 30900 E (Looking Grid East)	120	
22-7	Ahmavaara Deposit Section 31575 E (Looking Grid East)	121	
22-8	Suhanko Project Site Layout	126	
23-1	PGE Minerals within the Suhanko Deposits	131	
23-2	PGE Mineral Associations	133	
23-3	Sulphide Mineral Textures	135	
23-4	Sulphide Content	139	
23-5	Suhanko & LDI PGE Sulphide Associations	140	
24-1	Mass Pull vs Recovery	145	
24-2	Konttijärvi Mineralization	146	
24-3	Ahmavaara Mineralization	147	
24-4	Flash Flotation Bulk Rougher Float	148	
24-5	Metal Content in Flash Concentrate	149	
24-6	Two-stage Grinding	149	
24-7	Percentage Mass Pull Versus Recovery for the Ahmavaara Sample	150	
24-8	Mass Pull Versus Grade for the Ahmavaara Sample	150	
24-9	Mass Pull Versus Recovery for the Konttijärvi Sample	150	
24-10	Percentage Mass Pull versus Grade for the Konttijärvi Sample	150	
24-11	Metal Distribution of each Concentrate for Both Samples	151	
24-12	Results of Magnetic Separation on the Copper Cleaner Tails for the Ahmavaara Copper Concentrate	152	
24-13	Results from Magnetic Separation on a Bulk Concentrate	152	
24-14	Gravity Separation of Cu Cone	152	
24-15	Influence of Regrinding on Metal Recovery	154	
24-16	Grade % of Metals after Regrinding	154	
24-17	Copper Pre-float Impact on Recoveries	155	
24-18	Process Flow Sheet used for LCT on Samples	157	
26-1	DWG-000-F-001 Potential Process Flow Diagram	165	
26-2	DWG-000-G-001 Concentrate General Arrangement	166	
30-1	Sensitivity Analysis	181	

 

List of Tables	 	
 	 	 	
1.1	NAP 2006 Global Resource Inventory Suhanko Project (1.0 g/t PdEq Cut-off)	ii	
1.2	Estimated, Diluted, Potentially Mineable Portion of the Mineral Resources First 10 Years of Mine Life	iii	
1.3	Total Mineable Portion of the Mineral Resources Contained within the Ultimate Pit Shells, Ahmavaara and Konttijärvi Deposits	iv	
1.4	Suhanko Project Production Schedule – First 10 Years of Mine Life	vi	
1.5	Suhanko Project – Project Cost Summary ears of Production	viii	
1.6	Process Plant Operating Costs	ix	
1.7	Financial Results 100% Debt, Pre-tax Basis, First 10 years of Production	ix	
1.8	Effect of 50% Equity Financing, Pre-tax Basis, First 10 Years of Production	ix	
3.1	Gold Fields Arctic Platinum Oy – List of Tenements as at 23 October 2006	8	
5.1	GFAP Drilling Since Inception to June 2004	21	
5.2	GFAP 2004 2PGE+AU Recoverable Resource Inventory – Konttijärvi Deposit	25	
5.3	GFAP 2004 2PGE+AU Recoverable Resource Inventory – Ahmavaara Deposit	25	
6.1	Stratigraphy of the Suhanko-Konttijärvi Intrusion	28	
6.2	Stratigraphic Unit of the Ahmavaara Deposit	30	
13.1	Konttijärvi Verification Samples – Brown 2006	44	
13.2	Ahmavaara Verification Sampling, 2006	45	
14.1	Metal Concentrations and Ratios of Various Prospects in Koillismaa Complex	47	
15.1	Main Flotation Design Criteria	56	
16.1	Drilling History	64	
16.2	Modeling lithology	65	
16.3	Specific Gravity Values	67	
16.4	Domain Composite Statistics	67	
16.5	Grade Element Correlation and Ratios	70	
16.6	South block short-range grade continuity	71	
16.7	South block estimation semi-variograms	71	
16.8	NAP 2006 block model	73	
16.9	PdEq NSR parameters	73	
16.10	Estimation Domains	74	
16.11	NAP 2006 Global Resource Inventory – Konttijärvi and Ahmavaara Deposits	74	
16.12	Marginal Domain mineralization - composite samples	75	
16.13	Comparison of composite averages and model estimates	77	
16.14	Comparison of OK and NN block model estimates	77	
16.15	Comparison of blasthole averages and NAP block model estimates	78	
16.16	Average informing distance of the closest sample for Pd	78	
16.17	GFAP Feasibility Study Pit Design Resource Inventory	78	
16.18	NAP 2006 Pit Design 2PGE+Au Resource Inventory	78	
16.19	NAP 2006 Pit Design PdEq Resource Inventory	79	
16.20	Comparison of Measured and Indicated 2PGE+Au resources (1.0 g/t cut-off)	79	
16.21	Drilling history	80	
16.22	Modeling lithology	81	
16.23	Mineralization codes	81	
16.24	Average composite and accumulation values	82	
16.25	Specific gravity values	82	
16.26	Mineralization zone composite grade statistics	87	
16.27	Grade element correlation and Pd ratios (1m composites)	88	
16.28	Ahmavaara estimation semi-variograms	88	
16.29	NAP 2006 block model	92	
16.30	PdEq NSR parameters - Ahmavaara	92	
16.31	NAP 2006 PdEq resource inventory	93	
16.32	Lithological unit mineralization – 1.0 g/t 2PGE+Au cut-off (Meas. + Indic + Infer)	94	
16.33	Comparison of composite averages and model estimates	95	
16.34	Comparison of OK and NN block model estimates	96	
16.35	Comparison of blasthole averages and NAP block model estimates	96	
16.36	Average informing distance of the closest sample for Pd	96	
16.37	GFAP Feasibility study pit design constrained resource inventory	96	
16.38	NAP 2006 open pit design 2PGE+Au resource inventory	97	

 

16.39	NAP 2006 open pit design PdEq resource inventory	97	
16.40	Comparison of Measured and Indicated 2PGE+Au resources (1.0 g/t cut-off)	98	
22.1	Konttijärvi Pit Design Parameters	113	
22.2	Ahmavaara Pit Design Parameters	114	
22.3	Estimated Grades of Diluting Material	121	
22.4	Estimated Diluted Potentially Mineable Portion of Mineral Resources	122	
22.5	Total Mineable Portion of the Mineral Resources Contained within the Ultimate Pit Shells, Ahmavaara and Konttijärvi Deposits	123	
22.6	Production Schedule, First 10 Years of Mine Life	124	
22.7	Estimated Haulage Truck Cycle Times	127	
22.8	Key Pit Equipment	127	
22.9	Estimated Mine Direct and Indirect Labour Summary	128	
22.10	Estimated Mine Operations Direct Labour Requirements	128	
23.1	QemScan Analyses	137	
23.2	QemScan Bulk Mineral Analyses	138	
24.1	Metallurgical Balance – Konttijärvi	148	
24.2	Metallurgical Balance – Ahmavaara	148	
24.3	Metal Recoveries	143	
24.4	Locked-cycle Program Test Results	156	
25.1	Capital Cost Estimates	158	
25.2	5 Mt/a Mine Production Schedule	159	
25.3	7.5 Mt/a Production Schedule	160	
25.4	Process Operating Cost Estimates	161	
25.5	Metal Prices	161	
25.6	Capacity Trade-off IRR	161	
28.1	Suhanko Project – Project Cost Summary	168	
28.2	Plant and Infrastructure Breakdown	169	
28.3	Estimated Pit Capital Cost – 7.5 Mt/a	173	
29.1	Estimated Operating Cost – 7.5 Mt/a	174	
29.2	Process Plant Operating Cost Summary	175	
29.3	Bulk Concentrate for Sale - Labour	175	
29.4	Reagents and Consumables – Konttijärvi	176	
29.5	Reagents and Consumables – Ahmavaara	176	
29.6	Power Costs	176	
30.1	Metal Recoveries and Mass Pulls	178	
30.2	Metal Prices for Plant Capacity 7.5 Mt/a	178	
30.3	Ore Grades	179	
30.4	Information from Feasibility Study	180	
30.5	Financial Results – 100% Debt	180	
30.6	Sensitivity Analysis	181	

 
EXECUTIVE SUMMARY
 
INTRODUCTION
 
Gold Fields Arctic Platinum OY (GFAP) is interested in developing the Suhanko PGE deposit as part of the Arctic Platinum Project (APP) located in northern Finland. Considerable work has been carried out on this project by GFAP including a final draft report for a Feasibility Study (FS) by Lycopodium that was prepared in March 2005. However, the report did not show an acceptable rate of return for the project.
 
North American Palladium (NAP) in conjunction with GFAP has retained Aker Kvaerner E&C, a division of Aker Kvaerner Canada Inc. (Aker Kvaerner), P&E and F.H. Brown to carry-out a Technical Report, Mineral Resource Estimate, and Preliminary Economic Assessment (PEA) for the Suhanko Project based on producing a bulk concentrate for sale. As part of this assignment, a Capacity Trade-off Study was carried out to evaluate 5 and 7.5 Mt/a alternatives to pick the optimum capacity for executing the PEA.
 
The base case for the process plant uses a SAG/ball milling crushing circuit and conducts a preliminary evaluation of a High Pressure Grinding Mill (HPGR)/ball milling that has been developed as an alternative.
 
This report presents the results of the Preliminary Economic Assessment (Scoping Study) that covers mineral resource estimates, mining, mineralogical studies, metallurgical testwork, processing, infrastructure, capital and operating cost estimates, preliminary financial evaluation and risk assessment. The mineral resource estimates and mining sections were prepared by P&E Mining Consultants Inc. (P&E). The mineralogy section was prepared by GFAP.
 
The reader is cautioned that the Preliminary Economic Assessment as outlined in this report contains an economic assessment that is preliminary in nature, and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary assessment will be realized.
 
MINERAL RESOURCE ESTIMATES
 
High-grade hydrothermally remobilized Platinum Group Elements (PGE) mineralization at the Suhanko Project such as that at the Konttijärvi deposit can extend up to 70 m into the underlying basement. The Suhanko Project is focused on development of the Platinum Group Elements (PGE) + Gold (Au) + Copper (Cu) + Nickel (Ni) mineralization found at the Konttijärvi and Ahmavaara deposits. Primary PGE mineralization is hosted within the Portimo Layered Mafic Intrusive Complex and consists of disseminated and locally semi-massive, base-metal PGE-rich sulphides occurring near the basal contact of the Suhanko intrusion with the Archean basement. The Portimo Intrusive Complex comprises the Suhanko and Narkaus Intrusive bodies.
 
Since its discovery by Outokumpu Mining Oy in 1981, the total amount of diamond drilling on the Konttijärvi and Ahmavaara deposits to the beginning of 2006 amounts to 119,963 m in 1,053 drillholes. In addition to drilling, the deposits have been the subject of numerous mineral processing and metallurgical studies, trial mining exercises and mineral resources estimates.
 
In 2006, an independent evaluation and a NI 43-101 compliant resource model of the Suhanko project comprising the Konttijärvi and Ahmavaara deposits, at the scoping level of confidence, was requested by North American Palladium (NAP). Using data supplied by GFAP and NAP, updated Measured, Indicated plus Inferred resource inventories for the
 
North American Palladium Ltd - Suhanko PGE Project
Preliminary Economic Assessment
 
i

 
Konttijärvi and Ahmavaara deposits, shown in Table 1.1, were prepared by F.H. Brown (2006) as follows.
 
Table 1.1: Konttijärvi and Ahmavaara Deposits 2006 Global Resource Inventory
1.0 g/t PdEq Cut Off

 	 	 	 	 	 	 	 	 	 	 	 	 	 	PdEq	 	 	 	
 	 	Tonnes	 	Pd g/t	 	Pt g/t	 	Au g/t	 	Ni %	 	Cu %	 	g/t	 	PdEq Oz	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Ahmavaara	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	38,197,000	 	0.707	 	0.141	 	0.086	 	0.061	 	0.161	 	3.425	 	4,206,000	 	
Indicated	 	114,816,000	 	0.851	 	0.175	 	0.104	 	0.071	 	0.181	 	4.027	 	14,867,000	 	
Inferred	 	34,757,000	 	0.867	 	0.187	 	0.103	 	0.070	 	0.170	 	4.015	 	4,486,000	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Konttijärvi	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	26,031,000	 	1.064	 	0.307	 	0.081	 	0.059	 	0.100	 	4.173	 	3,493,000	 	
Indicated	 	37,571,000	 	0.902	 	0.255	 	0.071	 	0.043	 	0.095	 	3.452	 	4,170,000	 	
Inferred	 	11,638,000	 	0.867	 	0.241	 	0.062	 	0.027	 	0.097	 	3.014	 	1,128,000	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Combined	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	64,228,000	 	0.852	 	0.208	 	0.084	 	0.060	 	0.136	 	3.728	 	7,699,000	 	
Indicated	 	152,387,000	 	0.864	 	0.195	 	0.096	 	0.064	 	0.160	 	3.885	 	19,037,000	 	
Inferred	 	46,395,000	 	0.867	 	0.201	 	0.093	 	0.059	 	0.152	 	3.764	 	5,614,000	 	

(1) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
 
It should be noted that the mineral resources in this estimate were calculated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council December 11, 2005.
 
MINE DESIGN
 
The Konttijärvi and Ahmavaara deposits are located approximately 3 km from each other. Net Smelter Return (NSR) block models, where metal prices, process recoveries, concentrate freight, smelter charges, refining charges, penalties and marketing costs are considered, were developed from the nickel, copper, gold, platinum and palladium grades within each pit and were assessed for potential mining recovery using the Whittle 4X pit optimizing software. The optimization process was utilized to develop initial 10-year starter pits in order to maximize early project cash flow. The Whittle NSR cut-off grade for the Konttijärvi deposit utilized in the pit optimization was USD $6.18/t and for Ahmavaara USD $6.37/t, being the sum of the estimated costs of processing and administration.
 
The Konttijärvi designed open pit has dimensions of approximately 500 m wide, 900 m long and up to 160 m deep while the Ahmavaara designed open pit has dimensions of approximately 500 m wide, 1,800 m long and up to 225 m deep.
 
During the pre-stripping period for the two open pits, waste rock that is non-acid generating could be used for tailings dam and road construction. All potentially acid generating (PAG) waste rock from the two pits will be stored near the pits and hauled back into the Konttijärvi pit after production ceases. The mineable portion of the mineral resource total is shown in Table 1.2 and is 75,000,000 tonnes that, at the proposed milling rate of 21,500 t/d, would provide a project life of 10.0 years.
 
MINEABLE PORTION OF THE MINERAL RESOURCES
 
The estimated diluted, potentially mineable portion of the mineral resources contained within the Konttijärvi and Ahmavaara pit area design perimeters are set out in Table 1.2 together
 
ii

 
with the associated waste rock quantities. The mineable portion of the mineral resource total is 75,000,000 t that at the proposed milling rate of 21,500 t/d, would provide a project life of 10.0 years.
 
Table 1.2:  Estimated Diluted Potentially Mineable Portion of the Mineral Resources for the First 10 Years of Mine Life
 
May 28/07

 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	Waste &	 	 	 	 	 	
 	 	 	 	 	 	 	 	Au	 	Pt	 	Pd	 	NSR	 	OVB	 	Total	 	Strip	 	
 	 	Tonnes	 	Ni%	 	Cu%	 	g/t	 	g/t	 	g/t	 	$ US/t	 	Tonnes	 	Tonnes	 	Ratio	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Konttijärvi	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	12,445,000	 	0.075	 	0.113	 	0.104	 	0.390	 	1.346	 	$	23.75	 	 	 	 	 	 	 	
Indicated	 	4,698,000	 	0.065	 	0.116	 	0.117	 	0.361	 	1.298	 	$	22.19	 	 	 	 	 	 	 	
Inferred	 	357,000	 	0.056	 	0.127	 	0.157	 	0.543	 	1.987	 	$	29.14	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	21,300,000	 	38,800,000	 	1.22	 	
Ahmavaara	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	17,771,000	 	0.079	 	0.197	 	0.108	 	0.187	 	0.933	 	$	21.67	 	 	 	 	 	 	 	
Indicated	 	35,370,000	 	0.090	 	0.210	 	0.134	 	0.248	 	1.166	 	$	25.57	 	 	 	 	 	 	 	
Inferred	 	4,359,000	 	0.076	 	0.171	 	0.119	 	0.239	 	1.043	 	$	22.30	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	113,400,000	 	170,900,000	 	2.01	 	
Combined	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	30,216,000	 	0.078	 	0.162	 	0.106	 	0.271	 	1.103	 	$	22.53	 	 	 	 	 	 	 	
Indicated	 	40,068,000	 	0.087	 	0.199	 	0.132	 	0.261	 	1.181	 	$	25.17	 	 	 	 	 	 	 	
Inferred	 	4,716,000	 	0.075	 	0.168	 	0.122	 	0.262	 	1.114	 	$	22.82	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	134,700,000	 	209,700,000	 	1.8	 	
 
Note: Mineral resources that are not mineral reserves do not have demonstrated economic viability. The reader is cautioned that inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that any value from such resources will be realized either in whole or in part.
 
The mineral resources in this estimate were calculated and classified using the CIM Definitions and Guidelines on Mineral Resources and Reserves, December 11, 2005.
 
The NSR values in the above table are based on 18 month trailing average prices as at November 30, 2006 and an exchange rate of USD $1.225 = 1 Euro.
 
At the metal prices utilized in the November 30, 2006 18-month trailing average metal price estimate, it was seen that nickel provides 33% total NSR value, while copper contributes 20%, gold 5%, platinum 19% and palladium 23%.
 
ADDITIONAL MINEABLE RESOURCES
 
The purpose of this study was to evaluate the Konttijärvi and Ahmavaara pits for a period of 10 years to determine the economic viability of the project. There are additional mineable resources beyond the 10 years that were not subject to mine design or production scheduling because it was deemed that the first 10 years of operations would be adequate to determine project viability.
 
Table 1.3 shows the total combined estimated diluted, potentially mineable portion of the mineral resources.
 
iii

 
Table 1.3:  Total Mineable Portion of the Mineral Resources Contained within the Ultimate Pit Shells, Ahmavaara and Konttijärvi Deposits

Category	 	Tonnes	 	Pd g/t	 	Pt g/t	 	Au g/t	 	Ni%	 	Cu%	 	
Ahmavaara Deposit	 	
Measured	 	27,230,000	 	0.900	 	0.179	 	0.108	 	0.077	 	0.199	 	
Indicated	 	87,602,000	 	1.017	 	0.211	 	0.123	 	0.084	 	0.210	 	
Sub - Total, M&I	 	114,832,000	 	0.989	 	0.203	 	0.119	 	0.082	 	0.207	 	
Inferred	 	22,168,000	 	0.988	 	0.225	 	0.108	 	0.074	 	0.168	 	
Waste & OVB	 	269,800,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	406,800,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.97	 	 	 	 	 	 	 	 	 	 	 	
Konttijärvi Deposit	 	
Measured	 	18,844,000	 	1.342	 	0.387	 	0.101	 	0.069	 	0.121	 	
Indicated	 	18,826,000	 	1.262	 	0.353	 	0.102	 	0.054	 	0.131	 	
Sub -Total, M&I	 	37,670,000	 	1.302	 	0.370	 	0.101	 	0.062	 	0.126	 	
Inferred	 	3,230,000	 	1.411	 	0.390	 	0.109	 	0.040	 	0.159	 	
Waste & OVB	 	76,300,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	117,200,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.87	 	 	 	 	 	 	 	 	 	 	 	
Combined Ahmavaara Plus Konttijärvi Deposits	 	
Measured	 	46,074,000	 	1.081	 	0.264	 	0.105	 	0.074	 	0.167	 	
Indicated	 	106,428,000	 	1.060	 	0.236	 	0.119	 	0.079	 	0.196	 	
Sub - Total, M&I	 	152,502,000	 	1.067	 	0.245	 	0.115	 	0.077	 	0.187	 	
Inferred	 	25,398,000	 	1.042	 	0.246	 	0.108	 	0.070	 	0.167	 	
Waste & OVB	 	346,100,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	524,000,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.95	 	 	 	 	 	 	 	 	 	 	 	
 
Note: Mineral resources that are not mineral reserves do not have demonstrated economic viability.
 
MINE PLANNING BASIS
 
The mine planning process began with the development of preliminary capital and operating expenditure estimates for 5.0 and 7.5 Mt/a processing rates based on pit optimizations. The results of the assessment indicated that a processing rate of 7.5 Mt/a offered an effective mining rate and lower unit costs. The results were then used to further delineate the ultimate pit limits, haul roads, crusher and waste rock dump locations, haul distances, and to identify cost saving opportunities such as:
 
•               the interactive development of Konttijärvi and Ahmavaara pits
•               the sequencing of production in order to reduce pre-stripping and increase revenues in the initial years of mining
•               the use of waste rock in tailings starter dam construction and mine/plant site roads
•               sizing the Owner-Operators’ mine equipment fleet based on typical annual total material production requirements.
 
In consideration of the above, the proposed mining operations are based on a 21,500 t/d, 350 d/a processing rate over a 10 year period. The mine planning was thus, done on an initial 10 year basis using the assumption that the first phase of mining would be considered a “starter” pit operation.
 
MINE PRODUCTION SCHEDULE
 
The mine planning process began with the development of preliminary capital and operating expenditure estimates based on 3.5, 5.0 and 7.5 Mt/a processing rates based on pit optimizations. The result of the assessment indicated that a processing rate of 7.5 Mt/a offered a more effective mining rate and lower unit costs.
 
Table 1.4 shows the sources of material scheduled for each year of the mine life with production in the first two years of the mine life coming from the Konttijärvi mine, year three from both pits and the remaining 7 years thereafter from Ahmavaara.
 
iv

 
MINERALOGICAL DEPORTMENT
 
A mineralogical study of the different lithologies that host the platinum group metals (PGE), copper, and nickel mineralization at Ahmavaara and Konttijärvi was undertaken. The purpose of this study was to identify the various PGE mineral associations, sulphide textures and grain interlocking textures. Of particular interest were the host sulphide and silicate minerals and the relative amount of PGEs locked up in silicates, as these are factors which may have an impact on metal recoveries. Other factors that could have an impact on metal recoveries were also investigated, including the grain size of the PGEs, the distribution of the PGEs within the rocks and their host minerals, the types of sulphide minerals present and the fine grained intergrowth of these sulphides with gangue minerals.
 
The study was completed by analyzing the in-situ PGE distribution within the different rock types from the Konttijärvi and Ahmavaara deposits. For simplicity, these two deposits have been referred to here as the Suhanko Project.
 
Data from various historical studies was also used to generate some of the observations contained in this section. Some of the historical information was edited and put into a format that was used to generate comparative graphs and figures between the various data sets. Care was taken, as much as possible, to avoid inconsistent comparisons.
 
METALLURGICAL TESTWORK
 
Extensive testwork was carried out on the APP prior to the involvement of Aker Kvaerner in early 2006. After review of this work, a preliminary testwork program was carried out by SGS Lakefield to evaluate several options for improving the metallurgical recoveries and concentrate grades. From this work, a flow sheet was chosen as a basis for a further metallurgical test work program that was also performed by SGS Lakefield. The work included bench-scale and locked-cycle test work. The results of this work were used to aid in the development of the design criteria of the flow sheet for this Preliminary Economic Assessment.
 
CAPACITY TRADE-OFF EVALUATION
 
As part of the PEA, NAP/GFAP requested that Aker Kvaerner carry out a Capacity Trade-off Study to evaluate 5 and 7.5 Mt/a throughput alternatives in an attempt to identify the optimum throughput capacity for executing the rest of the study. This Trade-off Study was carried out at the beginning of the PEA in mid November 2006. (see section 7.0).
 
P&E developed preliminary mine production plans for each throughput option.
 
The 7.5 Mt/a option returned a significantly better IRR, from 11.3 to 13.6% for the 18-month trailing average metal prices, and only a modest increase in capital cost of USD50.8 M. Consequently, the decision was made to select this throughput rate as the plant capacity for the PEA.
 
v

 
Table 1.4: Suhanko Project Production Schedule – First 10 Years of Mine Life

Year	 	Pit	 	Ore	 	Ni	 	Cu	 	Pt	 	Pd	 	Au	 	NSR	 	Waste	 	Overburden	 	Total	 	Waste/Ore	 	 	 	
 	 	 	 	(t)	 	(%)	 	(%)	 	(g/t)	 	(g/t)	 	(g/t)	 	(USD/t)	 	(t)	 	(t)	 	(t)	 	Ratio	 	t/d	 	
Pre-strip	 	Konttijärvi	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	1 139 134	 	2 360 866	 	3 500 000	 	 	 	70 000	 	
 	 	Konttijärvi	 	7 432 838	 	0.076	 	0.108	 	0.371	 	1.289	 	0.098	 	22.85	 	9 236 575	 	1 047 211	 	17 716 625	 	1.38	 	 	 	
 	 	Ahmavaara	 	67 162	 	0.105	 	0.159	 	0.303	 	1.207	 	0.148	 	27.12	 	432 527	 	11 533 687	 	12 033 376	 	178.17	 	 	 	
1	 	 	 	7 500 000	 	0.076	 	0.109	 	0.371	 	1.289	 	0.099	 	22.89	 	9 669 102	 	12 580 898	 	29 750 000	 	2.97	 	85 000	 	
 	 	Konttijärvi	 	7 500 000	 	0.069	 	0.106	 	0.361	 	1.252	 	0.098	 	21.63	 	6 977 223	 	 	 	14 477 223	 	0.93	 	 	 	
 	 	Ahmavaara	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	11 940 920	 	3 331 857	 	15 272 777	 	 	 	 	 	
2	 	 	 	7 500 000	 	0.069	 	0.106	 	0.361	 	1.252	 	0.098	 	21.63	 	18 918 143	 	3 331 857	 	29 750 000	 	2.97	 	85 000	 	
 	 	Konttijärvi	 	2 560 823	 	0.072	 	0.153	 	0.503	 	1.788	 	0.166	 	30.48	 	573 486	 	 	 	3 134 309	 	0.22	 	 	 	
 	 	Ahmavaara	 	4 939 177	 	0.077	 	0.150	 	0.266	 	1.110	 	0.112	 	22.14	 	20 651 232	 	1 025 282	 	26 615 691	 	4.39	 	 	 	
3	 	 	 	7 500 000	 	0.075	 	0.151	 	0.347	 	1.342	 	0.131	 	24.99	 	21 224 718	 	1 025 282	 	29 750 000	 	2.97	 	85 000	 	
4	 	Ahmavaara	 	7 500 000	 	0.078	 	0.163	 	0.276	 	1.160	 	0.125	 	23.24	 	22 250 000	 	 	 	29 750 000	 	2.97	 	85 000	 	
5	 	Ahmavaara	 	7 500 000	 	0.077	 	0.172	 	0.250	 	1.093	 	0.119	 	22.53	 	17 000 000	 	 	 	24 500 000	 	2.27	 	70 000	 	
6	 	Ahmavaara	 	7 500 000	 	0.073	 	0.173	 	0.228	 	1.030	 	0.106	 	21.18	 	8 250 000	 	 	 	15 750 000	 	1.10	 	45 000	 	
7	 	Ahmavaara	 	7 500 000	 	0.077	 	0.191	 	0.194	 	0.958	 	0.101	 	21.07	 	6 500 000	 	 	 	14 000 000	 	0.87	 	40 000	 	
8	 	Ahmavaara	 	7 500 000	 	0.095	 	0.230	 	0.174	 	0.971	 	0.111	 	23.97	 	4 750 000	 	 	 	12 250 000	 	0.63	 	35 000	 	
9	 	Ahmavaara	 	7 500 000	 	0.105	 	0.241	 	0.212	 	1.109	 	0.124	 	27.18	 	4 750 000	 	 	 	12 250 000	 	0.63	 	35 000	 	
10	 	Ahmavaara	 	7 500 000	 	0.109	 	0.302	 	0.247	 	1.281	 	0.213	 	32.14	 	975 000	 	 	 	8 475 000	 	0.13	 	24 214	 	
Total	 	 	 	75 000 000	 	0.083	 	0.184	 	0.266	 	1.148	 	0.123	 	24.08	 	115 426 097	 	19 298 903	 	209 725 000	 	1.80	 	 	 	
 
vi

 
PROCESS DESCRIPTION
 
For the purposes of this PEA, Aker Kvaerner has evaluated the following circuit at a throughput of 7.5 Mt/a:
 
•       primary crushing
•       SAG milling
•       ball milling
•       copper flotation
•       copper concentrate column cleaning
•       copper/nickel rougher and scavenger flotation
•       copper/nickel concentrate regrinding
•       copper /nickel concentrate cleaning
•       bulk concentrate thickening and filtration.
 
The preliminary plant design is based on the design criteria derived in the recent testwork carried out by SGS Lakefield.
 
In an effort to reduce operating costs, particularly power costs, the option of using a high pressure rolls crusher in place of the SAG mill in the grinding circuit has been evaluated.
 
INFRASTRUCTURE
 
The infrastructure for the Suhanko Project consists of the following elements:
 
•      Site Development
•      Main Access Road
•      Site Road
•      Tailings Storage and Reclaim
•      Power Supply
•      Main Substation Distribution
•      Fuel Storage and Distribution
•      Water Supply (Fresh, Process and Fire)
•      Sewage Treatment
•      Port and Other Costs
•      Communication
•      Mobile Equipment
•      Ancillary Buildings
 
The design for the infrastructure is primarily based on the work carried out for the Feasibility Study prepared in 2005. However, Aker Kvaerner made a visit to site in November 2006 to review the infrastructure and meet with various local organizations. As a result, changes were made to some parts of the infrastructure that would result in some cost savings.
 
CAPITAL COST ESTIMATE
 
An order of magnitude capital cost has been developed for the mining and surface facilities for this PEA. The estimated costs for the mine, concentrator and infrastructure for these options are summarized in Table 1.5 and are inclusive of the costs up to and including plant commissioning and start-up. Sustaining capital costs and deferred capital costs are excluded from these estimates as they are included in the cash flow model in the years in which they are incurred.
 
vii

 
Table 1.5
Suhanko Project – Project Cost Summary

Project Area Description	 	Cost	 	
 	 	(USD)	 	
Direct Costs	 	 	 	
Process Mechanical Equipment	 	76,157,000	 	
Factored Commodities Process Plant	 	64,738,000	 	
Infrastructure	 	103,543,000	 	
Mining	 	 	 	
•      Roads	 	515,000	 	
•      Mine Operations Complex	 	1,194,800	 	
•      Mine Maintenance Complex	 	8,749,850	 	
•      Mine Support Facilities	 	1,354,450	 	
•      Mine Support Equipment	 	3,687,400	 	
•      Mining Equipment Fleet (lease down payment)	 	6,654,830	 	
Total Direct Costs	 	266,594,000	 	
Indirect Costs	 	 	 	
Project Indirect Costs	 	62,715,600	 	
Subtotal for Contingency	 	329,308,880	 	
Contingency @ 20%	 	65,862,120	 	
Subtotal Direct and Indirect	 	395,171,000	 	
Owner Costs	 	33,621,000	 	
TOTAL PROJECT CAPITAL COST	 	428,792,000	 	
 
All capital costs are expressed in third quarter 2007 USD with no allowance for escalation, interest or financing during construction. The capital cost estimate, based on the designs presented in this PEA, has a predicted accuracy level of ±30%.
 
OPERATING COST ESTIMATE
 
MINING
 
The estimated mining cost for a 7.5 Mt/a throughput rate is USD1.84/tonne rock. The estimated cost per tonne of mineralized material based on an overall stripping ratio of 1.8 is USD $3.31/tonne processed.
 
PROCESS PLANT
 
Preliminary process plant operating cost estimates have been developed in third quarter 2007 USD for the Konttijärvi and Ahmavaara deposits. The operating costs were estimated by developing the individual costs for labour, reagents and consumables, electrical power and maintenance supplies.
 
The estimated processing plant operating costs are summarized below in Table 1.6.
 
viii

 
Table 1.6
Process Plant Operating Costs Summary

 	 	Ahmavaara	 	Konttijarvi	 	
Cost Area	 	Annual Costs	 	Costs	 	Annual Costs	 	Costs	 	
 	 	(USD/a)	 	(USD/t ore)	 	(USD/a)	 	(USD/t ore)	 	
Labour	 	4,770,810	 	0.64	 	4,770,810	 	0.64	 	
Reagents and Consumables	 	17,652,862	 	2.35	 	16,421,145	 	2.19	 	
Assay laboratory	 	257,500	 	0.03	 	257,500	 	0.03	 	
Electrical Power	 	13,262,833	 	1.77	 	13,262,833	 	1.77	 	
Maintenance Supplies	 	4,194,459	 	0.56	 	4,194,459	 	0.56	 	
Contigency 10 %	 	4,013,846	 	0.54	 	3,890,675	 	0.52	 	
TOTAL	 	44,152,310	 	5.89	 	42,797,421	 	5.71	 	
 
GENERAL AND ADMINISTRATION
 
The annual estimated cost for general and administration is USD $5.15M or USD $0.67/tonne at an annual rate of 7.5 Mt/a.
 
FINANCIAL ANALYSIS
 
A preliminary financial model for the first 10 years of the project life has been developed. The different options are shown in Table 1.7.
 
A sensitivity analysis has been carried out and the results are presented in Section 32.
 
Table 1.7
Financial Results Pre-tax Basis, First 10 years of Production

 	 	18 Month Trailing	 	24 Month Trailing	 	
IRR	 	23.9	%	19.1	%	
NPV @ 7.5% (USD)	 	501,108,752	 	337,961,416	 	
NPV @ 10% (USD)	 	359,875,902	 	225,041,891	 	
NPV @ 12.5% (USD)	 	252,143,456	 	139,727,512	 	
 
RISK ASSESSMENT
 
A risk assessment has been carried out on the Suhanko Project by considering the following factors:
 
•        Political/social
•        Permitting/approvals
•        Geological complexity
•        Commodity prices/exchange rates
•        Mineral resources
•        Mining
•        Metallurgy
•        Capital cost/operating cost estimate
•        Generation of a saleable concentrate
 
Each of these factors is described more fully in Section 31.
 
The highest risk factor appears to be associated with concentrate marketing and metallurgy and the lowest risk is associated with political/social, permitting/approvals and mineral resources.
 
ix

 
RECOMMENDATIONS FOR FUTURE WORK
 
The following recommendations have been made for future work on the project.
 
MINERAL DEPOSIT ISSUES
 
(a)           An estimated diluted, potentially mineable portion of the mineral resources has been developed, which is a significant step forward from the 2005 Lycopodium draft Feasibility Study. Additional infill drilling is required to upgrade as much as possible of the Inferred resources to the Measured and Indicated categories and this work is currently under way. Further work is then required to covert the mineral resources to reserves.
 
(b)           Additional investigations are required to better understand the mineralogical characteristics of the mineralization.
 
(c)           The PEA Scoping Study has been based on mining a portion of the Konttijärvi deposit, first followed by a portion of the Ahmavaara deposit. However, due to the variability of the mineralized lithologies, a detailed mine plan is required to optimize the metallurgical characteristics of the material that will be fed to the concentrator for each operating period of the schedule. This work is required before any additional metallurgical testwork is carried out so that the samples used will reflect the mining schedule.
 
METALLURGY/PLANT ENGINEERING
 
(a)           The mineralization appears to be difficult material from which to generate a saleable product. Piloting of the flow sheet is recommended. Particular care needs to be taken with the selection of the sample and close supervision of the sample collection and the operation of the Pilot Plant is recommended.
 
(b)           The metal recoveries used in this report seem low compared to what was achieved in previous studies. This may have been due to the use of oxidized ore samples in the testwork. It is recommended that more metallurgical testwork be done on both ore deposits using fresh samples.
 
(c)           The reagent suite needs to be optimized so as to reduce operating costs.
 
(d)           Additional mineralogical work should be done on the flotation residue to examine the location of metal values within the tailings. This should lead to improved metal recoveries.
 
(e)           It may be possible to improve the recoveries at the expense of concentrate grade. This would depend upon finding a smelter that would accept the lower grade concentrate, or the use of an alternative concentrate leach technology such as Platsol. This avenue should be explored further.
 
(f)            The use of a high-pressure rolls crusher in the primary comminution circuit seems to improve the project economics. These numbers need to be confirmed using actual crushing test data as recommended by the equipment manufacturer.
 
(g)           The effects of increasing the annual throughput rate to 10 Mt/a should be further evaluated.
 
MARKETING
 
(a)           Additional work on the marketability and smelting cost of the concentrate should be done.
 
x

 
1.0           INTRODUCTION AND TERMS OF REFERENCE
 
1.1           INTRODUCTION
 
The Suhanko Project is situated adjacent to the Arctic Circle, in Northern Finland and has as its primary focus the exploration and development of the Konttijärvi and Ahmavaara deposits. The Konttijärvi and Ahmavaara deposits are situated approximately three kilometres from each other and about sixty kilometres south of the town of Rovaniemi. In 2006 North American Palladium, Ltd. (“NAP”) entered into a option joint venture with Gold Fields Arctic Platinum Oy (“GFAP”), a 100% owned subsidiary of Gold Fields Limited (“Gold Fields”), whereby NAP can acquire a 50% interest in the project and may acquire 60% in certain circumstances. NAP is the Project Operator.
 
The following report was prepared to provide an independent NI 43-101 compliant Preliminary Economic Assessment (“PEA”) of the Suhanko Project comprising the Konttijärvi and Ahmavaara deposits, Rovaniemi region, Northern Finland (the “Property (ies)”). The primary focus of the PEA is to prepare an independent technical appraisal of the potential economic viability of the platinum group element (PGE), gold, copper and nickel mineralization present on the Konttijärvi and Ahmavaara properties in conformance with the standards required by NI 43-101 and Form 43-101F.
 
This report was jointly prepared for NAP by Aker Kvaerner, P&E and Mr. F. H. Brown, Pr. Sci. Nat. at the request of Mr. Bruce Mackie, Vice-President, Exploration and Development of North American Palladium Ltd., a Toronto based company trading on the Toronto Stock Exchange (TSX) under the symbol of “PDL” and on the American Stock Exchange under the symbol “PAL” with its corporate office at:
 
Suite 2116, 130 Adelaide Street West,
Toronto, Ontario,
Canada, M5H 3P5
 
This report is considered current as of September 1, 2006.
 
The PEA is based on an independent global resource model and estimate of the Konttijärvi and Ahmavaara deposits, prepared at a scoping level of confidence by Mr. F.H Brown CPG, Pr. Sci. Nat. as documented in an NI 43-101 Technical report by P&E Mining Consultants Inc (“P&E”) and Mr. F.H. Brown CPG, Pr. Sci. Nat dated January 31, 2007. This document has been incorporated into Sections 1 through 21 of the current report.
 
Aker Kvaerner was retained by GFAP In April 2006 to initially review the testwork that has been undertaken on the Suhanko Project over the past several years in order to come up with recommendations regarding possible development of the project into an operating mine. A report was presented on May 18, 2006, presenting the findings of the initial investigation.
 
Subsequently, Aker Kvaerner was retained by NAP/GFAP to carry out a trade-off study to compare the process alternatives that were identified during the initial investigation with the original concentrator flow sheet used in the 2005 Lycopodium Feasibility Study for a capacity of 5Mt/a. The study also included a comparison of using the Platsol Process to treat a bulk concentrate compared to selling the concentrate to a smelter. Preliminary designs and capital and operating cost
 
North American Palladium Ltd - Suhanko PGE Project
Technical, Resource Estimate, and Preliminary Economic Assessment Report No. 135
 
1

 
estimates were developed for each of these options and a preliminary financial evaluation was developed to determine the optimum case for further evaluation at the PEA (Scoping Study) stage. The optimum case, determined from the study, was to use the Aker Kvaerner flow sheet to produce a single bulk concentrate for sale.
 
Based on the results of the Trade-off Study, GFAP asked Aker Kvaerner to carry-out a PEA for the Suhanko Project based on producing a bulk concentrate for sale. As part of this, a Capacity Trade-off Study was carried out to evaluate 5 and 7.5 Mt/a operations to pick the optimum capacity for executing the PEA.
 
The current report presents the results of the PEA that covers mineral resources, mining, mineralogical deportment, metallurgical testwork, processing, infrastructure, capital and operating cost estimates, a preliminary financial evaluation and risk assessment. The mineral resources and mining sections were prepared by P&E and F.H Brown CPG, Pr. Sci. Nat and the mineralogical deportment section was prepared by GFAP.
 
Mr. F. H. Brown, Pr. Sci. Nat., a qualified person under the terms of NI 43-101, conducted a site visit to the Konttijärvi and Ahmavaara Properties on May 04, 2006. An independent verification sampling program on both properties was conducted by Mr. Brown at that time.
 
In addition to the site visit, a study of all relevant parts of the available literature and documented results concerning the project as well as discussions with technical personnel from the company regarding all pertinent aspects of the project, was undertaken. The reader is referred to these data sources, which are outlined in Section 20.0 “References” of this report, for further detail on the project.
 
1.2           SOURCES OF INFORMATION
 
This report is based, in part, on internal company technical reports, and maps, published government reports, company letters and memoranda, and public information as listed in the “References” Section 20.0 of this report. Several sections from reports authored by other consultants have been directly quoted in this report, and are so indicated in the appropriate sections.
 
1.3           UNITS AND CURRENCY
 
Unless otherwise stated all units used in this report are metric. PGE and gold assay values are reported in grams per tonne (“g Pt/t, g Pd/t or g Au/t”). Nickel and copper assay values are reported in percent (“%”) unless some other unit is specifically stated. The USD$ is used throughout this report.
 
1.4           USE OF REPORT
 
This report was jointly prepared in conformance with National Instrument 43-101and Ni 43-101F1 for North American Palladium, Ltd (“NAP”) by Aker Kvaerner E&C, a Division of Aker Kvaerner Canada Inc. (“Aker Kvaerner”), P&E Mining Consultants Inc (“P&E”) and F.H. Brown CPG, Pr. Sci. Nat, pursuant to separate contract agreements (“Agreement(s)”) between NAP and Aker Kvaerner, P&E and F.H. Brown. The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in providing the necessary services for preparation of this report and is based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended to be used by NAP, subject to the terms and
 
2

 
conditions of its contractual obligations with the service providers, which allows it to file this document as a Technical Report with Canadian Securities Regulatory Authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects.
 
The Preliminary Economic Assessment section of this report was prepared for North American Palladium (Client) by Aker Kvaerner E&C, a Division of Aker Kvaerner Canada Inc. (Aker Kvaerner) pursuant to the contract agreement (Agreement) between the Client and Aker Kvaerner.
 
The report is based in whole or in part on information and data provided to Aker Kvaerner by Client and/or third parties. Aker Kvaerner represents that it exercised reasonable care in the preparation of this report and that the report complies with published industry standards for such reports, to the extent such published industry standards exist and are applicable. However, Client agrees that, except to the extent specifically stated in writing in the Agreement, Aker Kvaerner is not responsible for confirming the accuracy of information and data supplied by Client or third parties and that Aker Kvaerner does not attest to or assume responsibility for the accuracy of such information or data. Aker Kvaerner also does not attest to or assume responsibility for the accuracy of any recommendations or opinions contained in this report or otherwise expressed by Aker Kvaerner or its employees or agents, which recommendations or opinions are based in whole or in part upon such information or data.
 
The recommendations and opinions contained in this report assume that unknown, unforeseeable, or unavoidable events, which may adversely affect the cost, progress, scheduling or ultimate success of the Project, will not occur.
 
Any discussion of legal issues contained in this report merely reflects technical analysis of Aker Kvaerner and does not constitute legal opinions or the advice of legal counsel.
 
Aker Kvaerner makes no representations, guarantees, or warranties except as expressly stated herein or in the Agreement and all other representations, guarantees, or warranties, whether express or implied, are specifically disclaimed.
 
Except to the limited extent that may be required for this report to qualify as a “technical report” by a “qualified person” in accordance with National Instrument 43-101 as adopted by rulemaking authority of the Ontario Securities Commission and entering into force on December 30, 2005, the use of this report or the information contained herein is at the users sole risk. Client specifically agrees to release, defend, indemnify and hold Aker Kvaerner, its affiliated companies, and its/their officers, directors, employees and agents harmless from any and all liability, damages, or losses of any type, including consequential and punitive damages, suffered by Client or any third party, even if such damages or losses are contributed to or caused by the sole or concurrent fault or negligence of Aker Kvaerner, its affiliated companies, or its/their officers, directors, employees, or agents; provided, however, such release, limitation and indemnity provisions shall be effective to, and only to, the maximum extent allowable by law.

1.5           GLOSSARY OF TERMS

%	 	percent	
µm	 	micrometer, micron	
APP	 	Arctic Platinum Project	
Au-Ag alloy	 	gold-silver alloy	
 
3

 

BMA	 	Bulk Modal Analysis	
BMGS	 	BM Geological Services Pty Ltd	
CIM	 	Canadian Institute of Mining, Metallurgy, and Petroleum	
Cp	 	chalcopyrite	
Cr	 	chromium	
Cu-Ni	 	copper-nickel	
d/a	 	days per annum	
EDS	 	Energy Dispersive Spectrometry	
EPCM	 	Engineering, Procurement, and Construction Management	
Ga	 	gallium	
GFAP	 	Gold Fields Arctic Platinum OY	
GTK	 	Geological Survey of Finland	
HPGR	 	high pressure grinding rolls	
ICP	 	inductively coupled plasma	
IP	 	induced polarization	
Ir	 	iridium	
IRR	 	Internal Rate of Return	
JORC Code	 	means the Australasian Code for Reporting of Mineral	
 	 	Resources and Ore Reserves prepared by the Joint Ore	
 	 	Reserves Committee of the Australasian Institute of Mining	
 	 	and Metallurgy, Australian Institute of Geoscientists and	
 	 	Mineral Council of Australia as amended or supplemented	
M	 	million	
m3	 	cubic metre(s)	
Mg	 	magnesium	
MgO	 	magnesium oxygen	
mm	 	millimetre(s)	
MS	 	mass spectrometer	
Mt	 	million tonnes	
Mt/a	 	million tonnes per annum	
n	 	nano	
NAP	 	North American Palladium	
Ni	 	nickel	
NI	 	National Institute	
Ni-Cu	 	nickel-copper	
NSR	 	net smelter return	
OES	 	optical emission spectrometer	
Os	 	Osmium	
P&E	 	P&E Mining Consultants Inc.	
PAG	 	potentially acid generating	
Pd	 	palladium	
Pd-arsenides	 	palladium-arsenides	
Pd-sulphides	 	palladium-sulphides	
Pd-tellurides	 	palladium-tellurides	
PGE(s)	 	platinum group element(s) including Pt, Pd, Os, Rh and Ru	
PGM(s)	 	platinum group metal(s) synonymous with PGE’s	
Pn	 	pentlandite	
Po	 	pyrrhotite	
ppb	 	parts per billion	
ppm	 	parts per million	
Pt	 	Platinum	
Pt-arsenides	 	platinum-arsenides	
Py	 	pyrite	
Rh	 	rhodium	
ROM	 	run of mine	
Ru	 	Ruthenium	
 
4

 

S.D.	 	standard deviation	
SAG	 	semi-autogenous grinding	
SEM	 	scanning electron microscope	
SMU	 	selective mining unit	
t/d	 	tonnes per day	
TIN	 	Triangulated Irregular Network	
USD	 	US Dollars	
VAT	 	Value added tax	
Wt%	 	weight percentage	
 
5

 
2.0           RELIANCE ON OTHER EXPERTS
 
The authors of this report have relied on the fact that all the information and existing technical documents listed in Section 20 “References” of this report are accurate and complete in all material aspects. While all of the available information presented to the authors was examined and is believed to be reliable as determined to the best of their professional abilities, the accuracy and completeness of such information can not be guaranteed and the authors have relied on the belief that the previous documents have been subject to peer review and prepared in a professional and ethical manner. The authors reserve the right, but will not be obligated, to revise the report and conclusions if additional information becomes known subsequent to the date of this report.
 
Although copies of the underlying tenure documents, operating licenses, permits, and work contracts were reviewed, an independent verification of land title and tenure as summarized in Section 3 of this report was not performed. The authors have not verified the legality of any underlying agreement(s) that may exist concerning the licenses or other agreement(s) between third parties but have relied on the client’s solicitor(s) to have conducted the proper legal due diligence.
 
Recent and on-going environmental assessment and permitting work on the Suhanko Project was conducted by NAP/GFAP and the authors have relied on their legal and political expertise in such matters.
 
A draft copy of the report has been reviewed for factual errors by the client and the authors have relied on NAP’s historical and current knowledge of the property in this regard. The statement and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are not false and misleading at the date of this report.
 
6

 
3.0           PROPERTY DESCRIPTION AND LOCATION
 
3.1           LOCATION
 
The Suhanko Project is located approximately 60 km south of the town of Rovaniemi and 30 km NW of Ranua on the Arctic Circle in northern Finland. The town of Rovaniemi has a population of approximately 34,400 and as the capital of Lapland, is a major regional centre. The town is serviced by rail, road and air with multiple flights daily to and from Helsinki. The port of Kemi on the Gulf of Bothnia is kept open throughout winter and is located 120 km south of Rovaniemi as shown in the Figure 3-1.
 
3.2           TENURE
 
In order to acquire the mineral rights of the highly prospective margin of the Portimo Intrusive Complex (composed of the Suhanko, Konttijärvi, and Narkaus intrusive bodies), GFAP has acquired a total of 258 contiguous claims with a total area of 22 260.2 ha as detailed in Table 3.1. Fifteen of these claims (1388.1 ha) are subject to a purchase agreement with South Atlantic Resources Ltd. GFAP also has nine active claim applications, covering an area of 740.4 ha, that are undergoing the renewal process. With its large landholding GFAP controls virtually the entire extent of the Marginal Phases of the Portimo Complex
 
The Suhanko project is one of several active GFAP projects within the Portimo Intrusive Complex and covered by the regional claim holdings as shown in Figure 3-2. Being the most active of the projects, the Suhanko Project area  is protected not only by the claims but also by a mining license application covering 4145.4 hectares (this lease application has gone through the approval process and only the concession certificate is pending).
 
All properties are currently in good standing or are undergoing the lease renewal process. The tenements covered by mining licenses have been legally surveyed using the UTM Finnish KKJ Zone 3 grid while individual mining claims outside of the mining leases, having been acquired through a government regulated staking process, have not been surveyed.
 
The Finnish government has no free carried interest in the partnership and no royalties are payable to the State. Regular corporate income tax will be payable at a rate which is currently 29% but which is scheduled to be reduced incrementally. For the claim rights GFAP must pay annual claim compensations to the private landowners and annual surface area based fees to the State. The total annual amount for current claim coverage is approximately 394,654.17 €.
 
7

 
TABLE 3.1:  GOLD FIELDS ARCTIC PLATINUM OY –  LIST OF TENEMENTS AS AT 23 OCTOBER 2006
 
1. MINING LICENCES AND MINING LICENCE APPLICATIONS

Licence	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
No.	 	Licence name	 	Project	 	Date applied	 	Valid until	 	Status	 	Area (ha)	 	Remarks	 	
4691/1a	 	Konttijärvi	 	Suhanko	 	9/24/1996	 	10/8/2006	 	Valid	 	56.40	 	Extension applied	 	
3556/1a	 	Kilvenjärvi	 	Narkaus	 	11/28/1991	 	11/28/2006	 	Valid/ext	 	20.21	 	Extension granted 15.01.2002	 	
4689/1a	 	Sompuoja	 	Penikat	 	10/8/1996	 	10/8/2006	 	Valid	 	34.22	 	Extension applied	 	
 	 	Suhanko	 	Suhanko	 	4/23/2001	 	 	 	Application	 	4145.40	 	Pending concession certificate	 	
Totals:	 	3 valid mining licences	 	 	 	 	 	 	110.83	 	ha	 	
 	 	1 mining licence application	 	 	 	 	 	 	4145.40	 	ha	 	
 
2. CLAIMS

Claim No.	 	Claim name	 	Project	 	Date granted	 	Expiry Date	 	Status	 	Area (ha)	 	Remarks	 	
7044/1	 	Kämä 1	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.10	 	Extension granted 09.08.2005	 	
7044/2	 	Kämä 2	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	74.30	 	Extension granted 09.08.2005	 	
7044/3	 	Kämä 3	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	37.30	 	Extension granted 09.08.2005	 	
7044/4	 	Kämä 4	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	66.20	 	Extension granted 09.08.2005	 	
7044/5	 	Kämä 5	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	90.84	 	Extension granted 09.08.2005	 	
7044/6	 	Kämä 6	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	97.01	 	Extension granted 09.08.2005	 	
7045/1	 	Narkaus 1	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.23	 	Extension granted 09.08.2005	 	
7045/2	 	Narkaus 2	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.30	 	Extension granted 09.08.2005	 	
7045/3	 	Narkaus 3	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	97.90	 	Extension granted 09.08.2005	 	
7045/4	 	Narkaus 4	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	93.80	 	Extension granted 09.08.2005	 	
7045/5	 	Narkaus 5	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	96.50	 	Extension granted 09.08.2005	 	
7045/6	 	Narkaus 6	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	95.10	 	Extension granted 09.08.2005	 	
7045/7	 	Narkaus 7	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	100.00	 	Extension granted 09.08.2005	 	
7045/8	 	Narkaus 8	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.90	 	Extension granted 09.08.2005	 	
7045/9	 	Narkaus 9	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.90	 	Extension granted 09.08.2005	 	
7045/10	 	Narkaus 10	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.70	 	Extension granted 09.08.2005	 	
7045/11	 	Narkaus 11	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	98.70	 	Extension granted 09.08.2005	 	
7045/12	 	Narkaus 12	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	85.00	 	Extension granted 09.08.2005	 	
7045/13	 	Narkaus 13	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.60	 	Extension granted 09.08.2005	 	
7045/14	 	Narkaus 14	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	99.40	 	Extension granted 09.08.2005	 	
7045/15	 	Narkaus 15	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	98.30	 	Extension granted 09.08.2005	 	
7045/16	 	Narkaus 16	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	94.70	 	Extension granted 09.08.2005	 	
7045/17	 	Narkaus 17	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	95.40	 	Extension granted 09.08.2005	 	
7045/18	 	Narkaus 18	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	98.40	 	Extension granted 09.08.2005	 	
7045/19	 	Narkaus 19	 	Narkaus	 	7/4/2000	 	7/4/2008	 	Valid/ext	 	43.20	 	Extension granted 09.08.2005	 	
7060/1	 	Kuohunki 1	 	Narkaus	 	8/22/2000	 	8/22/2008	 	Valid/ext	 	77.10	 	Extension granted 09.08.2005	 	
7060/2	 	Kuohunki 2	 	Narkaus	 	8/22/2000	 	8/22/2008	 	Valid/ext	 	84.00	 	Extension granted 09.08.2005	 	
7060/3	 	Kuohunki 3	 	Narkaus	 	8/22/2000	 	8/22/2008	 	Valid/ext	 	95.40	 	Extension granted 09.08.2005	 	
7060/4	 	Kuohunki 4	 	Narkaus	 	8/22/2000	 	8/22/2008	 	Valid/ext	 	57.90	 	Extension granted 09.08.2005	 	
7104/1	 	Saunakivalo 1	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/2	 	Saunakivalo 2	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/3	 	Saunakivalo 3	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/4	 	Saunakivalo 4	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.80	 	Extension granted 11.11.2005	 	
7104/5	 	Saunakivalo 5	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/6	 	Saunakivalo 6	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.40	 	Extension granted 11.11.2005	 	
7104/7	 	Saunakivalo 7	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/8	 	Saunakivalo 8	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/10	 	Saunakivalo 10	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7104/11	 	Saunakivalo 11	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.40	 	Extension granted 11.11.2005	 	
7104/12	 	Saunakivalo 12	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.90	 	Extension granted 11.11.2005	 	
7104/13	 	Saunakivalo 13	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	92.70	 	Extension granted 11.11.2005	 	
7104/14	 	Saunakivalo 14	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	85.40	 	Extension granted 11.11.2005	 	
7104/15	 	Saunakivalo 15	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.50	 	Extension granted 11.11.2005	 	
7104/16	 	Saunakivalo 16	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	31.00	 	Extension granted 11.11.2005	 	
7105/1	 	Peräjärvi 1	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	43.40	 	Extension granted 11.11.2005	 	
 
8

 

Claim No.	 	Claim name	 	Project	 	Date granted	 	Expiry Date	 	Status	 	Area (ha)	 	Remarks	 	
7105/2	 	Peräjärvi 2	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	92.60	 	Extension granted 11.11.2005	 	
7105/3	 	Peräjärvi 3	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	92.60	 	Extension granted 11.11.2005	 	
7105/4	 	Peräjärvi 4	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7105/5	 	Peräjärvi 5	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	99.30	 	Extension granted 11.11.2005	 	
7105/6	 	Peräjärvi 6	 	Narkaus	 	11/3/2000	 	11/3/2008	 	Valid/ext	 	91.20	 	Extension granted 11.11.2005	 	
7106/1	 	Kämä 7	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	76.70	 	Extension granted 11.11.2005	 	
7106/2	 	Kämä 8	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	54.80	 	Extension granted 11.11.2005	 	
7106/3	 	Kämä 9	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	93.80	 	Extension granted 11.11.2005	 	
7106/4	 	Kämä 10	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7106/5	 	Kämä 11	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	62.00	 	Extension granted 11.11.2005	 	
7106/6	 	Kämä 12	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	98.10	 	Extension granted 11.11.2005	 	
7106/7	 	Kämä 13	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	94.80	 	Extension granted 11.11.2005	 	
7106/8	 	Kämä 14	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7106/9	 	Kämä 15	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	95.20	 	Extension granted 11.11.2005	 	
7106/10	 	Kämä 16	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	99.10	 	Extension granted 11.11.2005	 	
7106/11	 	Kämä 17	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	87.00	 	Extension granted 11.11.2005	 	
7106/12	 	Kämä 18	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	55.90	 	Extension granted 11.11.2005	 	
7106/13	 	Kämä 19	 	Narkaus	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	95.10	 	Extension granted 11.11.2005	 	
7119/1	 	Portimo 80	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	80.20	 	Extension applied	 	
7119/2	 	Portimo 81	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	97.20	 	Extension applied	 	
7119/3	 	Portimo 82	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.40	 	Extension applied	 	
7119/4	 	Portimo 89	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.80	 	Extension applied	 	
7119/5	 	Portimo 90	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	98.80	 	Extension applied	 	
7119/6	 	Portimo 101	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.00	 	Extension applied	 	
7119/7	 	Portimo 102	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	98.80	 	Extension applied	 	
7119/8	 	Portimo 105	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/9	 	Portimo 106	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	98.70	 	Extension applied	 	
7119/10	 	Portimo 107	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	58.30	 	Extension applied	 	
7119/11	 	Portimo 108	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.40	 	Extension applied	 	
7119/12	 	Portimo 115	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/13	 	Portimo 116	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.00	 	Extension applied	 	
7119/14	 	Portimo 117	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.10	 	Extension applied	 	
7119/15	 	Portimo 121	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.40	 	Extension applied	 	
7119/16	 	Portimo 122	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.30	 	Extension applied	 	
7119/17	 	Portimo 123	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.20	 	Extension applied	 	
7119/18	 	Portimo 124	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/19	 	Portimo 128	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/20	 	Portimo 129	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	99.70	 	Extension applied	 	
7119/21	 	Portimo 130	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/23	 	Portimo 135	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	96.60	 	Extension applied	 	
7119/24	 	Portimo 141	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	34.60	 	Extension applied	 	
7119/25	 	Portimo 145	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	98.70	 	Extension applied	 	
7119/26	 	Portimo 146	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/27	 	Portimo 148	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	100.00	 	Extension applied	 	
7119/29	 	Portimo 162	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	46.60	 	Extension applied	 	
7119/30	 	Portimo 163	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	9.00	 	Extension applied	 	
7119/31	 	Kuohunki - 1	 	Narkaus	 	10/3/2001	 	10/3/2006	 	Valid	 	35.60	 	Extension applied	 	
7409/1	 	Portimo 164	 	Narkaus	 	5/7/2002	 	5/7/2007	 	Valid	 	100.00	 	 	 	
7409/2	 	Portimo 165	 	Narkaus	 	5/7/2002	 	5/7/2007	 	Valid	 	99.90	 	 	 	
7518/1	 	Saariaapa 1	 	Narkaus	 	1/20/2003	 	12/31/2007	 	Valid	 	96.70	 	 	 	
7518/2	 	Saariaapa 2	 	Narkaus	 	1/20/2003	 	12/31/2007	 	Valid	 	98.20	 	 	 	
7518/3	 	Saariaapa 3	 	Narkaus	 	1/20/2003	 	12/31/2007	 	Valid	 	97.80	 	 	 	
7545/1	 	Ristiaapa 1	 	Narkaus	 	1/20/2003	 	12/31/2007	 	Valid	 	99.00	 	 	 	
7545/2	 	Ristiaapa 2	 	Narkaus	 	1/20/2003	 	12/31/2007	 	Valid	 	98.50	 	 	 	
6998/1	 	Kilkka 1	 	Penikat	 	3/27/2000	 	3/27/2008	 	Valid/ext	 	94.90	 	Extension granted 30.03.2005	 	
6998/2	 	Kilkka 2	 	Penikat	 	3/27/2000	 	3/27/2008	 	Valid/ext	 	92.20	 	Extension granted 30.03.2005	 	
7019/1	 	Penikat 1	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	92.50	 	Extension granted 30.03.2005	 	
7019/2	 	Penikat 2	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	98.80	 	Extension granted 30.03.2005	 	
 
9

 

Claim No.	 	Claim name	 	Project	 	Date granted	 	Expiry Date	 	Status	 	Area (ha)	 	Remarks	 	
7019/3	 	Penikat 3	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	97.20	 	Extension granted 30.03.2005	 	
7019/4	 	Penikat 4	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	72.80	 	Extension granted 30.03.2005	 	
7019/5	 	Penikat 5	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	64.00	 	Extension granted 30.03.2005	 	
7019/6	 	Penikat 6	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	72.60	 	Extension granted 30.03.2005	 	
7019/7	 	Penikat 7	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	91.60	 	Extension granted 30.03.2005	 	
7019/8	 	Penikat 8	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	57.10	 	Extension granted 30.03.2005	 	
7019/9	 	Penikat 9	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	56.60	 	Extension granted 30.03.2005	 	
7019/10	 	Penikat 10	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	89.10	 	Extension granted 30.03.2005	 	
7019/11	 	Penikat 11	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	94.20	 	Extension granted 30.03.2005	 	
7019/12	 	Penikat 12	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	52.00	 	Extension granted 30.03.2005	 	
7019/13	 	Penikat 13	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	18.30	 	Extension granted 30.03.2005	 	
7019/14	 	Penikat 14	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	78.70	 	Extension granted 30.03.2005	 	
7019/15	 	Penikat 15	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	29.10	 	Extension granted 30.03.2005	 	
7019/16	 	Penikat 16	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	98.60	 	Extension granted 30.03.2005	 	
7019/17	 	Penikat 17	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/18	 	Penikat 18	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	92.80	 	Extension granted 30.03.2005	 	
7019/19	 	Penikat 19	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	99.70	 	Extension granted 30.03.2005	 	
7019/20	 	Penikat 20	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	82.20	 	Extension granted 30.03.2005	 	
7019/21	 	Penikat 21	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	93.10	 	Extension granted 30.03.2005	 	
7019/22	 	Penikat 22	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	58.00	 	Extension granted 30.03.2005	 	
7019/23	 	Penikat 23	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/24	 	Penikat 24	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	93.20	 	Extension granted 30.03.2005	 	
7019/25	 	Penikat 25	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	96.40	 	Extension granted 30.03.2005	 	
7019/26	 	Penikat 26	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	97.90	 	Extension granted 30.03.2005	 	
7019/27	 	Penikat 27	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	93.40	 	Extension granted 30.03.2005	 	
7019/28	 	Penikat 28	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	98.00	 	Extension granted 30.03.2005	 	
7019/29	 	Penikat 29	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/30	 	Penikat 30	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	93.50	 	Extension granted 30.03.2005	 	
7019/31	 	Penikat 31	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/32	 	Penikat 32	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/33	 	Penikat 33	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	91.70	 	Extension granted 30.03.2005	 	
7019/34	 	Penikat 34	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/35	 	Penikat 35	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	94.20	 	Extension granted 30.03.2005	 	
7019/36	 	Penikat 36	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/37	 	Penikat 37	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/38	 	Penikat 38	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	97.30	 	Extension granted 30.03.2005	 	
7019/39	 	Penikat 39	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/40	 	Penikat 40	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	91.50	 	Extension granted 30.03.2005	 	
7019/41	 	Penikat 41	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	97.50	 	Extension granted 30.03.2005	 	
7019/42	 	Penikat 42	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	92.10	 	Extension granted 30.03.2005	 	
7019/43	 	Penikat 43	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	96.40	 	Extension granted 30.03.2005	 	
7019/44	 	Penikat 44	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	80.60	 	Extension granted 30.03.2005	 	
7019/45	 	Penikat 45	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	85.80	 	Extension granted 30.03.2005	 	
7019/46	 	Penikat 46	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	100.00	 	Extension granted 30.03.2005	 	
7019/47	 	Penikat 47	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	95.00	 	Extension granted 30.03.2005	 	
7019/48	 	Penikat 48	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	96.50	 	Extension granted 30.03.2005	 	
7019/49	 	Penikat 49	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	99.10	 	Extension granted 30.03.2005	 	
7019/50	 	Penikat 50	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	92.90	 	Extension granted 30.03.2005	 	
7019/51	 	Penikat 51	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	93.30	 	Extension granted 30.03.2005	 	
7019/52	 	Penikat 52	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	99.60	 	Extension granted 30.03.2005	 	
7019/53	 	Penikat 53	 	Penikat	 	5/11/2000	 	5/11/2008	 	Valid/ext	 	77.50	 	Extension granted 30.03.2005	 	
7208/1	 	Penikat 57	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	98.00	 	Extension applied	 	
7208/2	 	Penikat 65	 	Penikat	 	1/30/2002	 	12/31/2006	 	Valid	 	91.60	 	 	 	
7208/3	 	Penikat 66	 	Penikat	 	1/30/2002	 	12/31/2006	 	Valid	 	85.90	 	 	 	
7208/4	 	Penikat 67	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	100.00	 	Extension applied	 	
7208/5	 	Penikat 72	 	Penikat	 	1/30/2002	 	12/31/2006	 	Valid	 	83.10	 	 	 	
7208/6	 	Penikat 80	 	Penikat	 	1/30/2002	 	12/31/2006	 	Valid	 	92.90	 	 	 	
7208/7	 	Penikat 81	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	99.80	 	Extension applied	 	
 
10

 

Claim No.	 	Claim name	 	Project	 	Date granted	 	Expiry Date	 	Status	 	Area (ha)	 	Remarks	 	
7208/8	 	Penikat 82	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	99.60	 	Extension applied	 	
7208/9	 	Penikat 93	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	98.10	 	Extension applied	 	
7208/10	 	Penikat 94	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	99.80	 	Extension applied	 	
7208/11	 	Penikat 99	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	95.90	 	Extension applied	 	
7208/12	 	Penikat 100	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	79.70	 	Extension applied	 	
7208/13	 	Penikat 105	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	94.30	 	Extension applied	 	
7208/14	 	Penikat 106	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	97.00	 	Extension applied	 	
7208/15	 	Penikat 110	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	100.00	 	Extension applied	 	
7208/16	 	Penikat 111	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	97.70	 	Extension applied	 	
7208/17	 	Penikat 116	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	70.30	 	Extension applied	 	
7208/18	 	Penikat 117	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	69.70	 	Extension applied	 	
7208/19	 	Penikat 123	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	98.80	 	Extension applied	 	
7208/20	 	Penikat 165	 	Penikat	 	9/21/2001	 	9/21/2006	 	Valid	 	42.60	 	Extension applied	 	
7631/1	 	Takakangas 2	 	Penikat	 	8/20/2003	 	8/20/2008	 	Valid	 	44.40	 	Renewed claim	 	
7631/2	 	Vihvilöjänkä 2	 	Penikat	 	8/20/2003	 	8/20/2008	 	Valid	 	42.70	 	Renewed claim	 	
7103/1	 	Yliportimo 1	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	94.20	 	Extension granted 11.11.2005	 	
7103/2	 	Yliportimo 2	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	87.10	 	Extension granted 11.11.2005	 	
7103/3	 	Yliportimo 3	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	92.00	 	Extension granted 11.11.2005	 	
7103/4	 	Yliportimo 4	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	88.70	 	Extension granted 11.11.2005	 	
7103/5	 	Yliportimo 5	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	95.60	 	Extension granted 11.11.2005	 	
7103/6	 	Yliportimo 6	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	95.80	 	Extension granted 11.11.2005	 	
7103/7	 	Yliportimo 7	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	95.80	 	Extension granted 11.11.2005	 	
7103/8	 	Yliportimo 8	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	96.00	 	Extension granted 11.11.2005	 	
7103/9	 	Yliportimo 9	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	93.40	 	Extension granted 11.11.2005	 	
7103/10	 	Yliportimo 10	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7103/11	 	Yliportimo 11	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	98.20	 	Extension granted 11.11.2005	 	
7103/14	 	Yliportimo 14	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	93.30	 	Extension granted 11.11.2005	 	
7103/15	 	Yliportimo 15	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	94.20	 	Extension granted 11.11.2005	 	
7103/18	 	Yliportimo 18	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7103/19	 	Yliportimo 19	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7103/20	 	Yliportimo 20	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	91.80	 	Extension granted 11.11.2005	 	
7103/22	 	Yliportimo 22	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	92.80	 	Extension granted 11.11.2005	 	
7103/23	 	Yliportimo 23	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	100.00	 	Extension granted 11.11.2005	 	
7103/24	 	Yliportimo 24	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	69.60	 	Extension granted 11.11.2005	 	
7103/25	 	Yliportimo 25	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	82.00	 	Extension granted 11.11.2005	 	
7103/26	 	Yliportimo 26	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	83.00	 	Extension granted 11.11.2005	 	
7103/27	 	Yliportimo 27	 	Suhanko	 	11/15/2000	 	11/15/2008	 	Valid/ext	 	40.50	 	Extension granted 11.11.2005	 	
7206/1	 	Portimo - 1	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	30.10	 	Extension applied	 	
7206/4	 	Portimo - 4	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	9.30	 	Extension applied	 	
7206/17	 	Portimo - 18	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	60.30	 	Extension applied	 	
7206/18	 	Portimo - 19	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	36.40	 	Extension applied	 	
7206/19	 	Portimo - 20	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	36.50	 	Extension applied	 	
7206/20	 	Portimo - 21	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	5.60	 	Extension applied	 	
7206/23	 	Portimo - 24	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	31.90	 	Extension applied	 	
7206/24	 	Portimo - 25	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	31.60	 	Extension applied	 	
7206/25	 	Portimo - 26	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	38.70	 	Extension applied	 	
7206/26	 	Portimo - 27	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	68.10	 	Extension applied	 	
7206/27	 	Portimo - 28	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	96.50	 	Extension applied	 	
7206/28	 	Portimo - 29	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	99.40	 	Extension applied	 	
7206/29	 	Portimo - 30	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	99.70	 	Extension applied	 	
7206/30	 	Portimo - 31	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	94.70	 	Extension applied	 	
7206/31	 	Portimo - 32	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	36.20	 	Extension applied	 	
7206/32	 	Portimo - 33	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	100.00	 	Extension applied	 	
7206/33	 	Portimo - 38	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	93.60	 	Extension applied	 	
7206/34	 	Portimo - 40	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	48.50	 	Extension applied	 	
7206/35	 	Portimo - 41	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	94.60	 	Extension applied	 	
7206/36	 	Portimo - 42	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	93.30	 	Extension applied	 	
7206/37	 	Portimo - 43	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	92.90	 	Extension applied	 	
 
11

 

Claim No.	 	Claim name	 	Project	 	Date granted	 	Expiry Date	 	Status	 	Area (ha)	 	Remarks	 	
7206/38	 	Portimo - 44	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	61.40	 	Extension applied	 	
7206/39	 	Portimo - 45	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	81.80	 	Extension applied	 	
7206/40	 	Portimo - 47	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	96.20	 	Extension applied	 	
7206/41	 	Portimo - 48	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	91.60	 	Extension applied	 	
7206/42	 	Portimo - 49	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	96.20	 	Extension applied	 	
7206/43	 	Portimo - 50	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	89.90	 	Extension applied	 	
7206/44	 	Portimo - 51	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	95.70	 	Extension applied	 	
7206/45	 	Portimo - 53	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	95.10	 	Extension applied	 	
7206/46	 	Portimo - 54	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	98.20	 	Extension applied	 	
7206/47	 	Portimo - 55	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	91.70	 	Extension applied	 	
7206/48	 	Portimo - 56	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	95.70	 	Extension applied	 	
7206/49	 	Portimo - 57	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	98.10	 	Extension applied	 	
7206/50	 	Portimo - 58	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	91.40	 	Extension applied	 	
7206/51	 	Portimo - 59	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	91.10	 	Extension applied	 	
7206/52	 	Portimo - 60	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	100.00	 	Extension applied	 	
7206/53	 	Portimo - 61	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	90.50	 	Extension applied	 	
7206/54	 	Portimo - 62	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	64.70	 	Extension applied	 	
7206/55	 	Portimo - 63	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	37.70	 	Extension applied	 	
7206/56	 	Portimo - 65	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	97.70	 	Extension applied	 	
7206/57	 	Portimo - 67	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	96.90	 	Extension applied	 	
7206/58	 	Portimo - 69	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	100.00	 	Extension applied	 	
7206/59	 	Portimo - 70	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	95.60	 	Extension applied	 	
7206/60	 	Portimo - 160	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	38.20	 	Extension applied	 	
7206/61	 	Portimo - 161	 	Suhanko	 	10/22/2001	 	10/22/2006	 	Valid	 	35.00	 	Extension applied	 	
7197/1	 	KPP 38	 	Suhanko N	 	8/28/2001	 	8/28/2006	 	Valid	 	90.5	 	JV claim / extension applied	 	
7197/2	 	KPP 39	 	Suhanko N	 	8/28/2001	 	8/28/2006	 	Valid	 	97.8	 	JV claim / extension applied	 	
7197/3	 	KPP 40	 	Suhanko N	 	8/28/2001	 	8/28/2006	 	Valid	 	98.3	 	JV claim / extension applied	 	
7198/3	 	KPP 60	 	Suhanko N	 	8/28/2001	 	8/28/2006	 	Valid	 	72.2	 	JV claim / extension applied	 	
7198/4	 	KPP 61	 	Suhanko N	 	8/28/2001	 	8/28/2006	 	Valid	 	79.5	 	JV claim / extension applied	 	
7200/1	 	KPP 76	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	97.6	 	JV claim / extension applied	 	
7200/2	 	KPP 77	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.0	 	JV claim / extension applied	 	
7200/3	 	KPP 78	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.2	 	JV claim / extension applied	 	
7200/4	 	KPP 79	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.0	 	JV claim / extension applied	 	
7200/5	 	KPP 80	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	64.3	 	JV claim / extension applied	 	
7200/6	 	KPP 81	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	98.0	 	JV claim / extension applied	 	
7200/7	 	KPP 82	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.8	 	JV claim / extension applied	 	
7200/8	 	KPP 83	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.7	 	JV claim / extension applied	 	
7200/9	 	KPP 84	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	99.6	 	JV claim / extension applied	 	
7200/10	 	KPP 85	 	Narkaus E	 	8/28/2001	 	8/28/2006	 	Valid	 	93.6	 	JV claim / extension applied	 	
Total	 	258	 	valid claims	 	 	 	 	 	22,260.2	 	ha	 	
Including	 	15	 	GFAP - South Atlantic Resources JV claims (KPP-)	 	1,388.1	 	ha	 	
 
3. CLAIM APPLICATIONS

Claim
name	 	Claim No.	 	Project	 	Date granted	 	expiry date	 	Status	 	Area (ha)	 	Remarks	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	99.70	 	 	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	98.8	 	 	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	80.4	 	 	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	81.8	 	 	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	82.2	 	JV claim	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	70.9	 	JV claim	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	99.7	 	JV claim	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	33.1	 	JV claim	 	
Suhanko	 	 	 	Suhanko	 	 	 	 	 	Application	 	93.8	 	JV claim	 	
Total	 	 	 	claim applications	 	 	 	 	 	740.4	 	ha	 	
Including	 	 	 	GFAP - SAR JV claims (Suhanko 5-9)	 	 	 	379.7	 	ha	 	
 
4. CLAIM RESERVATIONS
   None
 
12

 

Figure 3.1 Location of Suhanko Project
 
3.2.1        PREVIOUS AND CURRENT TENEMENT OWNERSHIP
 
Gold Fields and Outokumpu Mining Oy (“Outokumpu”) formed a joint venture company, (the “Partnership”), by partnership agreement on 17 March 2000. The Partnership’s objective was to explore for and produce PGEs in northern Finland. Outokumpu’s initial share of the Partnership was 100% subject to a pending business arrangement, which came into force on 1 January 2000 with the following items:
 
•                   Outokumpu earned its initial interest of 100% by contributing 19 Claim Reservations (144 003 ha), 3 Claims (185.6 ha) and 3 Mining Licenses (110.8 ha), together with all available geological data. The Partnership agreement covered an Area of Interest that was originally explored by Outokumpu from the 1960s and consequently resulted in the development of an extensive geological database (that was made available to The Partnership).
 
•                   Gold Fields, as operator, could earn a 30% share by spending USD $5,000,000 on operating costs by 31 December 2002. Gold Fields could withdraw at any time but would retain no interest in the project unless this criterion was met.
 
•                   After earning its 30% interest, which diluted Outokumpu to 70%, Gold Fields was able to earn a 49% share by spending an aggregate amount of USD $11,000,000 by 31 December 2004.
 
13

 
•                   Thereafter, Gold Fields would have the option to earn a 51% majority interest by making an additional contribution of USD $2,000,000 within 12 months.
 
Exploration was more rapid than expected and Gold Fields earned its 49% share in November 2001 and increased it to 51% in January 2002.
 
Gold Fields acquired 100% of this project on September 10, 2003, by purchasing Outokumpu’s 49% share of the Arctic Platinum Partnership for USD $31.0 M.
 
In July 2004, the company was transferred from partnership status to a limited liability company – GFAP – which was a 100% owned subsidiary of Gold Fields Limited. The search for platinum group elements within the Rovaniemi District became known as the Artic Platinum Project or “APP”.
 
Under the terms of an agreement dated March 24, 2006, North American Palladium and Gold Fields Limited agreed to form an option joint venture to further explore and possibly develop a mining operation under the auspices of the APP. Included under this arrangement were several advanced stage PGE projects such as the Konttijärvi and Ahmavaara Open Pit Projects, collectively known as the Suhanko Project.
 
3.3           ENVIRONMENTAL PERMITING
 
3.3.1        SUHANKO MINING LEASE AND ENVIRONMENTAL PERMIT CURRENT STATUS
 
The final meeting regarding the execution of the Suhanko Mining Lease was held on 29 th May 2006. The minutes of the meeting were published on 24 th July and became legally valid in early September. The executing engineer has forwarded the minutes and related documents of the final meeting to the Ministry of Trade and Industry in September. Granting of a concession certificate is pending in the Ministry.
 
GFAP has purchased the surface rights of an area of 3,701.2 hectares out of the total Mining Lease surface area of 4,143.805 hectares. Arbitration regarding these lands has been completed with the landowners.
 
The Northern Finland Environmental Permit Authority granted the Environmental Permit for the Suhanko Project on 7 th December 2005. Two appeals were made against the Permit concerning the amounts of compensation for land for two small properties. Gold Fields Arctic Platinum Oy submitted its statement of defence in February 2006. The decision from the Administrative court is currently being awaited The Suhanko Mining Lease and Area subject to the Environmental Permit is shown on Figure 3-3.
 
The Energy Market Authority issued a construction permit for the Petäjäskoski – Konttijärvi 110 kV overhead transmission line on 12th August 2004. Yet to be applied for are the acceptance for general plan for mining (KTM 1975/921) from the Safety Technology Authority, the construction licence for buildings and workshops, the environmental permit for fuel storage as well as a number of other minor permits.
 
14

 

Figure 3-2: Suhanko Project Tenement Map	 	P&E Mining Consultants Inc.	

 
The rehabilitation bond to be deposited upon initiation of the construction of the tailings basin and waste rock stockpiling areas is 6,620,000 €.
 
3.4          ENVIRONMENTAL OBLIGATIONS
 
There are no current environmental obligations of significance outstanding. At the Ahmavaara trial mine the pumping waters and their impact on local surface waters are required to be monitored until the 2006 year end or as long as dewatering is continued. In case the Suhanko Project stops, the pit can be left to flood to ground water level. Rehabilitation of other test pits has been completed.
 
The Suhanko Environmental Permit and the rehabilitation bond have been decided and, despite the two appeals, are effective. As long as the appeals remain unresolved, they cause restrictions for construction of certain water areas and hence influence parts of the TSF and waste rock areas.
 
16

 

Figure 3-3: Suhanko Mineral Lease and Environmental Permit Area (Source GFAP 2006)	 	P&E Mining Consultants Inc.	

 
4.0           ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY
 
4.1           LOCATION, AND ACCESS
 
The Konttijärvi and Ahmavaara properties are examined in the context that they are the most significant discoveries found to-date within the Suhanko Project. The Project area is located approximately 60 km south (via Route 78) of the town of Rovaniemi which is located on the Arctic Circle in northern Finland.
 
Access to both Konttijärvi and Ahmavaara is via a sealed road to within 15 kilometres of the project location, and thence by a well-maintained gravel road passable by light vehicles. All-weather access on the non-sealed roads will be guaranteed by the use of snow ploughs during the autumn and winter months. Figure 4-1 is a detailed location map.
 
Project co-ordinates are relative to a localized grid system that has been rotated from the Finnish national grid. Elevations are relative to Mean Sea Level.
 
4.2           CLIMATE
 
The site is situated just south of the Arctic Circle where the climate is characterized by seasonal changes in temperature and daylight. The climate is, however, moderated by the Gulf Stream. Useful daylight varies from continuous light for a period of around 6 weeks in summer to around 6 hours per day in late December. Average temperatures range from -20°C to +20°C with occasional short-duration cold spells of below -40°C
 
4.3           INFRASTRUCTURE
 
With a population of 35,000, and a regional population of approximately 55,000, the city of Rovaniemi provides all major and modern amenities expected of a capital city. Transportation services include road, rail, and air, with multiple flights daily from Helsinki. The port of Kemi 120 km south (via Route 126) of Rovaniemi, on the Gulf of Bothnia provides modern facilities and year round open water access is maintained through the winter.
 
There are no major towns or villages, on or close to, the layered complexes. However, there are a few scattered homesteads and summer cottages on farms and near lakes located in the project area. The average population density in the area is about 2 per km 2 . Forestry is the main industry and very little primary forest is preserved except in proclaimed reserves.
 
4.4           PHYSIOGRAPHY
 
The topography of the lease area is generally flat as a result of various glaciation events. The layered complexes are covered by overburden that ranges in thickness from 0 - 30 m. Outcrop is rare and the average thickness of till is 5 - 10 m. As a result of the low relief and poor drainage, peat bogs cover large areas such that off road vehicle access in some areas is only possible during the winter months when the surface is frozen.
 
18

 

Figure 4-1: Detailed Access Map (Source Lycopodium 2005)	 	P&E Mining Consultants Inc.	

 
5.0           HISTORY
 
The earliest recorded exploration of the Suhanko project area describes bedrock mapping in the Portimo area by Hackman in 1910 (Hackman, 1914). In the early 1960s, a local farmer noticed mineralization in boulders on the eastern shore of Lake Yli-Portimojarvi, which is located in the Portimo Complex area.
 
Outokumpu Oy commenced a Cu-Ni exploration program in the Suhanko area using magnetic and electro-magnetic ground geophysical survey methods in conjunction with extensive geological mapping and drilling in 1964. The discovery hole, RN/YP 1, drilled on Lake Yli-Portimojarvi, intersected the first Marginal series mineralization in the Suhanko-Konttijärvi intrusion. Exploration continued until 1981, and focused on the disseminated and massive sulphide mineralization in the basal part of the Suhanko Intrusion. Exploration of the Marginal series mineralization was extended to other prospective areas, of the Portimo Intrusive Complex which included Niittylampi and Suhanko. The smaller Konttijärvi intrusion block was discovered following geological mapping and assaying, which led to the first indications of the presence of PGE from sulphide-bearing samples.
 
A brief summary of the historical exploration on both the Konttijärvi and Ahmavaara deposits is given below.
 
5.1           PAST EXPLORATION KONTTIJÄRVI
 
Further work by Outokumpu Oy in 1981 that verified the historical PGE assays of outcrop samples taken in the 1970s determined that true values were three to four times higher than first reported. This motivated a small drilling campaign and the first drillhole, RN/KOJ-1, drilled in September 1981 intersected significant sulphide-hosted PGE mineralization. A further four drill holes were completed all of which intersected significant PGE mineralization. Thus, Konttijärvi became the first significant PGE discovery in Finland.
 
Following these successful results, the area was covered by a ground magnetic survey and also by a small number of geochemical till sampling lines. The magnetic survey detected the location of the Peridotite Marker and the geochemical sampling suggested the existence of a significant occurrence of PGE mineralization within the intrusion block.
 
During the next phase of exploration, Outokumpu Oy completed 14 drillholes with the objective of infilling the mineralized area. The results obtained confirmed the regular and continuous character of the Konttijärvi deposit.
 
The Mining Services department of Outokumpu Oy organised the next exploration phase, which was designed to collect data for an economic evaluation of the Konttijärvi and Kilvenjarvi deposits.
 
The final exploration phase by Outokumpu included the drilling and sampling of drill holes RN/KOJ-160 to 161 (two drill holes) to test the PGE-bearing potential of some outcrops to the south of the pre-stripped area in rocks that stratigraphically form the basement of the Marginal series mineralization and were considered to have a low probability of hosting PGE mineralization. Sulphides and PGE values were discovered within these basement rocks at some distance from the known basal contact of the intrusion.
 
Since 1981, Outokumpu completed 189 diamond drill holes with a total length of 13,617 m at the Konttijärvi and Ahmavaara deposits. This drilling was conducted during three phases, from 1981 to 1983; 1986 to 1989; and finally in 1995.
 
20

 
5.2           PAST EXPLORATION AHMAVAARA
 
The Ahmavaara deposit was discovered by applying the exploration model developed from the Konttijärvi data. Based on this model, and with the results from a ground magnetic survey that identified the location of the all-important peridotite marker, the first seven drill holes, RN/YP-138 to 144 were completed. Laboratory-scale metallurgical test results, supported by a microscopic study of the mineralogy in samples taken from the later drill holes, showed that Ni, Cu and PGE could be concentrated at fairly good recovery rates at Ahmavaara. An additional 12 holes, RN/YP-172 to 183, were drilled at the Ahmavaara area in 1995 to search for nickel-rich occurences.
 
The total amount of drilling at Ahmavaara before the formation of the Arctic Platinum Partnership was 29 holes for 3,322 m.
 
5.3           APP - GFAP ACTIVITY
 
Since its inception on March 17, 2000, GFAP has carried out both exploration and sterilization drilling, and has also used a variety of other exploration techniques to develop and explore the Konttijärvi and Ahmavaara deposits. These include airborne magnetic, electro-magnetic and ground induced polarization (IP) surveys.
 
5.4           HISTORICAL DRILLING
 
Up to June 2004, GFAP completed 1,246 drill holes with a total length of 154,021 m in the Suhanko project area. Of this total, 986 diamond drill holes totalling 108,369 m in length were completed on the Konttijärvi and Ahmavaara deposits and 260 drill holes totalling 45,742 m in length on the SK reef. A recent program of “in-seam” drilling conducted on the Konttijärvi and Ahmavaara deposits since the Lycopodium Feasibility Study in 2005 and prior to the involvement of NAP in February 2006, has added a further 11,594 m of drilling and 67 drill holes, bringing the total amount of diamond drilling on the Suhanko project to 119,963 m and 1,053 drill holes. In seam drilling is a technique whereby drillholes are designed to intersect the target stratigraphy at a shallow angle rather than the traditional approach of intersecting the stratigraphy in a perpendicular orientation. In this manner, a greater volume of sample is provided for analysis by each drillhole.
 
Some 121 drill holes, with a total length of 8,502 m (6.97% of the total Suhanko drilling), were drilled for geotechnical, metallurgical and site investigation requirements.
 
The drilling completed by GFAP is summarized on a year by year and deposit basis in Table 5.1.
 
Table 5.1:Summary of Drilling by GFAP Since Inception to June 2004

 	 	2000	 	2001	 	2002	 	2003	 	2004	 	
Area	 	No.	 	Metres	 	No.	 	Metres	 	No.	 	Metres	 	No.	 	Metres	 	No.	 	Metres	 	
Konttijärvi	 	66	 	6,238	 	118	 	16,491	 	78	 	11,065	 	116	 	5,437	 	92	 	9,607	 	
Ahmavaara	 	17	 	2,190	 	144	 	18,157	 	99	 	14,332	 	44	 	5,699	 	103	 	12,528	 	
Sterilisation drilling	 	0	 	0	 	13	 	477	 	85	 	5807	 	0	 	0	 	11	 	251	 	
Sub-Total (K & A)	 	83	 	8,428	 	275	 	35,125	 	262	 	31,204	 	160	 	11,136	 	206	 	22,386	 	
SK Reef	 	11	 	1,147	 	87	 	16,092	 	88	 	16,638	 	74	 	11,865	 	0	 	0	 	
Suhanko Project Total	 	94	 	9,575	 	362	 	51,217	 	350	 	47,842	 	234	 	23,001	 	206	 	22,386	 	
 
21

 
All drilling by Outokumpu, and GFAP, excluding some blastholes in the test pits, has been by diamond core. Core diameter is typically 57.5 mm. Of 260 drill holes drilled at the SK deposits there have been 23 deflection holes.
 
5.4.1        DRILL CORE MANAGEMENT
 
GFAP has a well-defined core processing system for the Suhanko and Narkaus projects as summarized in Figure 5-1, which illustrates in flow chart form the logging and cutting procedures undertaken. The core storage and logging facility is indoors, allowing safe and efficient logging and processing year-round.
 
Core is processed in the following order:
 
•       meter marking of drill cores;
•       magnetic susceptibility measuring;
•       geological logging;
•       geotechnical logging;
•       marking of the samples for assaying;
•       core photographing;
•       core cutting;
•       sampling;
•       bulk density measuring; and
•       storage.
 
All activities in the flowchart are carried out by GFAP employees. Snowden (2004c) has studied the core processing procedures and considers them to be commensurate with industry best practice.
 
5.5           PAST PRODUCTION
 
No mining activities have been carried out at either the Konttijärvi or Ahmavaara deposits other than the excavation of a small test pit at each prospect by GFAP in April 2004. The objective of the test pits was twofold:
 
•       to collect a bulk sample for metallurgical test work at the GTK (or Geological Survey of Finland) owned concentrator in the town of Outokumpu
 
•       to undertake material properties testwork for drillability and blasting of the various rock types present within the Suhanko mine environment.
 
The results from the test pits demonstrated largely that the grades of the elements of interest (palladium, platinum, gold, copper and nickel) matched both the exploratory diamond drilling and the blasthole grade control drilling for the trial pits. Geological details of the test mining are provided in BMGS (2004). There has been no production from any of the SK deposits.
 
22

 

Figure 5-1 Core Processing System
 
5.6           MINERAL RESOURCE ESTIMATES
 
Note: Most of the following mineral Resource Estimates are stated to be classified as Measured, Indicated, and Inferred resources according to the JORC code in place as of September 1999 and later amended in 2004. NI 43-101 sec 7.1 provides for the disclosure and filing of a technical report that “utilizes the mineral resource and mineral reserve categories of the JORC Code…. If a reconciliation to the mineral resource and mineral reserve categories set out in sections 1.2 and 1.3 is disclosed “ Although the existing resource estimates written in conformance with the JORC Code are believed to accurately reflect the contained mineralization on the properties, the work necessary to reconcile and verify the classification of the mineral resources, in compliance with the requirements of NI 43-101, has not been completed by either Aker Kvaerner, F.H. Brown, or P&E and the resource estimates,
 
23

 
therefore, can not be treated as NI 43-101 compliant resources. Consequently, the estimates should not be relied upon and there can be no assurance that any of the resources, in whole or in part, will ever become economically viable.
 
The most recent mineral resource estimates produced by Snowden Mining Industry Consultants, from Perth Australia (Snowden, 2004a, b) are classified to current (2004) Australian Joint Ore Reserves Committee (“JORC”) standards, and reportedly conform to NI 43-101 standards. However, neither Aker Kvaerner, P&E, nor F.H. Brown has verified the classification of any of Snowden’s mineral resources and these resource estimates, therefore, are not treated as NI 43-101 compliant resources verified by a qualified person for the purposes of this report.
 
There are no available mineral resource estimates prepared by Outokumpu for either the Konttijärvi or Ahmavaara deposits. The first published resource estimate was by GFAP in August 2000 (APP, 2002). Since then a number of more or less annual (2000, 2001 2003, July 2004, Dec 2004) mineral resource estimates have been produced on behalf of GFAP by Snowden using then current JORC standards.
 
In 2002, a Suhanko project (Konttijärvi and Ahmavaara deposits) resource update based on 436 diamond drill holes was produced by Snowden for use in a preliminary Bankable Feasibility Study (APP, 2002). The study produced negative results and the 2002 resource estimate was subjected to an external review by Quantitative Geoscience of Australia in 2003. Several criteria were highlighted for investigation; in particular it was noted that the estimation block size selected was too small and the kriging strategy selected was inappropriate. It was further suggested that non-linear methods of resource estimation be investigated, in order to provide local estimates of recoverable (i.e. selectively mineable) resources.
 
These “deficiencies” were corrected in subsequent resource estimate updates in 2003 and July 2004, culminating in a final update in December 2004 which was used in the Lycopodium 2005 Feasibility Study. That study was based on the Snowden December 2004, “recoverable resources” inventory, containing the following estimates for both the Konttijärvi and Ahmavaara deposits.
 
5.6.1        KONTTIJÄRVI
 
The Konttijärvi deposit was estimated by Snowden (Dec, 2004) to contain a total Measured + Indicated recoverable resource of 31.0 Mt with an average 2PGE+Au grade of 2.35 g/t utilizing a 1.0 g/t 2PGE+Au cut-off (Lycopodium, 2005) as shown in Table 5.2. This resource estimate was prepared utilizing a uniform conditioning “recoverable resource” inventory that was derived from the large panel Ordinary Kriging estimate of 25 m x 25 m x 4 m blocks. The recoverable resource inventory was based on a Selective Mining Unit size of 5 m x 5 m x 4 m, and incorporated an assumed 10 m x 10 m x 2 m grade control pattern.
 
24

 
Table 5.2:  GFAP 2004 2PGE+Au Recoverable Resource Inventory – Konttijärvi Deposit (non NI 43-101 compliant)

Konttijärvi	 	 	 	 	 	 	 	 	 	 	 	 	 	2PGE+Au	

 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	 	 	 	 	
Category (JORC Code)	 	(Mt)	 	(g/t)	 	(g/t)	 	(g/t)	 	(%)	 	(%)	 	(g/t)	 	(M oz)	 	
Measured	 	20.2	 	1.79	 	0.51	 	0.13	 	0.08	 	0.17	 	2.42	 	1.58	 	
Indicated	 	10.8	 	1.63	 	0.46	 	0.12	 	0.07	 	 	 	2.21	 	0.77	 	
Inferred	 	10.6	 	1.42	 	0.38	 	0.10	 	0.16	 	0.03	 	1.91	 	0.65	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
1.0g/t 2PGE+Au cut-off	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
 
5.6.2        AHMAVAARA
 
The Ahmavaara deposit was estimated by Snowden (Dec, 2004) to contain a Measured + Indicated recoverable resource total of 54.7 Mt with an average combined palladium + platinum + gold grade of 1.89 g/t utilizing a 1.0 g/t 2PGE+Au cut-off (Lycopodium, 2005) as shown in Table 5.3. The “recoverable resource” inventory was derived from a large-panel Ordinary Kriging estimate of 50 m x 50 m x 4 m blocks. The recoverable resource inventory was based on a Selective Mining Unit size of 5m x 5m x 4m, and incorporated an assumed 10m x 10m x 2m grade control pattern.
 
Table 5.3:  GFAP 2004 2PGE+Au Recoverable Resource Inventory – Ahmavaara
Deposit (non NI 43-101 compliant)

Ahmavaara	 	 	 	 	 	 	 	 	 	 	 	 	 	2PGE+Au	 	

 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	 	 	 	 	
Category (JORC Code)	 	(Mt)	 	(g/t)	 	(g/t)	 	(g/t)	 	(%)	 	(%)	 	(g/t)	 	(M oz)	 	
Measured	 	18.8	 	1.69	 	0.36	 	0.17	 	0.12	 	0.29	 	2.22	 	1.34	 	
Indicated	 	35.9	 	1.29	 	0.27	 	0.15	 	0.09	 	0.27	 	1.71	 	1.98	 	
Inferred	 	22.1	 	1.21	 	0.27	 	0.15	 	0.09	 	0.23	 	1.62	 	1.15	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
1.0g/t 2PGE+Au cut-off	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
 
25

 
6.0           GEOLOGY
 
6.1           GENERAL GEOLOGICAL SETTING
 
The Suhanko project area is located in northern Finland, about 60 km south of Rovaniemi, and comprises the 2.4 Ga old-Suhanko-Konttijärvi layered intrusion rock types and adjacent areas. The Suhanko-Konttijärvi, Narkaus, and Siika-Kama intrusions form part of a group of Paleoproterozoic layered intrusions, which are exposed as an east-west oriented, widespread band extending from north-eastern Sweden to the Kola region in Russia. All known intrusions are located along the northern margin of the Archean basement close to, or in contact with, the Paleoproterozoic supracrustal rocks. These mafic rocks have been interpreted as being emplaced during representing the initial stage of the continental rifting. Many of the intrusions are located close to each other, thus forming igneous complexes in which the individual intrusions were probably connected by dykes or intermediate magma chambers at the time of emplacement. The current distribution of layered intrusions suggests that they are remnants of larger bodies that were preserved within folds and faulted structures.
 
The Archean and early Proterozoic-aged rocks underwent amphibolite facies metamorphism during the Svecofennian deformation period some 1850-1800 million years ago. The Archean basement is mainly composed of tonalite-trondhjemite-granodiorite series, granitoid gneisses and migmatitic gneisses with mafic and felsic enclaves. The Paleoproterozoic supracrustal rocks consist of metamorphosed sedimentary and volcanic rocks, which presently have been metamorphosed to amphibolites, schists and gneisses. Conglomerates containing clasts of layered mafic rocks are exposed along the roof contact of Kemi, Penikat and Narkaus intrusions attesting to erosion of basement and layered intrusions.
 
6.2           REGIONAL GEOLOGICAL SETTING
 
The Suhanko-Konttijärvi layered intrusion comprises the south-western parts of the Portimo layered intrusion complex (Figure 6-1) which consists of the Suhanko-Konttijärvi, Narkaus and Siika Kama intrusions, and dykes of approximately the same age that are named Portimo dykes. The Portimo complex is composed of several megacyclic units that have crystallized from two distinct parental magma types. The first intruded magma type was a high Cr-high MgO magma, which was followed by intrusion of low Cr-low MgO magma. Hanging wall gabbroic cumulates exposed at the present erosion level in the Suhanko-intrusion belong to a low Cr-low MgO megacyclic unit known as the third cycle. The north-eastern parts of the Portimo complex were uplifted to a higher elevation by the tectonic movements after crystallization, and are more deeply eroded. In spite of the regional amphibolite-facies-metamorphism and mineral alteration, igneous textures are generally well preserved in the Suhanko-Konttijärvi layered series rocks. The basement rocks are mainly banded gneisses and late Archean-aged granitoids, which are mostly comprised of diorite and granodiorite.
 
To the southwest of the Suhanko Project, quartzite and impure meta-sediments belonging to the north-south oriented, volcanic rock dominated Archean Oijarvi greenstone belt are encountered.
 
26

 

Figure 6-1:  Suhanko Project Regional Geology
 
The present Archean-Proterozoic contact zone occurs north of the Suhanko-Konttijärvi intrusion. Younger albite diabase dykes and sills cut Archean basement and Proterozoic rocks. Basement granitoids underlying the Suhanko-Konttijärvi region are also cross-cut by mafic to ultramafic sills and veins, which are co-magmatic with layered intrusion rocks, and by granite pegmatite veins.
 
6.3           LOCAL GEOLOGICAL SETTING
 
The Suhanko-Konttijärvi intrusion consists of two bodies separated by Archean basement rocks. The small Konttijärvi deposit is situated 3.5 km northwest of the larger Suhanko body while the Ahmavaara deposit is located at the western edge of the Suhanko Intrusion. The Konttijärvi intrusive is regarded as a faulted offset and subsequent erosional remnant of the Suhanko intrusion, because of the similar stratigraphic succession and style of mineralization as Ahmavaara.
 
6.4           KONTTIJÄRVI PROPERTY
 
The stratigraphy of the Suhanko-Konttijärvi intrusion (Table 6.1) consists of a thick layered series sequence that is dominated by gabbroic rocks with nearly horizontal layering, and of a marginal series with a mineralized unit at the base, extending into the basement.
 
27

 
Table 6.1: Stratigraphy of the Suhanko-Konttijärvi Intrusion

Stratigraphic Unit	 	 	 	Average Thickness	 	Description	
 	 	 	 	 	 	 	
Hangingwall	 		 	Up to 70 m	 	Gabbroic rocks, ranging from melanocratic to leucocratic varieties, with thinner or minor units of pyroxenite and olivine pyroxenite, dominate the Konttijärvi hangingwall rocks. Not mineralised.	
 	 	 	 	 	 	 	
Peridotite Marker	 		 	15 – 50m, avg. 25 m	 	Pyroxenite and olivine Pyroxenite at the top, through peridotite in the middle to olivine pyroxenite at the bottom. Highly magnetic. Sulphide content (mainly pyrrhotite and chalcopyrite) increases towards the base	
 	 	 	 	 	 	 	
Konttijärvi Pyroxenite	 		 	2 – 7 m, avg. 5 m	 	Olivine pyroxenite at the top and pyroxenite at the base. Shows a decrease in magnetic susceptibility compared to PM. Fine sulphides.	
 	 	 	 	 	 	 	
“Chlorite Schist”	 		 	0.2 - 1.0 m	 	Narrow, irregular/anastomosing chloritic shear zone. Generally defines the contact between Pyroxenite and Marginal Series. Marker horizon.	
 	 	 	 	 	 	 	
Marginal Series Gabbro	 		 	0 – 20 m, avg. 10 m	 	Gabbronorite, feldspathic pyroxenite, pyroxenite and assimilated hybrid gabbro. Assimilation features very common. Well mineralized with disseminated chalcopyrite and pyrrhotite, strong hydrothermal overprint.	
 	 	 	 	 	 	 	
Transitional Zone	 		 	0 – 60 m, avg. 10 m	 	Comprises assimilated Marginal Series rocks, mainly hybrid gabbros, and contaminated and mineralized basement rocks. The unit is well mineralized with chalcopyrite and disseminated pyrrhotite.	
 	 	 	 	 	 	 	
Basement	 		 	 	 	Variable Archean rocks, i.e. diorites, banded gneisses and amphibolites, interlayered with hybrid gabbros and mafic dykes/sills (Portimo dykes). Massive hydrothermal and structural alteration evident. The basement rocks host approximately 45% of the mineralization at Konttijärvi.	
 
6.4.1        KONTTIJÄRVI DEPOSIT
 
The Konttijärvi deposit has a strike length of approximately 1000 m. The deposit dips north at between 10° and 20° in the eastern part, to between 30° and 40° in the central and western parts.
 
The northern part of the mineralized body, Konttijärvi North, is divided from the main body by the Konttijärvi fault (normal or N1-fault), which brings the mineralized succession back to surface on the northern side of the fault. Konttijärvi North is divided into two fault blocks, by a reverse fault (R1-fault). The northern extent of Konttijärvi North is terminated against a second normal fault (N2-fault). Both normal faults are inclined towards the south at approximately 50°, although local variations and a faulted zone are present.
 
A reverse fault (R1-fault), in the western part of the Konttijärvi deposit, results in fault block 1, which hosts the deepest known mineralization present. The deepest known layered series mineralization is found approximately 250 m below surface. A cross section through the Konttijärvi deposit showing the boundary faults is shown in the Figure 6-2.
 
28

 

Figure 6-2:  Cross Section of the Konttijärvi Body Showing Fault Blocks
 
6.5           AHMAVAARA PROPERTY
 
The stratigraphy of the Ahmavaara Deposit which summarizes the local geology is given in Table 6.2.
 
6.5.1        AHMAVAARA DEPOSIT
 
The Ahmavaara deposit has an approximate strike length of 2,700 m, of which the easternmost 1000 m is termed Ahmavaara East. The western part of the deposit dips to the northeast with dips ranging from 70° along the south-western margin to 5° in the north-eastern side of the deposit. Ahmavaara East dips north with fairly constant 10 25° dips. The deposits’ western and southern limits are defined by fault-disrupted lithological contacts with Achaean basement rocks. The northern extent is limited by a normal fault (N5-fault), and the north-eastern part is currently terminated against a smaller reverse fault (R3-fault). The deepest known layered series mineralization at Ahmavaara is approximately 500 m below surface. A typical cross section is presented in the Figure 6-3.
 
29

 
Table 6.2: Stratigraphic Unit of the Ahmavaara Deposit

Stratigraphic Unit	 	 	 	AverageThickness	 	Description	
 	 	 	 	 	 	 	
Hangingwall 2	 		 	>50 m	 	Gabbroic rocks, varying between melanocratic to leucocratic varieties with narrower units of pyroxenite and olivine pyroxenite. Not mineralised.	
 	 	 	 	 	 	 	
Upper Peridotite	 		 	10 m	 	 	
 	 	 	 	 	 	 	
Hangingwall 1	 		 	10 – 20 m, avg. 15 m	 	Gabbroic rocks, varying between melanocratic to leucocratic varieties with narrower units of pyroxenite and olivine pyroxenite. Not mineralised.	
 	 	 	 	 	 	 	
Peridotite Maker	 		 	15 – 50 m, avg. 25 m	 	Pyroxenite and olivine Pyroxenite at the top, through peridotite in the middle to olivine pyroxenite at the bottom. Highly magnetic. Sulphide content, mainly pyrrhotite and chalcopyrite, increases towards the base	
 	 	 	 	 	 	 	
Ahmavaara Pyroxenite	 		 	2 – 7 m, avg. 5 m	 	Olivine pyroxenite at the top and pyroxenite at the base. Lower magnetic susceptibility than PM. Fine sulphides	
 	 	 	 	 	 	 	
“Chlorite Schist”	 		 	0.2 - 1.0 m	 	Narrow, irregular/anastomosing chloritic shear zone. Generally defines the contact between Pyroxenite and Marginal Series. Marker horizon.	
 	 	 	 	 	 	 	
Marginal Upper Gabbro	 		 	0 – 30 m, avg. 15 m	 	Gabbros and gabbronorite. Feldspathic pyroxenite are also present at the top contact. The unit is well-mineralized with disseminated chalcopyrite and pyrrhotite. Common xenoliths of Marginal central in lower parts.	
 	 	 	 	 	 	 	
Marginal Central Microgabbro	 		 	0 – 50 m, avg. 25 m	 	Fine-grained, micro-gabbronorite ranging to feldspathic pyroxenite, with thin veins of coarser- grained marginal lower and/or upper, near the top and bottom contacts. The marginal central zone is poorly mineralized. Interpreted as a precursor to the main intrusive.	
 	 	 	 	 	 	 	
Marginal Lower Gabbro	 		 	0 – 30 m, avg. 20 m	 	Skeletal gabbronorite, evidence of assimilation near basement. The contact with the basement may be gradational. Common xenoliths of Marginal central in upper parts. The unit is well- mineralized, with matrix-disseminated and blotchy chalcopyrite and pyrrhotite, occasional massive sulphide lenses.	
 	 	 	 	 	 	 	
Basement	 		 	 	 	Archean gneisses, diorites, banded gneisses, amphibolites and, in the western part of Ahmavaara, quartz-sericite schists	
 
30

 

Figure 6-3: Schematic Cross Section - Ahmavaara Deposit Showing Mineralized Horizons
 
31

 
7.0           DEPOSIT TYPES
 
7.1           KONTTIJÄRVI AND AHMAVAARA DEPOSIT TYPE AND MODEL
 
The Konttijärvi and Ahmavaara Deposits are hosted within the Portimo Intrusive Complex, which is part of a group of 2450 Ga Finnish mafic intrusions derived from a parental magma that is characterized by high MgO and Cr contents and a relatively high SiO 2 content when compared to the MgO content. These types of magma have been referred to either as siliceous high magnesium basalts, basaltic Komatiites or Boninites and have been found to have potential for hosting Cr, Ni-Cu-PGE sulphide, PGE and Fe-Ti-V oxide mineralization. Examples of well known economic deposits derived from similar parental magmas include the deposits hosted by the South African Bushveld, Zimbabwean Great Dyke, Chinese Jinchuan. The types of mineralization mentioned above can all be found in the Finnish intrusions. (http://en.gtk.fi/ExplorationFinland/Commodities/PGE. Nov 2006
 
In the Portimo Complex the magma mixing model for reef-type PGE ore genesis requires that two parental magmas inevitably mixed, at least in the Narkaus chamber. The magma mixing model entails, however, certain problems. One problem is that Reef-type mineralization can sometimes be located in cumulates or gabbroic-pegmatites far below the first signs of a new magma. Another problem is found in the Konttijärvi and Ahmavaara areas, where higher tenor of PGE is sporadically found in the footwall granitoids and Portimo Dikes. These PGE enrichments are disseminated in character and therefore not indicative of an accumulation of immiscible sulphide liquid. Local co-precipitation of U - and Pd - Se minerals suggests co-transportation and mineralization through a fluid phase. The above mentioned features rule out magma mixing as the sole cause of ore formation and it is suggested that the ore-forming processes started much earlier in some intermediate magma chamber in the crust below the Narkaus and Suhanko-Konttijärvi intrusions and the last batch of the Cr-richer magma was already highly enriched in PGE at the time of emplacement.
 
The PGE metals were possibly carried by a fluid phase and/or in small amounts of immiscible sulphide liquids. The multi-million ounce PGE deposit in the Konttijärvi deposit demonstrates that the small size of a mafic body does not make it non-prospective when it is related to large volumes of magma of contained within the whole intrusion belt.
 
7.2           GENETIC MODELS
 
Contact –Type Mineralization
 
The mineralization at the Konttijärvi and Ahmavaara deposits has been widely classified as “contact-type” mineralization. This contact-type of mineralization stands in contrast to the reef-type models by relying less upon magma mixing and more upon the extensive and prolonged interaction of the intrusion with country rock to accounts for the presence of enriched PGE sulphide mineralization (Iljina and Lee 2005). The composition of the country rocks is of significance in this style of mineralization, and Mungall (2005) presented the results of modelling of assimilation of different compositions. Assimilation of sulphur-rich sediment will result in a greater volume of sulphide, commonly forming larger tonnage base metal mineralization with lower PGE grades. In contrast, assimilation of sulphur-poor sediments will result in sulphide saturation of the magma at a lower sulphur concentration, with resultant possible formation of low tonnage, but high-grade PGE mineralization. The results of
 
32

 
Mungall’s (2005) modelling may explain the presence of PGE mineralization in some Fennoscandian mafic magmas such as the Suhanko-Konttijärvi intrusions.
 
The composition of the parental magma in either the contact- or reef-style of PGE mineralization is important in understanding the formation of the mineralization and the tectonic history of the intrusion. Alapieti (2005) suggested that the magmas forming the lowermost parts of the megacyclic type and the ultramafic to mafic type intrusions crystallized from a boninite-like parental magma with high MgO and Cr and low TiO2. A boninite is formed by more than 30% partial melting, at relatively low pressure and temperature, of a fertile mantle harzburgite that has undergone previous significant partial melting in the shallowest parts of subduction zones, or the fore arc area (Mungall 2005). However, the Fennoscandian mafic intrusions are interpreted to have formed in a plume system (Easton 2005).
 
Another possibility is that these magmas formed as a result of a komatiitic magma that assimilated more than half of its own weight in crustal material while crystallizing a similar mass of ultramafic cumulates (Mungall 2005). Distinguishing between the products of these 2 different processes is difficult, as either model would produce magmas of very similar composition.
 
Assimilation of country rocks may not have occurred at the current site of the mineralization, but rather at some mid-crustal level where the magma rose to higher crustal levels through a series of magma conduits. Formation of phases that concentrate the PGE (e.g., sulphides or chromite) may also occur at mid-crustal levels. These phases may then either be entrained and transported by the magma, or deposited and subsequently re-worked with final deposition at the current site of mineralization (Mungall 2005). In these dynamic systems, the physical form of the magma conduit becomes significant, as features such as embayments and transitions from restricted conduits to larger sills or dikes affect the transport of entrained phases and interaction with the wall rocks or previously deposited mineralization. Portions of these conduit systems may be indicated by excessive amounts of cumulate phases and a lack of mafic or felsic fractionates (Mungall 2005),
 
Contact Model Modifications at Konttijärvi and Ahmavaara
 
At Konttijärvi, approximately 50% of the PGM mineralization is contained within a pre-intrusion basement and it is evident that a hydrothermal component of mineralization must be present. Thus, a polygenetic, rather than a simple magmatic sulphide model, has to be applied at Konttijärvi (Robb, 2004) and presumably at Ahmavaara as well, although the basement–hosted mineralization is not as extensively studied.
 
Detailed petrographic work at Konttijärvi shows that mineralization in the mainly dioritic basement is predominantly associated with a hydrothermal epidote-biotite assemblage that is associated with a complex array of both steep and flat dipping ductile structures that appear to become progressively more brittle as successive increments of fluid have passed through the rock.
 
Down-the-hole grade distributions show that the layered mafic intrusion is “bottom-loaded” with mineralization typical of the peridotite and pyroxenitic units, and that high grades are concentrated largely within the lower (marginal) gabbro and transitional (assimilated) zones. The dioritic basement rocks immediately below the mafic intrusion are typified by a prominent hydrothermally altered low-grade mineralized section. Below this zone, the diorite is only sporadically altered and is largely barren, except where discrete chalcopyrite-rich veins and sulphidized amphibolitic zones occur.
 
33

 
The Pd/Pt and Cu/Ni ratios which are thought to reflect an increasing component of hydrothermal remobilization were modeled at the Konttijärvi deposit. The Pd/Pt ratio appears to be the more robust ratio in terms of consistency and for reflecting the polygenetic processes prevailing at Konttijärvi (Robb, 2004). Surprisingly, mineralization in the layered mafic intrusion is characterized by a high variability in the Pd/Pt ratio, with a broad tendency for the ratio to increase (from around 2 to 4) towards the bottom of the intrusion. The hydrothermally altered basement zone immediately below the base of the intrusion is typically characterized by very consistent Pd/Pt (and to a lesser degree Cu/Ni) ratios of around 4. Below the zone of hydrothermal alteration, the basement is either barren (characterized typically by a Pd/Pt ratio of 1.5 which reflects the assay detection limits for the two metals) or is mineralized in discrete zones represented either by amphibolitic (Fe-rich) layers or chalcopyrite rich veins/segregations in the basement (Robb, 2004).
 
Contact Type Mineralization - Model Characteristics
 
The following concluding remarks can be made regarding the characteristics of the “Contact type” mineralization (Easton and Hart, 2005):
 
•       The Marginal Series hosted, ‘contact-type’ PGE deposits are generally zones which are tens of metres wide and have developed at the base or sides of mafic layered intrusions. The zones are erratic in nature and in individual drillholes the highest PGE values can be found tens of metres above or below the contact of the intrusion; they are also variable along the strike.
 
•       The contact-type mineralization is typically base metal sulphide bearing, often enriched in copper, but Ni is dominant in the case of the Platreef and Sheba’s Ridge.
 
•       The PGE concentrations are lower than in the reef-type deposits and the exploitability is based on the large tonnages. These features lead to a broad geophysical signature and the possibility of using elements like copper, nickel and sulphur as pathfinders in geochemical exploration.
 
•       The PGE assemblage is dominated by the low temperature assemblages of Pd–As, Pd-As-Sb, Pd-Bi-Te and Pt-As minerals, with the sulfides and Fe-alloys being rarer.
 
•       Contact-type PGE enrichments seem to be related to an aerially larger igneous event, but the size of the hosting intrusion is not necessarily the unambiguous requirement.
 
•       The key conclusion is that the host rocks of contact-type PGE deposits are characterized by extensive and prolonged interaction of mafic magma with the surrounding host rock. This results in thick marginal zones, which also show compositional reversals in modal mineralogy and mineral chemistry. The absolute PGE contents and relative PGE content over the base metals vary from deposit to deposit in which the Platreef and Portimo stand as examples of more highly and, particularly, more pervasively mineralized.
 
34

 
8.0           MINERALIZATION
 
8.1           OVERVIEW AND BASE METAL SULPHIDES
 
The platinum-group element (PGE) mineralization at both Ahmavaara and Konttijärvi is associated with Cu and Ni mineralization in the form of base metal sulphides. The typical sulphide assemblage is pyrrhotite-chalcopyrite-pentlandite, and accessory sulphides include pyrite, sphalerite, galena and molybdenite. The main oxides are magnetite and ilmenite, with chromite present in trace amounts. The grades of PGE mineralization roughly correspond with the abundance of sulphides, particularly chalcopyrite, and are generally higher at Konttijärvi than at Ahmavaara. Four principal types of PGE mineralization have been identified.
 
•       Sulphide mineralization in the basal parts of the Layered Series (Peridotite Marker, Konttijärvi and Ahmavaara Pyroxenite);
 
•       Sulphide mineralization in the Marginal Series (Marginal Gabbro, Marginal Upper Gabbro);
 
•       Sulphide mineralization in the Konttijärvi Basement; and
 
•       Predominantly massive pyrrhotite mineralization located close to the basal contact of the Ahmavaara intrusion. Mineralization occurs parallel to the intrusion layers.
 
Sulphides present in the Peridotite Marker are pyrrhotite-dominated and are low in abundance. The sulphides in this horizon generally occur as fine-grained disseminations in trace amounts. Sulphide grains are interstitial to former cumulate olivine grains in an assemblage now dominated by talc-carbonate-chlorite after olivine. PGE grades vary from weak to moderate in the basal 2 to 8 meters of the Peridotite Marker, to higher amounts in the upper parts of the sequence. In the underlying Pyroxenite, sulphides are finely intergrown with amphibole (tremolite-actinolite) and chlorite, with minor talc. Sulphide assemblages in the Pyroxenite are dominated by fine-grained chalcopyrite, with PGE grades generally higher than in the Peridotite Marker.
 
Sulphides in the Ahmavaara and Konttijärvi marginal sequences are dominated by an assemblage of pyrrhotite-chalcopyrite-pentlandite and occur as medium grained (in Ahmavaara also coarse grained), disseminated aggregations. Generally the abundances of pyrrhotite and chalcopyrite in Ahmavaara are about equal, though with increasing sulphide abundance the ratio of pyrrhotite to chalcopyrite increases. PGE grades in the Ahmavaara Marginal Series vary, being moderate to high in the Marginal Upper and Marginal Lower units. In the Marginal Central unit both the abundance of sulphides and associated PGE grades are generally low. At Konttijärvi, in the Marginal Series and Transition Zone, chalcopyrite is generally dominant over pyrrhotite and pentlandite. PGE grades there usually vary from moderate to high.
 
Massive sulphide concentrations occur as veins or lenses close to the basal contact of the Ahmavaara intrusion, within both the Marginal Lower and the Basement units. These massive sulphide concentrations comprise dominantly pyrrhotite, with chalcopyrite lamellae and lenses present within the units. PGE grades are high, with individual assays up to 80g/t 2PGE+Au being recorded. In the Marginal Series and Basement units of the Konttijärvi intrusion the sulphides rarely become concentrated into massive sulphide veins or large patches.
 
35

 
The thickness of the Basement-hosted mineralization at Konttijärvi varies, reaching widths of 30 to 40 meters. Sulphides usually occur as fine-medium grained chalcopyrite and pyrrhotite disseminations in the Basement unit and in cross-cutting gabbroic-pyroxenitic intrusives. Pyrite is also present. Platinum-group elements are associated with the sulphides, and the highest values occur in chalcopyrite-rich domains. Upon moving deeper into the Basement, pyrite becomes a dominant sulphide and PGE values decrease to below detection limits. At Ahmavaara, the Basement-hosted mineralization occurs as fine-grained disseminations or as medium-grained sulphide veins. In contrast to the Konttijärvi Basement, the sulphides have relatively high pyrrhotite to chalcopyrite ratios and the PGE ratio is usually low.
 
8.2           PLATINUM GROUP MINERALS
 
Most of the platinum-group minerals at Ahmavaara and Konttijärvi are arsenides, bismutotellurides, and arsenoantimonides. Native forms and alloys are absent. PGE are included in base metal sulphides, magnetite, and silicates and also occur along gangue mineral grain boundaries. Pd-bearing minerals include isomertieite, members of the kotulskite-sobolevskite solid-solution, palladoarsenide, majakite, paolovite, and fengluangite. The principal platinum-bearing mineral is sperrylite.
 
Platinum-group sulphides are rare in the Suhanko-Konttijärvi intrusion and those that have been identified belong to the vysotskite - braggite series.
 
36

 
9.0           EXPLORATION
 
The following is a summary of the work completed by NAP since entering into an agreement with GFAP in early 2006. Previous exploration on the Konttijärvi and Ahmavaara properties is discussed in the History section 5.0.
 
In addition to the initial data compilation work as briefly summarized below, NAP, through its agreement with GFAP, initiated additional on-going metallurgical studies that were supervised by Aker Kvaerner, the results of which are partly included in this report.
 
9.1           GEOLOGICAL DATA COMPILATION
 
All of the available mineralogical data and some of the metallurgical data from Suhanko were compiled and organized into data-binders. These and other metallurgical reports were then documented in a spreadsheet.
 
The compilation exercise documented eight different “stages” of metallurgical work completed in five different “phases”, some of which included mineralogical studies. Most of the metallurgical related work was done on composite samples made up of drill core or bulk samples. It was initially very difficult to match up the mineralogical and metallurgical reports with the original samples since most of the original composites were given one of 11 different sample names (eg. “Konttijärvi Peridotite Marker” sample) that were reused throughout different phases of sampling. This had the potential of creating a great deal of confusion and thus required that the metallurgical and mineralogical sampling be mapped out using flow sheets to track the individual samples.
 
The various phases of sampling involved making composite samples of the 11 litho-types (5 from Konttijärvi, 6 from Ahmavaara). There were also “Master Composite Feed” and “Life of Mine” composite samples made. The vast majority of the mineralogical work was done on mill products resulting from various tests, with a lesser amount of mineralogical work done on the crushed composites themselves. The only QemSCAN analyses were done by Lakefield on 4 composite samples that were derived from 11 composite samples that were in turn derived from 318 individual drill core samples. Both bulk modal analysis and PGE deportment analysis was done. The results from all of these studies have been verified and incorporated into the overall project database.
 
Other studies incorporated into the database include:
 
•       PGE grain distribution study wherein 10 composite samples were analyzed using a SEM equipped with a Gunshot program that automatically searches for PGE grains. Though useful, only a small number of PGE grains were identified.
 
•       Two research studies and one field study by GFAP have been completed on broken rock sample material. These studies, although informative about the distribution of PGE minerals, are too limited in scope to be useful on a deposit scale.
 
•       There are a large number of complete whole rock (XRF) analyses available in the database (2,159 total analyses). These were randomly collected from different units during logging and were done by varying labs at different times and are both within and outside the zone of mineralization. There are also a number of MgO analyses done on “feed” composites and concentrate products.
 
37

 
9.2           MINERALOGICAL AND GEOCHEMISTRY STUDY
 
A mineralogical study titled “Mineralogical and Geochemical Study of the Konttijärvi and Ahmavaara Deposits, Arctic Platinum Project Finland” and dated September 28, 2006 was completed by J. Rickard on behalf of NAP. The results of this study are discussed in detail in Section 23 of this report.
 
The focus of this in-situ mineralogical study was to identify PGE mineral associations and grain interlocking textures that may prove to be problematic in milling and affect PGE recoveries. Of particular interest are the host sulphide and silicate minerals and the relative amount of PGEs locked-up in silicates. It is important to understand the grain size of the PGE, their distribution within the rocks and their host minerals. The mineralization was evaluated by analysis of the in-situ PGE distribution within different rock types from the Konttijärvi and Ahmavaara deposits. Data from various prior studies were used in this study. Some of the historic data was edited and put into a format that was useable to generate comparative graphs and figures between datasets.
 
9.3           SAMPLING AND METHODS
 
Thin section samples were prepared from new dedicated samples of both the Konttijärvi and Ahmavaara deposits. The sample locations were evenly distributed within the limits of the previously planned pits (Lycopodium feasibility study); Konttijärvi (n=39), Ahmavaara (n=43). Due to time constraints only a limited number of samples were scanned in the SEM; a total of 14 samples were analyzed from Konttijärvi and 13 samples were analyzed from Ahmavaara. The thin section slides were analyzed using a Scanning Electron Microscope (SEM) at Lakehead University in Thunder Bay.
 
The PGEs were very small and difficult to locate using a petrographic microscope, so the slides were scanned using backscattered electron imaging to highlight the PGEs. The PGEs were measured and counted, the host silicates and sulphides were identified and pictures taken of the textures. A quantitative analysis was collected for each PGE and the surrounding host minerals. When large clusters of PGE were encountered, a representative number of quantitative analyses were collected and the rest of the minerals were qualitatively identified using the EDS spectra.
 
Summary of Significant Results
 
•       PGEs are found i) completely enclosed by sulphide, ii) along sulphide-silicate grain boundaries, iii) completely enclosed by silicates, iv) scattered as “clusters” within sulphides and silicates.
 
•       The majority of PGE are associated with sulphides, with a smaller proportion being completely enclosed by silicates.
 
•       The PGE distribution (association with sulphides/silicates) is similar between the Konttijärvi and Ahmavaara deposits.
 
•       Average grain size of the PGEs is 12.06µm (n=272), with a range of 0.81µm to 118.0µm.
 
•       PGEs have a strong correlation with chalcopyrite with lesser amounts associated with pyrrhotite, pentlandite, and pyrite, and only a few grains were observed to be attached to sphalerite, galena or millerite.
 
38

 
•       PGEs have a strong correlation with amphibole, with a lesser numbers associated with feldspar, chlorite, epidote, and biotite, and a minor amount with quartz and pyroxene.
 
•       The three gabbroic units at Ahmavaara are all quite similar in terms of composition.
 
39

 
10.0         DRILLING
 
As of the effective date of this report, NAP, through its agreement with GFAP, the project operator, has not conducted any drilling on either the Konttijärvi or Ahmavaara deposits, although diamond drill programs were initiated on both deposits as this report was being finalized and are expected to be completed by mid-April, 2007. A summary of previous drilling on these deposits is contained in Section 5.0 of this report.
 
40

 
11.0         SAMPLING METHODS AND APPROACH
 
The sampling methods and approaches taken by GFAP on the Suhanko project were standardized early in the project and applied throughout the data collection procedures. It is the intent of NAP to continue to utilize the existing procedures after making some minor modifications. In this regard, the core samples from the Konttijärvi and Ahmavaara deposits are subject to a variety of quality assurance - quality control (QA-QC) measures. These measures are built into the sample preparation and analysis scheme, which is detailed in flow chart form in Figure 11-1.
 
GOLD FIELDS ARCTIC PLATINUM OY
CORE SAMPLING PROTOCOL

 
Figure 11-1:  Sample Preparation and Analysis Scheme
 
41

 
12.0                         SAMPLE PREPARATION, ANALYSIS AND SECURITY
 
12.1                         SAMPLE PREPARATION AND ANALYSIS
 
GFAP currently uses two main laboratories for assays from the Suhanko project – the GTK (Finnish Geological Survey) laboratory located in Rovaniemi, and Genalysis located in Perth. All sample preparation is being carried out at GTK and pulps are then shipped to Perth for umpire sampling. Traditionally, PGE (effectively Pd and Pt) and gold are determined by conventional fire assay fusions, using a standard lead flux at both GTK and at Genalysis. The analysis is carried out using the ICP method at both laboratories – GTK uses an optical emission spectrometer (ICP-OES) while Genalysis uses a mass spectrometer (ICP-MS).
 
Base metal assays are also carried out at both laboratories on a second split of the pulp. Both laboratories use an aqua regia digestion followed by ICP-OES finish. It is noted that the aqua regia only provides a partial digestion – in other words, base metals in sulphide and sulphate form are extracted, but not those silicate-hosted species.
 
Snowden carried out two studies of the assay quality assurance data generated by GFAP in 2002 and again in 2004. The 2002 report covered the historical Quality Assurance-Quality Control (QA/QC) data up to July 2002, and made a number of conclusions and recommendations which, for the most part were implemented and the issues peripherally re-examined by Dr Barry Smee in his 2003 investigation and again in the 2004 Snowden study.
 
Both the Snowden 2002 and 2004 studies note that, overall, the QA-QC data shows that the primary (GTK) laboratory is preparing and assaying for PGEs, Au, and base metals diligently. Snowden has also reviewed GFAPs internal core handling, data processing and QA-QC procedures and notes that these are of a very high standard.
 
Independent QA/QC Review – Smee and Associates 2003
 
In February 2003, Dr. Barry Smee of Smee and Associates (Vancouver, Canada) was commissioned to conduct a review of the projects QA/QC procedures. Dr. Smee is acknowledged as a leader in the field of assay QA/QC practices.
 
Dr. Smee concluded that, overall, analysis of quality control data for platinum, palladium and gold has not identified any critical issues which would adversely affect the material outcome of on-going resource estimations.
 
12.2                         SAMPLE SECURITY
 
In Snowden’s (2004) opinion, GFAP takes reasonable steps to ensure the security of samples from the drill site through to the database.
 
42

 
13.0                         DATA VERIFICATION
 
Mr. Fred Brown has visited both the Konttijärvi and Ahmavaara sites (May 4, 2006) and taken independent samples as part of the data verification process associated with the preparation of the 2006 resource estimation used in the current report. Table 13.1 and Table 13.2 summarize the Konttijärvi and Ahmavaara check sample comparisons.
 
In 2004, Snowden conducted an independent site visit of the trial excavations but did not conduct an independent sampling program. They did confirm that sulphide-rich PGE-base metal mineralization was plainly visible in situ. It is also noted that the reconciliation of the trial mining parcels through the GTK pilot plant (BMGS, 2004) demonstrates a reasonable correspondence with both diamond drill assays and blasthole grade control assays. Sampling independent of GFAP from both Ahmavaara and Konttijärvi has been carried out by Lycopodium for the purposes of metallurgical test work. This largely confirms the PGE, Ni and Cu grades as detailed in the current Mineral Resource estimate.
 
The results from numerous internal and independent external reviews (Smee (2003), Snowden (2002) and Snowden (2004)) of QA/QC sampling, sample preparation and sample assaying procedures attest to the validity of the project assay database. For this report and the contained resource estimate, the validity and integrity of the sampling database and the geological information supplied have been assumed. The author has not independently verified the accuracy of the data used; however, sufficient validation has taken place previously for a high level of confidence to be placed in the data, and no reason exists to doubt its veracity and reliability.
 
43

 
Table 13.1: Konttijärvi Verification Samples – Brown 2006

 	 	 	 	al Assays	 	
Drillhole	 	Sample #	 	PD g/t	 	PT g/t	 	AU g/t	 	NI%	 	CU%	 	
KOJ353	 	2053001	 	0.2500	 	0.0700	 	0.0100	 	0.0600	 	0.0400	 	
KOJ543	 	2053002	 	1.6000	 	0.5400	 	0.0800	 	0.0900	 	0.1300	 	
KOJ733	 	2053003	 	0.0060	 	0.0050	 	0.0210	 	0.0072	 	0.0408	 	
KOJ711	 	2053005	 	30.6000	 	9.5300	 	1.1000	 	0.1130	 	0.2090	 	
KOJ749	 	2053006	 	0.6570	 	0.1900	 	0.0180	 	0.0523	 	0.0068	 	
KOJ682	 	2053007	 	0.6700	 	0.1600	 	0.0900	 	0.0400	 	0.3600	 	
KOJ581	 	2053008	 	2.9100	 	0.9300	 	0.2900	 	0.2300	 	0.2400	 	
KOJ262	 	2053010	 	0.3800	 	0.1200	 	0.0040	 	0.0300	 	0.0010	 	
KOJ494	 	2053011	 	1.4300	 	0.4900	 	0.0800	 	0.0500	 	0.1400	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	
 	 	 	 	Assays	 	
KOJ353	 	2053001	 	0.2990	 	0.0708	 	0.0214	 	0.0679	 	0.0511	 	
KOJ543	 	2053002	 	1.4300	 	0.5080	 	0.1930	 	0.0938	 	0.1240	 	
KOJ733	 	2053003	 	0.0116	 	0.0053	 	0.0184	 	0.0070	 	0.0397	 	
KOJ711	 	2053005	 	2.8400	 	1.0100	 	0.0507	 	0.0953	 	0.2080	 	
KOJ749	 	2053006	 	0.7740	 	0.2230	 	0.0179	 	0.0518	 	0.0086	 	
KOJ682	 	2053007	 	2.4700	 	0.6400	 	0.1790	 	0.2580	 	0.6370	 	
KOJ581	 	2053008	 	3.6800	 	1.8200	 	0.2970	 	0.1280	 	0.2630	 	
KOJ262	 	2053010	 	0.2040	 	0.0779	 	0.0001	 	0.0407	 	0.0010	 	
KOJ494	 	2053011	 	1.3800	 	0.4910	 	0.1330	 	0.0531	 	0.1810	 	
Re-assay Crush	 	2053005	 	28.2000	 	8.9700	 	1.0700	 	NA	 	NA	 	
Re-assay Crush	 	2053005	 	27.6000	 	9.1700	 	1.0700	 	0.0930	 	0.1930	 	
 
44

 
Table 13.2: Ahmavaara Verification Sampling, 2006

 	 	 	 	al Assays	 	
Drillhole	 	Sample #	 	Pd g/t	 	Pt g/t	 	Au g/t	 	Ni%	 	Cu%	 	
RN/YP-510	 	2052255	 	1.5200	 	0.2860	 	0.2540	 	0.0932	 	0.4160	 	
RN/YP-290	 	2052256	 	12.5000	 	1.5000	 	0.1230	 	2.7200	 	0.6950	 	
RN/YP-676	 	2052257	 	11.4000	 	0.5080	 	2.0600	 	1.8300	 	1.5100	 	
RN/YP-335	 	2052258	 	1.4000	 	0.2760	 	0.1340	 	0.0854	 	0.2730	 	
RN/YP-176	 	2052260	 	2.5000	 	0.4700	 	0.1000	 	0.2180	 	0.1900	 	
RN/YP-546	 	2052261	 	1.4600	 	0.3970	 	0.0520	 	0.0818	 	0.0932	 	
RN/YP-371	 	2052262	 	8.1100	 	2.3400	 	0.2740	 	0.0921	 	0.1870	 	
RN/YP-463	 	2052264	 	0.1150	 	0.0320	 	0.0310	 	0.0130	 	0.0267	 	
RN/YP-552	 	2052265	 	30.9000	 	8.5700	 	1.8900	 	0.1710	 	1.0800	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	
 	 	 	 	Assays	 	
RN/YP-510	 	4102255	 	1.7100	 	0.3220	 	0.2820	 	0.0908	 	0.4110	 	
RN/YP-290	 	4102256	 	11.6000	 	1.2300	 	0.0360	 	2.7200	 	0.3630	 	
RN/YP-676	 	4102257	 	12.0000	 	0.6890	 	1.1700	 	1.5400	 	1.8400	 	
RN/YP-335	 	4102258	 	1.7700	 	0.3720	 	0.1370	 	0.1070	 	0.2610	 	
RN/YP-176	 	4102260	 	2.1100	 	0.4600	 	0.1420	 	0.2000	 	0.1860	 	
RN/YP-546	 	4102261	 	1.6500	 	0.3970	 	0.0569	 	0.0775	 	0.0789	 	
RN/YP-371	 	4102262	 	9.3600	 	2.1900	 	0.3500	 	0.1060	 	0.1920	 	
RN/YP-463	 	4102264	 	0.0669	 	0.0220	 	0.0286	 	0.0130	 	0.0271	 	
RN/YP-552	 	4102265	 	33.2000	 	9.1700	 	1.9600	 	0.1650	 	0.9680	 	
 
45

 
14.0                         ADJACENT PROPERTIES
 
The Konttijärvi and Ahmavaara deposits are the most investigated mineralized occurrences with the layered Portimo Intrusive Complex and as such are the main focus of the Artic Platinum Project. Given the highly prospective nature of the entire Portimo Complex the Suhanko Project has evolved to incorporate a large regional land position that protects virtually the entire margin of the Portimo Complex including the Konttijärvi, Ahmavaara and other mineralized properties. Consequently there are no immediately adjacent properties to Konttijärvi and Ahmavaara, such as those shown in Figure 14-1, that are not controlled by GFAP.
 
On a more regional scale there are other adjacent properties that are situated within proximal layered mafic intrusions that contain similar PGE or related styles of mineralization (e.g. layered Chromite mineralization). These regionally adjacent properties are displayed in Figure 14-2 while the distribution of related mafic layered intrusions is depicted in Figure 14-3. The varying style of mineralization hosted by the Portimo Complex and adjacent intrusions (Penikat-Kemi and Koillismaa) is briefly discussed below.
 
14.1                         REGIONAL METALLOGENIC SETTING
 
The geological survey of Finland in its Metallogenic map of Finland (Special Paper 35 “Metallogenic zones and metallic mineral deposits in Finland - Explanation to the Metallogenic Map of Finland” places the Konttijärvi and Ahmavaara PGE bearing deposits within the Suhanko PGE zone (38) of the Portimo Complex (see Figure 14-2).
 
14.1.1               MINERALIZATION OF THE PORTIMO COMPLEX
 
The Narkaus PGE zone (39) is also hosted within the Portimo Layered Complex (Iljina 1994). Although situated close to each other, the Suhanko and Narkaus intrusion groups differ in their shape and character of mineralization. The Suhanko intrusions resemble the Koillismaa Complex, while the Narkaus intrusion occupies a position strictly at the Archean/Proterozoic boundary, a position nearly identical to that of the nearby Penikat intrusive complex. Like the Penikat it also is broken up into several large blocks.
 
Five principal types of PGE mineralization have been identified in the Portimo Complex on ground controlled by GFAP.
 
In the Suhanko zone, there are;
 
•                                           PGE-bearing Cu-Ni sulphide disseminations, similar to those in Koillismaa;
•                                           massive pyrrhotite bodies; and
•                                           a reef called Rytikangas Reef, comparable to the Ala-Penikka Reef.
 
In the Narkaus zone, there are :
 
•                                           massive pyrrhotite bodies identical to those at Suhanko;
•                                           a PGE reef (Siika-Kämä Reef); and
•                                           off-set Cu-PGE mineralization in the Archean wall rocks.
 
These styles of mineralization, as briefly discussed below, are also found on properties hosted within adjacent Koillismaa and Penikat-Kemi Intrusive Complexes. These regionally adjacent properties include:
 
Kemi-Penikat Complex
 
The PGE mineralization in the Penikat intrusion is confined to four mineralised horizons or reefs all located inside, or at the boundaries of, one megacyclic unit. The
 
46

 
mineralized horizons are referred to as the Sompujärvi Reef, Ala-Penikka and the Paasivaara Reef. A detailed description and review of the geology of the intrusion and the contained mineralization is presented by Alapieti & Lahtinen (2002). Presently, exploration of the PGE-rich horizons reveals that they continue through the intrusion but, where explored, are mainly thin horizons (decimetres in thickness) with local swellings or potholes up to 20-m thickness. PGE levels are generally less than 10 g/t. It has been estimated by the GTK that the total in situ mineral inventories are around 2 340 t of Pt and 4 590 t Pd (Saltikoff et al., 2006).
 
Kemi Chromite Mine
 
The Kemi mine, established in 1966, operates continuously and up to the present time about 30.91 Mt of ore has been extracted at an average grade of 25.7% Cr2O3. The Kemi deposit is hosted by the single oval-shaped Kemi layered intrusion (Saltikoff et al., 2006). The Penikat intrusion immediately to the east is a long, platform-shaped massif tectonically broken into five blocks. It carries some chromium concentrations as well, but main mineral potential lies in the intrusion’s PGE content.
 
Koillismaa Complex
 
Initially, the Koillismaa complex was explored for nickel-copper and iron-vanadium ores. Indeed, large tonnages of low-grade Ni-Cu mineralization were encountered at Porttivaara and Kuusijärvi in the western subgroup of the intrusive blocks. The PGE potential of the Cu-Ni occurrences was soon realized and it was quickly confirmed that they represent sulphide type Ni-Cu-PGE mineralization in the marginal series of a layered mafic intrusion. Reef-type enrichments in the layered series were disclosed as well. The metal concentrations (Cu, Ni, Au and PGE) and ratios of various properties in the Koillismaa are summarized in the Table 14.1.
 
Table 14.1:  Metal Concentrations and Ratios of Various Prospects in Koillismaa Complex (Saltikoff et al., 2006)

 	 	 	 	Cu+Ni*	 	Cu:Ni	 	Pt+Pd+Au*	 	Pd:Pt	 	Pt:Au	 	Au eq	 	
 	 	(n)	 	(wt.%)	 	 	 	(ppb)	 	 	 	 	 	(ppm)	 	
Haukiaho marginal series	 	116	 	0.60	 	1.56	 	973	 	2.64	 	1.38	 	3.63	 	
Kaukua Marginal series	 	48	 	0.31	 	1.43	 	1,196	 	3.04	 	3.80	 	2.54	 	
Kaukua reef-type	 	9	 	0.33	 	1.67	 	1,293	 	4.03	 	3.35	 	2.68	 	

* Average metal concentrations are also expressed in gold equivalent.
 
47

 

Figure 14-1: Suhanko and Konttijärvi Intrusions - PGE Mineralized Occurrences (Source London Mining Jour, Feb 2005)	 	P&E Mining Consultants Inc.	

 

Figure 14-2: Metallogenic Zones of Finland (Source Geo Surv Finlanf SP 35)	 	P&E Mining Consultants Inc.	

 

Figure 14-3: Fennoscandian Shield Showing Layered Mafic Intrusive Complexes (Source: Andersen et al., 2006)	 	P&E Mining Consultants Inc.	

 
15.0                         MINERAL PROCESSING AND METALLURGICAL TESTING
 
A great deal of metallurgical testing of mineralization from the Konttijärvi and Ahmavaara deposits occurred in the period 2001 through 2005, prior to the involvement of NAP and is well documented in the Lycopodium 2005 Feasibility Study. Due to the amount of work that has been conducted only a brief summary of the results of these tests are presented below.
 
15.1                         INTRODUCTION
 
Lycopodium was commissioned by GFAP to undertake the management of the ongoing metallurgical testwork programme for the Suhanko Project. Metallurgical advice and consultation was provided by GFAP and P Taggart and Associates of Vancouver.
 
Metallurgical testwork has been conducted on the Ahmavaara and Konttijärvi deposits. The deposits contain Platinum Group Elements (PGEs) which includes platinum, palladium and gold, as well as nickel, copper and cobalt. Minor amounts of rhodium are contained in the ore but this element is not evaluated in the testwork and reports although it will provide minor additions to the Project cashflow. A concentrator incorporating primary crushing, grinding, flotation and concentrate dewatering is planned to produce a smeltable concentrate.
 
Previous metallurgical testwork was conducted on samples from the Ahmavaara and Konttijärvi deposits in the following programmes:
 
•                                           Stage 1 by Mintek (South Africa) in 2000.
•                                           Stage 2 by PRA (Canada) in 2001.
•                                           Stage 3 by VTT (now GTK, Finland) in 2001 and 2002.
•                                           Phases 1 and 2 by LRA (now SGS, South Africa) in 2001 and 2002.
 
A summary of these testwork programmes has been reported by Lycopodium in “Past Metallurgical Testwork Review, June 2002”.
 
The Phase 3 testwork programme was conducted by LRA in 2002 to develop a processing scheme suitable for treatment of the Suhanko ore types and this work was reported in the 2002 Draft Feasibility Study (FS2002).
 
The Phase 3 testwork results indicated that:
 
•                                           A grind size P80 of 75 µm produced the best economic return.
•                                           Single stage milling – flotation (MF1) was selected in preference to two stage milling – flotation (MF2).
•                                           Regrinding of the rougher and scavenger concentrates to a product P80 size of 20 microns was selected as producing the best economic return.
•                                           Rougher and cleaner flotation pH of 7.0 and 6.0 respectively, provided an optimum grade / recovery relationship.
•                                           Rougher / scavenger flotation times of 28 minutes produced optimum recoveries.
•                                           Four cleaning stages were selected to allow final concentrate specifications to be achieved.
 
The process design for the FS2002 study and the current feasibility study has been based on the Phase 3 results, but validated in the subsequent test programmes.
 
Three separate test phases have been conducted since the Phase 3 test programme:
 
•                                           The Phase 4 laboratory testing.
•                                           The GTK pilot plant and laboratory bench testing.
 
51

 
•                                           The Phase 5 laboratory bench testing at SGS Lakefield.
 
Phase 4 Test Programme
 
The overall objectives of the Phase 4 testwork programme were to:
 
•                                           Validate the processing scheme developed for treatment of the Suhanko ore types in the Phase 3 works particularly for the Life of Mine ore composite.
•                                           To evaluate alternative processing options.
•                                           Conduct development and variability testwork on a complete new set of Phase 4 variability lithological samples.
 
Whilst the majority of the Phase 4 works were conducted at SGS, other programmes covering aspects of mineralogical liberation size, copper nickel separation, nickel leaching and gravity separation were investigated by other laboratories.
 
GTK Test Programme
 
The GTK testwork covered two aspects, namely the pilot plant and laboratory bench testworks. The ore processed was obtained from the 2004 Suhanko Trial Pits.
 
The main objectives of the GTK pilot plant were two-fold:
 
•                                           Produce concentrate samples for evaluation by smelters.
•                                           Provide engineering design data.
 
The main objectives of the GTK bench scale testwork were:
 
•                                           To determine a baseline performance against which to measure the pilot plant performance.
 
Phase 5 Test Programme
 
The Phase 5 testwork programme was conducted at SGS Lakefield. The overall objectives were to conduct a series of rougher scavenger kinetic tests, open circuit cleaner tests and locked cycle tests on a range of samples in order to revise and upgrade the Grade – Recovery Modelling data base. The samples tested were:
 
•                                           2004 Suhanko Trial Pit lithology and blends of these lithological samples.
•                                           Phase 4 Variability samples and blends of these samples.
 
15.2                         SUMMARY
 
The main purpose of the testwork programme completed since the 2002 study was to build on the knowledge base and to address questions which were outstanding from that programme. The results of this programme have provided a sound basis for the design of the concentrator for the Study. No significant changes to the flowsheet have resulted from the additional testwork.
 
15.2.1               SAMPLE SELECTION AND PREPARATION
 
All stratigraphic and variability sample selection from drill core samples and the trial pit in the Konttijärvi deposit for the Phase 3, Phase 4 and Phase 5 (on the 2004 Trial Pits samples) and subsequent metallurgical testwork was undertaken by Arctic Platinum
 
•                                           Arctic Platinum has confirmed that the samples were collected using approved procedures and the samples are representative of the ore characteristics of each deposit (not withstanding any restrictions on core availability).
 
52

 
•                                           Sample preparation techniques applied by VTT for preparation of sub samples of the original Phase 3 drill core samples from both orebodies were not witnessed by Lycopodium. Nevertheless general written procedures reviewed by Lycopodium were in keeping with standard laboratory procedures
 
•                                           Sample preparation techniques applied by GTK for preparation of the samples from the 2004 Trial Pit lithological samples were witnessed by Lycopodium and are considered to have been conducted appropriately.
 
•                                           Phase 3, Phase 4 and Phase 5 testwork samples made up at SGS from the original drill core (Phase 3 and 4) and Trial Pit samples (Phase 5) selected by Arctic Platinum were composited based on data supplied by Arctic Platinum and Lycopodium. Selection criteria for samples made up to investigate particular characteristics were developed by Lycopodium and Arctic Platinum.
 
•                                           Variability samples were prepared to determine the metallurgical implications of stratigraphy, grade and feed composition.
 
•                                           The following limitations of samples tested exist:
 
•                                           Original sample selection of drill core for the Phase 3 and subsequent Phase 4 test programmes by Arctic Platinum did not spatially cover both deposits due to drilling limitations. Validation of metallurgical performance of stratigraphic units within sample areas not covered by the Phase 3 or Phase 4 testwork programme is recommended to confirm metallurgical characteristics for these stratigraphic units within these areas.
 
•                                           Selection by Arctic Platinum for the 2004 Trial Pit samples could not inherently spatially cover both deposits by nature of the small size mining pits. The subsequent GTK and Phase 5 metallurgical test programme results achieved with these samples may not represent the complete orebody.
 
The drill core samples collected and composited from the respective orebodies provide a reasonable coverage of those orebodies for the testwork conducted and subsequent plant design. Areas of the deposits not already included in the test programmes should be reviewed in the period following Project approval and before mining of those areas commences.
 
15.2.2               MINERALOGY
 
No further mineralogical investigations have been conducted since the Phase 1, 2 and 3 test programmes which demonstrated that:
 
•                                           Major sulphide minerals include chalcopyrite, pentlandite, and pyrrhotite.
 
•                                           The majority of the PGEs occur as platinum in sperrylite, gold in electrum and palladium in kotulskite and isomerteite.
 
•                                           Major gangue minerals include talc, chlorite and tremolite in the upper stratigraphy with additional actinolite, hornblende and quartz in the lower stratigraphy.
 
•                                           The liberation analyses conducted on the orebody lithology samples milled to P80 of 75 µm indicated that acceptable liberation had been
 
53

 
achieved and that further grinding to release small (5 µm) occluded particles is not justified.
 
•                                           Studies on flotation tailings samples indicated that the PGEs lost to tails occurred as very fine inclusions within silicates and release of these particles would require grinding to much finer sizes.
 
Some of the lithologies have not been subjected to detail mineralogical evaluation and whilst the mineralogical data collected is seen to provide a broad basis suitable for process design, it is recommended that lithologies not already subjected to mineralogy be reviewed in the period following Project approval to identify any similarities or differences with lithological types already tested.
 
15.2.3               COMMINUTION
 
Other than two Bond Ball Mill Work Index tests conducted on two lithological samples selected from the 2004 trial pit samples which demonstrated indices within the envelope of those historically tested no further comminution testing has been undertaken since the 2002 Phase 3 test programme. The results of the comminution assessment in 2002 are summarised below:
 
•                                           Similar mineralized stratigraphy is present at both Ahmavaara and Konttijärvi, and comminution testing indicated limited variability between ore types and locations. Samples were classified as possessing average to high compressive strength, moderate resistance to impact breakage, and requiring below average amounts of grinding energy. Properties ranged from the easiest, peridotite marker, to the toughest, hybrid gabbro.
 
•                                           Each of the samples tested indicated high to extreme competency at coarse size, competent enough to generate good media for autogenous grinding and likely to require upwards of 16 kWh/t in a primary autogenous mill. Critical size pebble generation will be likely during milling, although pebble strength is not high. A ball charge and/or recycle crusher should control any tendency to develop circulating loads. Control of feed size will be critical to successful SAG milling.
 
•                                           An SABC grinding circuit was proposed for these ores with a large diameter, high aspect SAG mill to facilitate impact breakage of the coarse lumps. Pebble crushing will be required to contend with the moderate impact resistance of the potential critical size material and the high rod: ball work index ratio observed.
 
•                                           The most comprehensive grinding assessment completed to date was that completed by OMC for FS2002 in which they specified SAG mill power of 16 MW. Metso reviewed the grinding in 2003 on request of GFAP and they recommended SAG mill power of 13.5 MW. This difference in approach prompted a further review which was conducted by SMCC in 2004 and they recommended a SAG mill power of 18.5 MW. GFAP decided to conclude the current feasibility study using the grinding circuit design as per FS2002 with SAG mill size of 16 MW as determined by OMC which would be able to process blended ores at the design rate but not process the harder ores (ie; hybrid gabbro) at any more than 1,100 tph.
 
54

 
It is recommended that:
 
•                                           If the Project proceeds to final design, SMCC be requested to complete the detailed modelling of the circuit to further define the grinding circuit requirements.
 
•                                           In light of the relatively high f SAG value that has been determined further review of means to reduce SAG mill feed size be considered.
 
15.2.4               PHASE 4 FLOTATION TESTWORK
 
The purpose of the Phase 4 flotation programme was to confirm the validity of the Phase 3 results on a wider spectrum of samples and evaluate additional processing options. The Phase 4 flotation testwork programme generally validated the processing scheme developed for treatment of the Suhanko ore types in the Phase 3 works and the main flotation design criteria are summarised in Table 15.1.
 
Acid Washing of Rougher – Scavenger Concentrate
 
Acid washing of the rougher-scavenger concentrate (prior to the cleaning circuit) from a talcose ore composite was reviewed and it was determined that:
 
•                                           There is potential to increase cash flows in the order of $1 .00/t of ore under the acid wash conditions evaluated when compared to non –acid wash regime.
 
The implications of the potential requirements for neutralisation of tailings with lime will need to be considered further if this process is to be pursued.
 
•                                           To confirm the benefits of the acid wash circuit the following items should be reviewed:
 
•                                           verification of the optimum acid wash conditions
•                                           capital cost of an acid wash circuit
•                                           operating cost (including maintenance) of an acid wash circuit.
 
Testwork on this process option was stopped on advice from GFAP.
 
55

 
Table 15.1:  Main Flotation Design Criteria

Criteria	 	Unit	 	Selection	
Grinding	 	 	 	Single Grind-Float	
Grinding Regime	 	 	 	75	
Primary Grind Product Size	 	P80, µm	 	 	
Roughing Scavenging Flotation	 	 	 	28	
 	 	 	 	 	
Flotation Times (laboratory)	 	minutes	 	33	
Pulp Density	 	Wt% solids	 	7	
Pulp pH	 	 	 	Sulphuric Acid	
pH Modifier	 	 	 	Guar	
Depressant	 	 	 	675	
Depressant addition rate	 	g/t	 	SIPX	
Collector	 	 	 	350	
Collector addition rate	 	g/t	 	Senkol2	
Promoter	 	g/t	 	200	
Promoter addition rate	 	 	 	DF 200	
Frother	 	g/t	 	30	
Frother addition rate	 	 	 	 	
Roughing Scavenging Concentrate Regrind	 	 	 	20	
Regrind Product Size	 	P80, µm	 	 	
Cleaning Flotation	 	 	 	 	
Number of Stages	 	No.	 	4	
Flotation Times (laboratory)	 	 	 	 	
Cleaner 1	 	minutes	 	10	
Cleaner 2	 	minutes	 	8	
Cleaner 3	 	minutes	 	6	
Cleaner 4	 	minutes	 	5	
Pulp pH	 	 	 	6	
pH Modifier	 	 	 	Sulphuric Acid	
Collector	 	 	 	SIPX	
Collector addition rate	 	g/t	 	60	
Promoter	 	 	 	Senkol2	
Promoter addition rate	 	g/t	 	15	
 
Variability Testing
 
Variability testwork on 16 individual Ahmavaara Phase 4 variability samples demonstrated that:
 
•                                           The high sulphur grade samples achieved the lowest sulphur recovery and the highest concentrate sulphur grades.
•                                           The lowest sulphur feed grade samples gave the highest PGE recovery (with the exception of the extremely high sulphur feed grades) at the final concentrate stage.
•                                           It is possible that at high sulphur feed grades increased losses of PGE may occur through the cleaning stages due to the rejection of sulphide minerals. The PGEs are believed to be lost due to entrainment or mineral associations with the sulphides.
 
Variability and optimisation testwork was also conducted on the Phase 4 variability samples using three different blends prepared to represent the Konttijärvi and the Ahmavaara ore bodies and a blend of the two in a nominal 35 to 65 wt% ratio. These
 
56

 
blends were selected to represent early, middle and late stages of mine development. The effects of changes in variable conditions tested can be summarised as follows:
 
•                                           Higher flotation pH’s caused small reductions in PGE performances and the standard regime of rougher pH 7.0 and cleaner pH 6.0 was confirmed.
•                                           The primary grind size P80 of 75 µm was confirmed.
•                                           Results for all three blends indicate that grades and recoveries were generally not affected when the P80 of the cleaner feed regrind was coarsened to 25 µm.
•                                           Minor recovery improvements resulted with increased collector and promoter additions. This should be optimised as part of plant-commissioning.
•                                           The use of Depramin 347 as talcose mineral depressant was not beneficial.
 
The variability testing generally validated the process design adopted as a result of the Phase 3 test programme, although variables such as concentrate regrind size, sulphide (pyrrhotite) mineralogy and reagent addition rates were highlighted as warranting further review in the period following Project approval and before commissioning to provide more information for planning purposes.
 
Gravity Recovery
 
The gravity testwork conducted by Knelson Concentrators indicated a gold recovery of 44% (at 3.2 g/t Au), a Pd recovery of 27% (at 16.7 g/t Pd) and a Pt recovery of 28% (at 4.1 g/t Pt). It was decided that although the Au recovery was reasonable, the Pt and Pd were marginal and on the basis that these metals could be recovered in flotation and with most losses to flotation testwork tailing associated with particles locked in silicates gravity testwork would not be pursued further.
 
15.2.5               PHASE 5 FLOTATION TESTWORK
 
Samples of the 11 different lithologies for the two orebodies extracted from the 2004 trial mining pits and prepared at the GTK pilot plant at Outokumpu (Finland) were sent to SGS LRA, Johannesburg, for testing. These samples and a range of lithological samples prepared from the Phase 4 variability samples were subjected to a series of standard flotation tests. This work was completed to extend the data base of metallurgical results for the range of lithologies and head grades to assist in Grade - Recovery Modelling and metallurgical performance prediction. In addition a perceived difference in testwork methodology between GTK and SGS LRA was examined, although the GTK results could not be successfully reproduced at SGS LRA.
 
The tests showed that the trial pit samples could be expected to yield lower recoveries than samples tested during Phase 3 and 4. Although results of locked cycle tests differed significantly for different lithologies, weighted average grade - recovery performance was similar to that achieved in open-circuit tests. The locked cycle tests consistently produced higher mass pulls than equivalent open-circuit tests, due to the presence of recirculating loads of sulphide minerals.
 
Due to the greater effect of mineralogical variation no direct relationship between feed grade and recovery could be established and recoveries have been assumed to be independent of 2PGE+Au head grades.
 
Results of additional tests on selected Phase 4 variability samples were incorporated into the grade - recovery models.
 
57

 
15.2.6               GTK PILOT PLANT TESTWORK
 
A programme of pilot testing was conducted at GTK in Finland using samples from the 2004 trial pit operations. At the same time a bench scale programme was conducted with the same samples.
 
Pilot Plant Operations
 
The pilot plant served the purpose of:
 
•                                           Preparing samples of concentrate for downstream processing and smelting tests by potential purchasers of the concentrate.
•                                           Providing valuable design data both from the physical data gathered at GTK and from ancillary testwork conducted by vendors and selected laboratories.
 
The pilot plant was not able to provide any definitive metallurgical performance data because the results were compromised by:
 
•                                           Head grades deviating from expected mill feed grades.
•                                           Poor operation and control of the plant resulting in spillage which compromised the complete metallurgical balance.
•                                           Suspected poor sampling in the plant.
•                                           Unreliable assaying.
•                                           Performance and reliability limitations of pilot plant equipment, particularly grinding mills, reagent distribution pumps, float cell level and air control and slurry pump hopper and pump ability to handle the flotation slurry froths.
 
Laboratory Bench Flotation Testing
 
In addition to the quality control tests providing a baseline for evaluation of the metallurgical performance of the ore blends treated in the pilot plant the main findings of the laboratory quality control testwork were:
 
•                                           The Ahmavaara blend open circuit cleaner tests demonstrated higher recoveries and lower concentrate grades for 2PGE+Au, Cu and Ni(s) compared to the other blends. This difference is suspected to be associated with the higher sulphide head grades of the Ahmavaara blend.
•                                           The Konttijärvi blend produced the lowest recoveries and highest concentrate grades for 2PGE+Au, Cu and Ni(s) compared to the other blends. As expected the AK blends lie in between the two.
•                                           The copper performance was very similar for all blends.
•                                           For the Konttijärvi blend, the alternative reagent regime of CMC, SEX and Danafloat produced slightly better 2PGE + Au metallurgical performance than the standard reagent scheme. Danafloat also appears to lift performance with the Konttijärvi / Ahmavaara blend.
•                                           For the Konttijärvi blend the best cleaning circuit performance was achieved by the use of the stainless steel regrind mill and without addition of depressant to the cleaning stages.
•                                           Higher guar additions produced lower rougher scavenger recoveries for all metals with both the Ahmavaara–Konttijärvi and Konttijärvi-Ahmavaara blends.
 
The pilot plant and laboratory testwork particularly highlighted the effect of talcose or non sulphide mineral depressants and as result it is recommended that the type, addition rate and point of addition of this reagent should be further investigated in
 
58

 
the period following Project approval and before commissioning. This is not expected to have any significant effect on capital or operating costs.
 
15.2.7               COMPARISON WITH PREVIOUS TEST PHASES
 
The results of standard open circuit cleaner batch flotation testwork conducted on Phase 3, 4 (Variability) and 5 (Trial Pit) samples have been compared for each of the 11 Suhanko lithologies. A global comparison of results for the three phases has been made by mathematically combining the data from test results on individual lithologies, weighted according to their LOM distribution. It should be noted that these results do not represent the outcome of tests on actual composites, but are used to compare the trends in the three test phases in a quantitative way.
 
Comparisons have been drawn between the cleaner 4 concentrates in different tests. Mass pulls generally differ by less than 0.2% and the difference in recoveries is the overriding feature of the comparison.
 
Recovery of (2PGE+Au) to cleaner 4 concentrate from Konttijarvi LOM weighted average lithologies was similar for the three phases (3, 4 and 5) at around 68%.
 
Ahmavaara Phase 3 samples, at 77% (2PGE+Au) recovery, yielded 8% higher recoveries than Phase 4 and 6% higher recoveries than Phase 5 samples, based on cleaner 4 concentrate production.
 
Phase 3 weighted LOM copper recoveries were 19% higher than Phase 4 and 17% higher than Phase 5, based on cleaner 4 concentrate production.
 
Phase 3 weighted LOM nickel (t) recoveries were 4% higher than Phase 4 and 13% higher than Phase 5, based on cleaner 4 concentrate production.
 
The differences in recovery point to likely differences in mineralogy of the samples as a result of their different origins. This was noticeable in higher MgO content and depressant addition requirements with Phase 4 and 5 samples.
 
In comparing these results the origin of samples and purpose of the testwork in each case should be taken into consideration.
 
•                                           Phase 3 samples were composited to approximate the average composition per lithology at the time and still approximate the currently scheduled average feed best in this regard. These samples originate from a limited area of the proposed pits.
•                                           Phase 4 samples were produced from a wider spatial distribution in the ore body, but these tests were conducted to probe the effect of variability in (2PGE+Au) head grade on recovery. Head grades are thus different from those of the other samples.
•                                           Phase 5 samples were obtained from the 2004 Trial Pit Mine. Compared to the average scheduled grades, these samples are often quite different. The main features of these Phase 5 samples are that:
 
•                                           Konttijärvi ore has lower base metal content than predicted by the mine schedule.
•                                           Ahmavaara lower lithologies contain significantly higher S (pyrrhotite) than the schedule.
 
59

 
15.2.8               ANCILLARY TESTWORK
 
To generate additional design data required for the feasibility study, samples were taken during the 2004 GTK Pilot Plant campaign and subjected to testing by various vendors and laboratories.
 
The following tests were conducted:
 
•                                           Determination of stream solids densities by GTK.
•                                           Regrind Testwork by Xstrata (Isamill), Metso (Vertimill and Stirred Media Detritor (SMD) and AMMTEC (ball mill).
•                                           Tailings Thickening Characteristics by Outokumpu and Dorr-Oliver Eimco.
•                                           Concentrate Thickening by Outokumpu and Dorr-Oliver Eimco.
•                                           Filtration by Larox and Metso.
•                                           Concentrate Characteristics for Bulk Shipping by ATSIS and Lycopodium / AMMTEC.
•                                           Concentrate Drying Testwork by Drytech and Metso.
•                                           Concentrate Assays for Marketing by Genalysis.
•                                           Tailings Acid Generating and Leaching Potential under direction of Knight Piésold.
 
Results of this testwork have been incorporated into the concentrator design.
 
15.2.9               RECOMMENDED ADDITIONAL TESTWORK AND PROCESS REVIEW
 
At this stage Lycopodium believes that the testwork completed to date is of a definitive standard and provides a good understanding of the metallurgical response of the orebodies. However to provide more information for planning purposes additional metallurgical testwork and process review in the following areas is recommended to be conducted following Project approval and prior to commencement of operations Most of this work should consist of standard flotation testwork, generally based on variability testing and optimisation of reagent schemes unless noted otherwise:
 
•                                           Following Project approval and prior to final design, SMCC should be requested to complete the detailed modelling of the comminution circuit to further define the grinding circuit requirements and in light of the relatively high f SAG value that has been determined review means and practicality of reducing SAG mill feed size.
 
•                                           The effect of talcose or non sulphide mineral depressants was particularly highlighted by the 2004 GTK pilot plant and laboratory testwork and as result it is recommended that the type, addition rate and point of addition of the depressant should be further investigated in the period following Project approval and before commissioning. This is not expected to have any significant effect on capital or operating costs.
 
•                                           Some of the lithologies have not been subjected to detail mineralogical evaluation and it is recommended that these should be subjected to mineralogical review to identify any previously unrecognised mineral associations or intergrowths.
 
•                                           Samples of ore from areas of the Konttijärvi and Ahmavaara deposits that are currently included in resource estimates and that have not been subject to any metallurgical testwork to date. Flotation and general grinding characteristics should be investigated.
 
60

 
•                                           Further flotation optimization of stratigraphic blends that have demonstrated lower recoveries for selected metals such as platinum and nickel may be warranted. Losses associated with stages within the flotation circuit (eg cleaning) should be addressed.
 
•                                           If target concentrate specifications received from the preferred smelter differ from the existing concentrate grades, flotation variability testwork should be conducted on each stratigraphy to confirm that the revised specifications can be met and to determine metal recoveries and reagent consumption required.
 
•                                           Flotation testing of ore body and stratigraphic blends representing periods of mine production (eg by quarter) may be warranted to determine the effect of blending and to derive expected metal recovery and reagent consumption for each period. This need not necessarily be done prior to commissioning but should be done before mining of that material.
 
•                                           The type of non silicate mineral depressant reagent and point of application in the circuit should be reviewed to examine opportunities for improved performance or reduced unit cost.
 
•                                           Concentrate regrind size has been identified as one area of potential review and the Phase 4 variability testwork programme identified similar results at a coarser grind P80 of 25µm compared to the design 20µm. Further review of concentrate regrind size following Project approval may identify potential capital and operating cost savings achievable with a coarser size.
 
•                                           The primary grind size P80 of 75 µm has been evaluated and selected based on extensive testwork and preliminary financial analyses. Coarsening of the primary grind size would result in considerable capital and operating cost savings and therefore warrants ongoing investigation following Project approval and the availability of more reliable cost data.
 
15.2.10 ALTERNATIVE PROCESS REVIEW
 
The alternative processes reviewed do not form a basis for the concentrator design and are presented to provide possible direction for future testwork. Whilst none of these alternative processes are currently being investigated, they are summarised as follows:
 
Copper-Nickel Separation
 
The preliminary test programmes conducted at SGS, AMMTEC, GTK and G&T Metallurgical Services (G&T) demonstrated the following outcomes for the different Cu-Ni separation processes investigated.
 
Sequential Copper Nickel Separation
 
The Sequential Copper Nickel Separation process has been demonstrated to be a process in which marketable Cu concentrate grades can be prepared with better than 70% recovery of Cu. The Ni concentrate grades and recoveries are lower than desired with concentrate grades varying between 0.7 to 4% Ni(t) and recoveries varying between 26 to 45% Ni(t). The Ni flotation stage performance has been demonstrated to be highly dependant on the Ni head grade
 
One of the better results was achieved in the final locked cycle test conducted by G&T in 2004 as shown in the following table.
 
61

 

 	 	 	 	Grade	 	 	 	Recovery %	 	
 	 	 	 	 	 	 	 	2PGE +Au	 	 	 	 	 	 	 	 	 	 	 	
Product	 	Wt %	 	Cu %	 	Ni %	 	g/t	 	Mg %	 	Cu	 	Ni	 	2PGE +Au	 	Mg	 	
Cu Conc.	 	0.5	 	24.6	 	0.9	 	131	 	1.8	 	73.6	 	6.1	 	26	 	0.1	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Ni Conc.	 	5.0	 	0.5	 	0.7	 	20.1	 	12.8	 	16.2	 	45.0	 	41	 	8.7	 	
 
The Ni(t) and PGE recoveries in the Sequential Copper Nickel Separation process are currently seen as being slightly lower than for the standard bulk base metal / PGE flotation as per the Study base case.
 
Further work is required to optimise and verify the sequential flotation process if GFAP wish to pursue this process.
 
Bulk Concentrate Reverse Flotation
 
The Bulk Concentrate Reverse Flotation process as originally tested during the 2001 VTT test programme has not been shown to be repeatable. None of the optional reagent strategies investigated has demonstrated acceptable performance and no further work is currently planned on this process.
 
Magnetic Separation
 
Magnetic separation was investigated by processing cleaner tailings samples from a laboratory bench test and the pilot plant. Whilst the tests were of a preliminary nature and no further tests are currently planned for this option, there would appear to be potential to utilise magnetic separation to:
 
•                                           Reject pyrrhotite from cleaner tails and intermediate streams.
 
•                                           Concentrate value metals in intermediate streams.
 
•                                           Reduce pyrrhotite and thereby levels of sulphur discharged to final tails.
 
Nickel Leaching
 
The possibility of leaching the low grade nickel concentrate from the sequential flotation process has been reviewed. A nickel selective leach process was targeted which would produce a nickel intermediate. Other value-elements would be recovered from the leach residue as a saleable concentrate followed by toll smelting.
 
Based on a literature review of available processes pressure leaching in the sulphate system was selected as the most likely technology to provide such a process with the least amount of development work.
 
Preliminary testwork investigated both the Activox® process and the pressure leaching process in the sulphate system.
 
The results indicate that:
 
•                                           Reaction rate is highly temperature dependent.
•                                           Ni dissolution at 200C is extremely rapid and results in total oxidation of sulphur to sulphate with 95% copper and 96% Ni extraction within 15 minutes.
•                                           At 140C the addition of sodium lignosulphonate (lignosol) has a positive effect on rate and extent of reaction.
•                                           At low temperature (1 10C), fine grinding is required to achieve the required extraction within practical leach times.
 
While the preliminary results were encouraging no further work was conducted on this process.
 
62

 
16.0        MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES
 
16.1                         INTRODUCTION
 
Since 2000, a number of mineral resource estimates have been produced for the Konttijärvi and Ahmavaara deposits by Snowden. The 2005 GFAP Feasibility Study by Lycopodium was based on the Snowden December 2004 “recoverable resources” inventory.
 
The recoverable resource inventory for the Konttijärvi deposit was a Measured plus Indicated total of 31.0 Mt with an average 2PGE+Au grade of 2.35 g/t at a 1.0 g/t 2PGE+Au cut-off. This resource estimate was prepared utilizing a uniform conditioning “recoverable resource” inventory that was derived from the large panel Ordinary Kriging estimate of 25-m by 25-m by 4-m blocks. The recoverable resource inventory was based on a Selective Mining Unit size of 5 m by 5 m by 4 m, and incorporated an assumed 10 m by 10 m by 2 m grade control pattern.
 
The recoverable resources inventory for the Ahmavaara deposit was a Measured plus Indicated resources total of 54.7 Mt with an average combined palladium plus platinum plus gold grade of 1.89 g/t at a 1.0 g/t 2PGE+Au cut-off. The “recoverable resources” inventory was derived using a uniform conditioning “recoverable resource” inventory derived from the large-panel Ordinary Kriging estimate of 50 m by 50 m by 4 m blocks. The recoverable resource inventory was based on a Selective Mining Unit size of 5 m by 5 m by 4 m, and incorporated an assumed 10 m by 10 m by 2 m grade control pattern.
 
Although the above mineral resource estimates are written in conformance with the JORC Code and are believed to accurately reflect the contained mineralization on the properties, the work necessary to reconcile and verify the classification of the mineral resources, in compliance with the requirements of NI 43-101, has not been completed by either Aker Kvaerner, F.H. Brown CPG, Pr. Sci. Nat, or P&E and the resource estimates, therefore, can not be treated as NI 43-101 compliant resources. Consequently, the estimates should not be relied upon and there can be no assurance that any of the previous resources, in whole or in part, will ever become economically viable.
 
16.2        2006 NI 43-101 MINERAL RESOURCE ESTIMATE
 
An independent evaluation and a NI 43-101 compliant resource model of the Suhanko project (Konttijärvi and Ahmavaara deposits), at the scoping level of confidence, was requested by NAP and prepared by Mr. F.H. Brown CPG, Pr. Sci. Nat., in mid-2006. The effective date for the resource estimates is September 01, 2006.
 
The resource estimates have been prepared for the use of North American Palladium, Ltd. according to the standards set out in Canadian National Instrument 43-101 and have been classified in conformance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council December 11, 2005.
 
All drillhole related data for both deposits are maintained by GFAP in an MSAccess database that was accessed as required for each of the deposits.
 
63

 
16.3        KONTTIJÄRVI PROPERTY
 
16.3.1               DATA SUPPLIED
 
GFAP geologists supplied Triangulated Irregular Network (TIN) files of the local topography, till/bedrock contact and a structural interpretation of the major fault systems. TIN files are used by the Gemcom mine modelling software program to generate three-dimensional digital models of surfaces (such as topography) and volumes such as a rock unit.
 
All data are maintained by GFAP in an Access database. GFAP supplied a Konttijärvi database containing 857 diamond drillholes, of which 719 diamond drillholes were used for grade modeling (Table 16.1). The extract file contains header, survey, lithology and assay data, with assay data reporting Pd, Pt, Au, Ni and Cu sample grades. The results of validation checks of the database were within industry standards.
 
Table 16.1: Drilling History.

HOLE ID	 	TYPE	 	COUNT	 	MODEL	 	
KOJ/GT	 	Geotech	 	4	 	4	 	
KOJ/MET	 	Metallurgy	 	1	 	0	 	
KOJ	 	Geology	 	555	 	555	 	
RN	 	Historical	 	160	 	160	 	
KRR	 	Blast Holes	 	135	 	0	 	
WH	 	Water Hole	 	2	 	0	 	
TOTAL	 	 	 	857	 	719	 	
 
In order to reduce the random effects of low grade assay values below the assay detection limits for the defined grade elements, a low grade transform was applied to data extracted from the database, approximately equivalent to half the lower detection limit. For Pd, Pt and Au, assay values less than 0.01 g/t were transformed to 0.005 g/t; for Ni and Cu assay values less than 0.005% was transformed to 0.0025%. Blank intervals were flagged as Not Sampled in the database.
 
16.3.2               CONFIRMATION SAMPLES
 
Nine (9) verification samples were randomly selected for analysis from existing core stored at the GFAP Arktikum storage site in Rovaniemi, Finland. The sampling procedures and assay results are summarized in Section 13.1 of this report and the reader is referred to this section for more detailed information.
 
16.3.3               GEOLOGICAL MODEL
 
The geological model was constructed using polygonal interpretations of each lithology on north-south orientated cross-sections that were on 10 m centers. Three-dimensional wireframe solid models of the individual lithological units were then created based on successive polygons, and were combined into a three- Dimensional layered geological model. A TIN model of the Portimo Dyke was supplied by GFAP.
 
Due to difficulties in interpreting continuity in historical data, the Transitional Zone was combined with the Marginal Series for lithological modeling purposes (Table 16.2). Mineralization envelopes and “low-grade zones” have not been modeled or incorporated into the estimation process. All lithological models were clipped (trimmed) to the till/bedrock contact. The Peridotite Marker, Konttijärvi Pyroxenite,
 
64

 
Marginal Series and Transitional Zone were then combined into the single “Marginal Domain” unit for estimation purposes (Figure 16-1).
 
Table 16.2:  Modeling lithology.

Lithology	 	Strat Code	 	Rock Code	 	Color	
Hanging Wall	 	HW	 	1100	 	Blue	
Peridotite Marker	 	PM	 	1200	 	Cyan	
Konttijärvi Pyroxenite	 	KPX	 	1300	 	Green	
Marginal Series + Transitional Zone	 	MS	 	1400	 	Brown	
“Marginal Domain”	 	LS	 	1900	 	Yellow	
Basement	 	BAS	 	1600	 	Magenta	
 
Major faulting divides the deposit into five fault blocks, which were incorporated into the geological model. The bulk of the deposit lies in the South fault block (Figure 16.2).
 
Figure 16-1: Konttijärvi Geological Model (Source: F.H. Brown 2006).

 
65

 
Figure 16-2: Konttijärvi Fault blocks (Source: F.H. Brown 2006).

 
16.3.4               MINERALIZATION MODEL
 
NAP requested that an unconstrained mineralization model be utilized, in order to accommodate the complex mineralization history of the deposit without reference to empirically defined grade boundaries. Two statistical populations were evident from the statistical results and variography. Each fault block was therefore divided into two mineralization domains where applicable, creating a total of nine domains for estimation purposes (see Table 16.10). Fault block boundaries were considered hard boundaries for estimation purposes. The Portimo Dyke was modeled as a zero grade waste intrusive.
 
16.3.5               COMPOSITING
 
Raw sample lengths from the Konttijärvi database ranged from 0.01m to 4.73m, with an average sample length of 1.01 m. A total of 82% of the samples were one meter in length. For estimation purposes the assays were composited to one- meter length-weighted intervals downhole from the collar.
 
16.3.6               GRADE ELEMENTS
 
Palladium (Pd), platinum (Pt), gold (Au), nickel (Ni) and copper (Cu) were estimated separately as individual grade elements. A derived 2PGE+Au (Pd + Pt + Au) parameter was then calculated by combining the appropriate grade elements for each block. As well, a Palladium Equivalent (PdEq) parameter incorporating Pd, Pt, Au, Ni and Cu values was calculated.
 
16.3.7               SPECIFIC GRAVITY
 
Specific gravity values for each lithological unit were taken from Snowden (2004c) as given in Table 16.3.
 
66

 
Table 16.3: Specific Gravity Values Konttijärvi.

Unit	 	Rock Code	 	Density (t/m 3 )	 	
Hanging Wall	 	1100	 	2.91	 	
Peridotite Marker	 	1200	 	2.94	 	
Konttijärvi Pyroxenite	 	1300	 	2.99	 	
Marginal Series + Transitional Zone	 	1400	 	2.95	 	
Basement	 	1600	 	2.80	 	
“Marginal Domain”	 	1900	 	2.95	 	
Portimo Dyke	 	2300	 	3.00	 	
 
16.3.8               SUMMARY STATISTICS
 
Composite grade statistics were compiled for each estimation domain. Samples were not de-clustered prior to analysis (Table 16.4).
 
Table 16.4: Domain Composite Statistics Konttijärvi.

 	 	DOMAIN	 	n	 	Min	 	Max	 	Mean	 	Media n	 	SD	 	COV	 	
Pd	 	Marginal	 	17343	 	0.010	 	77.110	 	1.351	 	0.800	 	2.003	 	1.483	 	
 	 	Basement	 	25902	 	0.010	 	245.000	 	0.724	 	0.170	 	2.885	 	3.987	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Pt	 	Marginal	 	17343	 	0.010	 	26.290	 	0.387	 	0.220	 	0.574	 	1.482	 	
 	 	Basement	 	25902	 	0.010	 	80.500	 	0.198	 	0.040	 	0.924	 	4.675	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Au	 	Marginal	 	17343	 	0.010	 	5.880	 	0.099	 	0.050	 	0.164	 	1.659	 	
 	 	Basement	 	25902	 	0.010	 	19.880	 	0.054	 	0.010	 	0.229	 	4.215	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Ni	 	Marginal	 	17343	 	0.005	 	1.240	 	0.069	 	0.050	 	0.057	 	0.826	 	
 	 	Basement	 	25902	 	0.005	 	1.040	 	0.025	 	0.010	 	0.044	 	1.752	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Cu	 	Marginal	 	17343	 	0.005	 	1.740	 	0.113	 	0.070	 	0.133	 	1.177	 	
 	 	Basement	 	25902	 	0.005	 	5.640	 	0.082	 	0.030	 	0.140	 	1.694	 	
 
Probability graphs, separated by lithological domains, indicate the presence of mixed population distributions, as shown in Figure 16-3, possibly as the result of multiple primary mineralization events combined with hydrothermal remobilization. Sensitivity studies were carried out on grade-capping and high grade outliers, but the results suggest that there is no need for grade-capping at this level of investigation and none was applied.
 
A strong correlation exists between composite Pd and Pt assay grades (R 2 = 0.91 for Basement data and R 2 = 0.94 for Marginal Domain data). Correlation was weaker for Pd and Au, and no linear correlation could be demonstrated between Pd and Ni, Pd and Cu or Ni and Cu (Table 16.5).
 
67

 
Figure 16-3:  Konttijärvi Grade Probability Graphs

 
68

 

69

 
Table 16.5: Grade Element Correlation and Ratios Konttijärvi .

 	 	Basement	 	Marginal Domain	 	
 	 	Correlation	 	Pd Ratio	 	Correlation	 	Pd Ratio	 	
Pd:Pt	 	0.911	 	3.3	 	0.936	 	3.6	 	
Pd:Au	 	0.541	 	17.2	 	0.476	 	17.5	 	
Pd:Ni%	 	0.153	 	167	 	0.182	 	82	 	
Pd:Cu%	 	0.148	 	84	 	0.362	 	25	 	
Ni%:Cu%	 	0.247	 	NA	 	0.185	 	NA	 	
 
Figure 16-4: Palladium Correlation Konttijärvi (Source: F.H. Brown 2006). .

 
16.3.9               VARIOGRAPHY
 
Variography was performed on composited values within the South fault block, which contains 82% of the sample data. Lognormal directional semi-variograms were investigated for both the Marginal Domain and Basement mineralization zones for each grade element, and the trend direction of maximum continuity in the horizontal plane determined. Within this trend, continuity was then determined by varying the plunge direction. Directional semi-variograms perpendicular and orthogonal to this trend were then modeled to identify effective ranges in each principle direction.
 
Short-range variography (~100 m or less) results suggest weak to moderate anisotropy, relative to the grade element and the lithological unit. Au was effectively isotrophic at short ranges for both lithological units, with moderate nugget effects. Pd and Pt Marginal Domain mineralization appears to be weakly anisotropic, and both elements display high nugget effects in the Marginal Domain. Ni also displayed a
 
70

 
moderate anisotropy in the Marginal Domain at the ranges investigated. In general, mineralization at short ranges is very weakly anisotropic in the Basement unit, and weakly anisotropic in the Marginal Domain. The horizontal grade continuity in the Marginal Domain has a trend approximately parallel to the local strike. The vertical grade continuity plunges -10° to -30° (Table 16.6). No clearly defined mineralization model could be discerned from the short-range variography. For this reason, short-range non-directional semi-variograms were deemed sufficient for estimation purposes in this study, as shown below in Figure 16-5 and tabulated in Table 16.7.
 
Table 16.6:  South block short-range grade continuity Konttijärvi

Element	 	Domain	 	Anisotrophy	 	Trend	 	Plunge	 	
Pd	 	Basement	 	Weak	 	105°	 	-30°	 	
 	 	Marginal Domain	 	Weak	 	90°	 	-10°	 	
 	 	 	 	 	 	 	 	 	 	
Pt	 	Basement	 	~Isotrophic	 	90°	 	-10°	 	
 	 	Marginal Domain	 	Weak	 	90°	 	-20°	 	
 	 	 	 	 	 	 	 	 	 	
Au	 	Basement	 	~Isotrophic	 	90°	 	-10°	 	
 	 	Marginal Domain	 	~Isotrophic	 	75°	 	-20°	 	
 	 	 	 	 	 	 	 	 	 	
Ni	 	Basement	 	~Isotrophic	 	90°	 	-10°	 	
 	 	Marginal Domain	 	Moderate	 	75°	 	-10°	 	
 	 	 	 	 	 	 	 	 	 	
Cu	 	Basement	 	~Isotrophic	 	75°	 	-10°	 	
 	 	Marginal Domain	 	Weak	 	75°	 	-10°	 	
 
Table 16.7:  South block estimation semi-variograms Konttijärvi .

Element	 	Domain	 	C 0	 	Semi-variogram	
Pd	 	Basement	 	38	%	C 0 3.2 + SPH(45, 5.1) + SPH(99, 0.09)	
 	 	Marginal Domain	 	74	%	C 0 3.25 + SPH(40, 1.1) + SPH(99, 0.06)	
 	 	 	 	 	 	 	
Pt	 	Basement	 	31	%	C 0 0.27 + SPH(42, 0.53) + SPH(99, 0.06)	
 	 	Marginal Domain	 	77	%	C 0 0.28 + SPH(40, 0.08) + SPH(99, 0.002)	
 	 	 	 	 	 	 	
Au	 	Basement	 	57	%	C 0 0.035 + SPH(13, 0.025) + SPH(99, 0.001)	
 	 	Marginal Domain	 	41	%	C 0 0.012 + SPH(10, 0.017) + SPH(99, 0.0002)	
 	 	 	 	 	 	 	
Ni	 	Basement	 	34	%	C 0 0.0011 + SPH(35, 0.0008) + SPH(999, 0.0002)	
 	 	Marginal Domain	 	50	%	C 0 0.0012 + SPH(28, 0.0021) + SPH(50, 0.00023)	
 	 	 	 	 	 	 	
Cu	 	Basement	 	39	%	C 0 0.008 + SPH(12, 0.0125) + SPH(99, 0.0001)	
 	 	Marginal Domain	 	45	%	C 0 0.008 + SPH(25, 0.0085) + SPH(99, 0.0014)	
 
71

 
Figure 16-5:  Konttijärvi Experimental Semi-Variograms

 
16.3.10  BLOCK MODEL
 
The Konttijärvi drilling grid is laid out to provide sample information on a nominal 50 m spacing, with infilling to 25 m and 12.5 m in places. A 15 m x 15 m x 8 m (L x W x H) estimation block size was selected as a compromise between (potential) mining selectivity; the narrow width of the Marginal Domain mineralized zone, the existing drilling grid and semi-variogram ranges. The block model framework contains 308,760 blocks covering the defined area of mineralization (Table 16.8).
 
72

 
Table 16.8:  NAP 2006 block model Konttijärvi.

 	 	Min	 	Max	 	Block Size	 	Count Blocks	 	
Easting	 	53300	 	54695	 	15	 	93	 	
Northing	 	36400	 	37645	 	15	 	83	 	
Elevation	 	-112	 	208	 	8	 	40	 	
Volume = 2162 M m 3	 	 	 	 	 	 	 	 	 	
 
Block model geological parameters were constructed from a percent Gemcom™ model, allowing for a high degree of accuracy in terms of volume estimates.
 
16.3.11  CUT-OFF GRADES
 
Due to the polymetallic nature of the deposit the resource inventory was tabulated relative to a 2PGE+Au cut-off for comparative purposes with the Snowden estimates. The resource inventory listing is an arithmetic accumulation of individual grade element estimates from blocks that meet the 2PGE+Au cut-off value.
 
Current NAP operating practice is to also express resources in terms of a combined NSR cut-off. A resource inventory has therefore also been calculated relative to a NSR Palladium Equivalent (PdEq) grade. PdEq grades were calculated from individual grade element values as follows:
 
PdEq = Pd + (3.68 x Pt) + (2.07 x Au) + (20.88 x Ni) + (5.88 x Cu)
 
PdEq parameters were derived from the data given in Table 16.9.
 
Table 16.9:  PdEq NSR parameters* - Konttijärvi.

 	 	 	 	 	 	 	 	 	 	Payable Metal	 	
 	 	Metal Price	 	Concentrate	 	 	 	Pd Eq	 	$US/tonne/% or	 	
Element	 	$US/lb or oz	 	Recovery	 	Net Revenue %	 	Factor	 	gram/t	 	
Pd	 	$	277.00	 	69.0	%	28	%	1.00	 	5.25	 	
Pt	 	$	1,057.00	 	65.0	%	30	%	3.68	 	19.31	 	
Au	 	$	546.00	 	75.0	%	4	%	2.08	 	10.90	 	
Ni	 	$	9.04	 	65.0	%	23	%	20.90	 	109.72	 	
Cu	 	$	2.56	 	73.0	%	15	%	5.89	 	30.91	 	

*Confidential smelter terms: Smelter recovery/payables, treatment and refining charges, penalties, freight costs, etc., used to determine NSR/net revenues per unit are not shown because of their proprietary nature.
 
16.3.12  ESTIMATION
 
The resource estimate was based on 719 diamond drillholes. Estimates were made from three-Dimensional block models using commercial Gemcom™ software version 5.45.
 
Separate in-situ global resource estimates were produced for the defined grade elements for 15 m x 15 m x 8 m blocks using Ordinary Kriging on composite grades. Nearest Neighbour (NN) block estimates were also calculated for comparison and validation purposes.
 
Estimation data were constrained within demarcated estimation domains, incorporating assay data and geological interpretations within the framework of the geological model. Estimation was not restrained to empirical mineralization zones as per the GFAP approach, but was instead allowed to continue on all blocks informed by the defined search ellipse within an individual estimation domain (Table 16.10).
 
The kriging neighbourhood was investigated for the Pd grade element within the South Block Basement and Marginal Domain mineralization zones. Search ellipses
 
73

 
were restricted to 50m, and estimation required a minimum of five samples and a maximum of fifteen samples. The discretization level was set at 4 x 4 x 4. No dilution factor was applied to the estimates.
 
NAP 2006 Global Resource Inventory –Konttijärvi Deposit
 
The Global Resource Inventory for the Konttijärvi deposit, based on a NSR palladium equivalent estimated at a cut-off of 1.0 g/t PdEq produced a Measured + Indicated total of 63.602 million tonnes with an average PdEq grade of 3.747 g/t and an Inferred total of 11.640 million tonnes at an average PdEq grade of 3.104 g/t . The 2006 Global Resource Inventory for the Konttijärvi and Ahmavaara Deposits (1.0 g/t PdEq cut-off ) as prepared by F.H. Brown (2006) is shown in Table 16.11
 
A complete resource inventory by cut-off grade for 2PGE+Au and PdEq estimates is given in Appendix 1.
 
Table 16.10: Estimation Domains .

ESTIMATION DOMAIN	 	 	 	 	 	
Fault Block	 	Mineralization	 	Domain Code	 	
North	 	Basement	 	5100	 	
South	 	Marginal Domain	 	6200	 	
South	 	Basement	 	5200	 	
East	 	Marginal Domain	 	6300	 	
East	 	Basement	 	5300	 	
West 1	 	Marginal Domain	 	6400	 	
West 1	 	Basement	 	5400	 	
West 2	 	Marginal Domain	 	6500	 	
West 2	 	Basement	 	5500	 	
 
Table 16.11: NAP 2006 Global Resource Inventory - Suhanko Project (Konttijärvi and Ahmavaara Deposits) 1.0 g/t PdEq Cut-off

Ahmavaara	 	Tonnes	 	Pd g/t	 	Pt g/t	 	Au g/t	 	Ni %	 	Cu %	 	PdEq g/t	 	PdEq Oz	 	
Measured	 	38,197,000	 	0.707	 	0.141	 	0.086	 	0.061	 	0.161	 	3.425	 	4,206,000	 	
Indicated	 	114,816,00	 	0.851	 	0.175	 	0.104	 	0.071	 	0.181	 	4.027	 	14,867,00	 	
 	 	0	 	 	 	 	 	 	 	 	 	 	 	 	 	0	 	
Inferred	 	34,757,000	 	0.867	 	0.187	 	0.103	 	0.070	 	0.170	 	4.015	 	4,486,000	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Konttijarvi	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	26,031,000	 	1.064	 	0.307	 	0.081	 	0.059	 	0.100	 	4.173	 	3,493,000	 	
Indicated	 	37,571,000	 	0.902	 	0.255	 	0.071	 	0.043	 	0.095	 	3.452	 	4,170,000	 	
Inferred	 	11,638,000	 	0.867	 	0.241	 	0.062	 	0.027	 	0.097	 	3.014	 	1,128,000	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Combined	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	64,228,000	 	0.852	 	0.208	 	0.084	 	0.060	 	0.136	 	3.728	 	7,699,000	 	
Indicated	 	152,387,00	 	0.864	 	0.195	 	0.096	 	0.064	 	0.160	 	3.885	 	19,037,00	 	
 	 	0	 	 	 	 	 	 	 	 	 	 	 	 	 	0	 	
Inferred	 	46,395,000	 	0.867	 	0.201	 	0.093	 	0.059	 	0.152	 	3.764	 	5,614,000	 	
 
The distribution of mineralization within the defined lithological units was also estimated (Table 16.12), as this directly affects the metallurgical characteristics of the mineralization.
 
74

 
Table 16.12: Konttijärvi Lithological unit mineralization – 1.0 g/t 2PGE+Au cutoff (Measured + Indicated + Inferred).

 	 	 	 	 	 	 	 	 	 	 	 	
 	 	 	 	Peridotite	 	 	 	Marginal Series +	 	 	 	
 	 	 	 	Marker	 	Pyroxenite	 	Transitional Zone	 	Basement	 	
Tonnes	 	 	 	15	%	10	%	37	%	39	%	
%
Mineralization	 	Pd	 	14	%	10	%	36	%	40	%	
 	 	Pt	 	15	%	10	%	36	%	39	%	
 	 	Au	 	16	%	11	%	36	%	37	%	
 	 	Ni	 	20	%	13	%	41	%	27	%	
 	 	Cu	 	10	%	9	%	41	%	40	%	
Avg Grade	 	Pd g/t	 	1.505	 	1.641	 	1.633	 	1.711	 	
 	 	Pt g/t	 	0.471	 	0.488	 	0.458	 	0.465	 	
 	 	Au g/t	 	0.133	 	0.134	 	0.124	 	0.121	 	
 	 	Ni %	 	0.085	 	0.081	 	0.072	 	0.045	 	
 	 	Cu %	 	0.102	 	0.133	 	0.163	 	0.163	 	
 
16.3.13  CLASSIFICATION
 
Global mineral resources were classified in accordance with guidelines established by the Canadian Institute of Mining, Metallurgy and Petroleum (CIM, 2005):
 
•                                           Inferred Mineral Resource
 
An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.
 
•                                           Indicated Mineral Resource
 
An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics, can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.
 
•                                           Measured Mineral Resource
 
A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.
 
Kriging efficiency (Krige, 1996) and the geological continuity of the Pd estimate were used as the primary criteria for classification. Kriging efficiency was calculated for each Pd block estimate using the OK estimation variance and the block variance.
 
75

 
Kriging efficiency (KE) and the estimation regression slope (R) were then used to establish the relative confidence of the Pd block estimates, based on the semivariogram parameters and the location of the informing data. Higher kriging efficiencies are associated with better informed blocks, with a value of 100% reflecting perfect theoretical estimation.
 
The following ranges were then used to assign an overall classification level to each block:
 
•                                           Measured: KE >= 75% and R >= 0.8
•                                           Indicated: 30% >= KE < 75% and R >= 0.6
•                                           Inferred: 0.01% > KE < 30%
 
Well-informed blocks, with a low conditional bias, were therefore classified as Measured. Moderately-informed blocks were classified as Indicated. Poorly-informed blocks with less than adequate sample information or a poor spatial distribution of sample information were classified as Inferred. Blocks with a zero or negative KE were deleted from the estimation inventory (Figure 16-6). Only Pd results were used to classify the resource, as Pd is the primary grade element of economic interest.
 
Figure 16-6: Block classification (Section 53802.5, looking east) (Source: F.H. Brown 2006).

 
16.3.14  MODEL VALIDATION
 
Five methods were used to test the validity of the OK model:
 
•                                           Visual inspection of rock codes, composite grades and block grades in multiple north-south sections. No systematic bias was evident in the model estimates.
 
76

 
•                                           Global comparison of average composite grades per block and model block grades at a nominal cut-off (Table 16.13). Results for all grade elements indicate that the model provides a reasonable smoothed global estimate.
 
•                                           Comparison of model block estimates with a Nearest Neighbor (NN) estimate. The average grade of the model estimate and the NN estimate were compared at a nominal cutoff grade of 0.01. The mean grades were within 4.0% for all grade elements, indicating that the estimate is globally unbiased (Table 16.14).
 
•                                           Model block estimates and composite block averages were also compared to average blasthole grades from the 2004 trial mine sampling program (GFAP, 2004). The fifteen blocks affected straddle the South Block Marginal Domain and Basement lithological contact (Table 16.15). Model block estimates were intermediate between the composite average for the blocks and the calculated blasthole average for the blocks.
 
•                                           As a further test of the reasonableness of the model, the average informing distance of the closest sample used for estimation of the Pd grade element was also examined (Table 16.16). The distances used are consistent with industry practice for classification.
 
Table 16.13: Comparison of composite averages and model estimates.

Grade Element	 	Composite Avg	 	Model Estimate	 	
Pd	 	0.945	 	0.818	 	
Pt	 	0.282	 	0.242	 	
Au	 	0.084	 	0.071	 	
Ni	 	0.056	 	0.044	 	
Cu	 	0.108	 	0.089	 	
2PGE+Au	 	1.311	 	1.131	 	
 
Table 16.14: Comparison of OK and NN block model estimates

 	 	 	 	OK	 	NN	 	VAR	 	
Pd	 	Basement	 	0.438	 	0.443	 	1.1	%	
 	 	Marginal Domain	 	0.984	 	0.976	 	-0.8	%	
Pt	 	Basement	 	0.133	 	0.131	 	-1.5	%	
 	 	Marginal Domain	 	0.288	 	0.285	 	-1.1	%	
Au	 	Basement	 	0.042	 	0.041	 	-2.4	%	
 	 	Marginal Domain	 	0.081	 	0.080	 	-1.3	%	
Ni	 	Basement	 	0.028	 	0.027	 	-3.7	%	
 	 	Marginal Domain	 	0.066	 	0.065	 	-1.5	%	
Cu	 	Basement	 	0.071	 	0.069	 	-2.9	%	
 	 	Marginal Domain	 	0.101	 	0.100	 	-1.0	%	
 
77

 
Table 16.15: Comparison of blasthole averages and NAP block model estimates

Grade Element	 	Blasthole Avg	 	Comp. Avg	 	OK	 	NN	 	
Pd	 	1.201	 	1.577	 	1.367	 	1.267	 	
Pt	 	0.390	 	0.520	 	0.446	 	0.414	 	
Au	 	0.094	 	0.114	 	0.099	 	0.098	 	
Ni	 	0.086	 	0.075	 	0.067	 	0.061	 	
Cu	 	0.095	 	0.115	 	0.109	 	0.095	 	
2PGE+Au	 	1.685	 	2.211	 	1.977	 	1.779	 	
 
Table 16.16:  Average informing distance of the closest sample for Pd

 	 	MINIMUM	 	AVERAGE	 	MAXIMUM	 	
Measured	 	0.1	 	7.7	 	20.9	 	
Indicated	 	1.9	 	16.4	 	39.9	 	
Inferred	 	9.8	 	18.4	 	47.1	 	
 
16.3.15  KONTTIJÄRVI MINERAL RESOURCE ESTIMATE
 
An optimized pit design was produced by AMC Consultants for the Snowden 2004 resource (Lycopodium, 2005). Because the deposit was uneconomic at the time, no reserve meeting NI43-101 standards could be declared. A “Pit Design Inventory” was listed for the optimized pit however, and is tabulated in Table 16.17. The “Pit Design Inventory” is based on a selective mining “recoverable resource” model, and incorporates an additional 2.5% dilution at zero grade to account for edge effects (AMC, 2003).
 
Table 16.17:  GFAP Feasibility Study Pit Design Resource Inventory

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	2PGE+Au	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
Snowden	 	MEAS	 	17.5	 	1.79	 	0.51	 	0.13	 	0.08	 	0.17	 	2.43	 	1.4	 	
2004(1)	 	IND	 	3.8	 	1.78	 	0.50	 	0.13	 	0.07	 	0.19	 	2.41	 	0.3	 	
5 x 5 x 4	 	INF	 	1.6	 	1.50	 	0.40	 	0.11	 	0.03	 	0.17	 	2.01	 	0.1	 	

(1) 1.0 g/t 2PGE+Au cutoff
 
The AMC pit design was applied to the NAP 2006 resource model for comparative purposes only, and the results are tabulated for a 2PGE+Au resource in Table 16.18 and a PdEq resource in Table 16.19. No dilution factors were applied to the NAP model.
 
Table 16.18: NAP 2006 Pit Design 2PGE+Au Resource Inventory.

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	2PGE+Au	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
OK(1)	 	MEAS	 	12.426	 	1.678	 	0.481	 	0.124	 	0.075	 	0.145	 	2.283	 	0.912	 	
15 x 15 x 8	 	IND	 	10.054	 	1.499	 	0.423	 	0.120	 	0.060	 	0.149	 	2.043	 	0.660	 	
 	 	INF	 	1.337	 	1.649	 	0.481	 	0.110	 	0.045	 	0.153	 	2.239	 	0.096	 	

(1) 1.0 g/t 2PGE+Au cutoff     NAP standard is to report grades to 3 decimal places.
 
78

 
Table 16.19: NAP 2006 pit design PdEq resource inventory.

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	PdEq	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
OK(1)	 	MEAS	 	21.877	 	1.113	 	0.323	 	0.086	 	0.062	 	0.102	 	4.378	 	3.079	 	
15 x 15 x 8	 	IND	 	20.224	 	0.942	 	0.267	 	0.078	 	0.047	 	0.102	 	3.675	 	2.390	 	
 	 	INF	 	2.935	 	1.001	 	0.284	 	0.071	 	0.033	 	0.112	 	3.536	 	0.334	 	

(1) 1.0 g/t PdEq cutoff                             NAP standard is to report grades to 3 decimal places.
 
16.3.16    KONTTIJÄRVI CONCLUSIONS
 
This report represents an exploratory, global resource model of the complex Konttijärvi deposit. In comparison with the Snowden 2004 OK resource model used for the 2005 Feasibility Study, the estimation strategy used produced a resource containing 12% less tonnes, with a 2PGE+Au grade 2% lower (Table 16.20).
 
Table 16.20: Comparison of Measured and Indicated 2PGE+Au resources (1.0 g/t cut-off)

 	 	SNOWDEN 2004	 	 	 	 	 	
 	 	“Recoverable	 	SNOWDEN 2004	 	 	 	
 	 	Resource”	 	OK	 	NAP 2006	 	
ELEMENT	 	5 x 5 x 4	 	25 x 25 x 4	 	15 X 15 X 8	 	
2PGE+Au (M t)	 	31.0	 	34.4	 	30.706	 	
2PGE+Au (g/t)	 	2.40	 	2.20	 	2.154	 	
2PGE+Au (M oz)	 	2.31	 	2.44	 	2.126	 	
Pd (g/t)	 	1.73	 	1.62	 	1.585	 	
Pt (g/t)	 	0.49	 	0.46	 	0.451	 	
Au (g/t)	 	0.13	 	0.11	 	0.118	 	
Ni (%)	 	0.07	 	0.08	 	0.065	 	
Cu (%)	 	0.17	 	0.16	 	0.143	 	
Volume (M m(3))	 	2109	 	2109	 	2162	 	
 
The model used serves to validate the previous GFAP resource estimates. Minor differences in the estimates can be attributed to changes in the support size, the geological model, block model layout and volumes, updated diamond drilling results, and differences in density modelling. Some differences in results between the NAP and Snowden estimates can also be attributed to the estimation approach taken. For the NAP estimate, in-situ resources were estimated using a conservative short-range estimation strategy. The Snowden models made use of empirical mineralization envelopes, which constrain the estimation process, while the NAP model was allowed to continue on all blocks within the Pd search ellipse. The Snowden “recoverable resource” estimate was also modified by the use of uniform conditioning to reduce conditional bias.
 
It must be kept in mind that the Konttijärvi deposit is an unique deposit in a geologically complex environment for which no production experience or knowledge base is available. Until a sufficient amount of mining has taken place to permit reconciliation of the resource model with actual production, the validity of any model cannot be proved. It should also be noted that a definitive mineralization model has not yet been established for the Konttijärvi deposit. For this reason NAP estimation strategy emphasized the conservative use of restrictive search ellipses combined with kriging efficiency for classification.
 
It is however possible to validate the resource models against standard industry practice. Sufficient information is available regarding the Konttijärvi deposit to make
 
79

 
preliminary financial decisions, and the approach taken in terms of diamond drilling and sampling is suitable for resource estimation applied to open-pit resources at the scoping level.
 
16.4                      AHMAVAARA PROPERTY
 
16.4.1               DATA SUPPLIED
 
GFAP geologists supplied Triangulated Irregular Network (TIN) files of the local topography and the lower till contact. Files representing the lithological contacts, as well as TIN solids of the mineralization domains used for the 2005 Feasibility Study were also supplied. TIN files are used by the Gemcom mine modelling software program to generate three-dimensional digital models of surfaces (such as topography) and volumes (such as a rock unit).
 
All data are maintained by GFAP in an Access database. GFAP supplied an extract file from the database which contained 1,179 drillholes, of which 506 diamond drillholes were used for grade modeling (Table 16.21). The extract contains header, survey, lithology and assay. Assay data reporting Pd, Pt, Au, Ni and Cu sample grades were extracted from the database for modeling. The results of validation checks of the extract were within industry standards.
 
Table 16.21: Drilling history

YEAR	 	PROGRAM	 	COUNT	 	MODEL	 	
1995	 	OUTOKUMPU	 	28	 	28	 	
2000	 	APP RESOURCE DEFINITION	 	17	 	17	 	
2001	 	APP GEOTECH	 	5	 	5	 	
2001	 	APP RESOURCE DEFINITION	 	138	 	138	 	
2001	 	PERCUSSION	 	14	 	0	 	
2001	 	WATER	 	2	 	0	 	
2002	 	APP METALLURGICAL	 	2	 	2	 	
2002	 	APP RESOURCE DEFINITION	 	93	 	93	 	
2003	 	APP GRADE CONTROL	 	20	 	20	 	
2003	 	APP INFILL	 	22	 	22	 	
2004	 	APP BLASTHOLE	 	124	 	0	 	
2004	 	APP GRADE CONTROL	 	36	 	36	 	
2004	 	APP INCLINED	 	11	 	11	 	
2004	 	APP INFILL	 	88	 	88	 	
2004	 	APP PIT	 	4	 	0	 	
2005	 	APP INCLINED	 	46	 	46	 	
2005	 	MINIDRILL	 	529	 	0	 	
 	 	TOTAL	 	1179	 	506	 	
 
In order to reduce the random effects of low grade assay values below the assay detection limits for the defined grade elements, a low grade transform was applied to data extracted from the database, approximately equivalent to half the lower detection limit. For Pd, Pt and Au, assay values less than 0.01 g/t were transformed to 0.005 g/t; Ni and Cu assay values less than 0.005% were transformed to 0.0025%. Blank intervals were flagged as “Not Sampled” in the database.
 
16.4.2               GEOLOGICAL MODEL
 
The geological model used for the 2005 feasibility study was checked against subsequent drilling and accepted for modeling purposes. The Hanging Wall 2 and Hanging Wall 1 units were combined into a single lithological unit, and the Transitional Zone was combined with the Marginal Lower Gabbro (Table 16.22).
 
80

 
16.4.3     MINERALIZATION MODEL
 
The mineralization zones defined for the 2005 feasibility study were checked against subsequent drilling and accepted for modeling purposes (Table 16.23). Five mineralization zones have been previously identified (Figure 16-7).
 
Figure 16-7: Ahmavaara Zones of Mineralization (not to scale)

 
Table 16.22: Modeling lithology .

Lithology	 	Strat Code	 	Rock Code	 	Color	 	
Basement Xenolith	 	XENO	 	1661	 	Lt Blue	 	
Upper Peridotite	 	UP	 	1261	 	Lt Green	 	
Hanging Wall 2 + Hanging Wall 1	 	HW	 	1151	 	Blue	 	
Peridotite Marker	 	PM	 	1251	 	Cyan	 	
Ahmavaara Pyroxenite	 	APX	 	1351	 	Green	 	
Marginal Upper Gabbro	 	MU	 	1451	 	Lt Red	 	
Marginal Central Gabbro	 	MC	 	1461	 	Yellow	 	
Marginal Lower Gabbro + Transitional Zone	 	ML	 	1471	 	Brown	 	
Basement	 	BAS	 	1651	 	Magenta	 	
 
Table 16.23: Mineralization codes.

Mineralization Zone	 	Zone Code	 	Min Code	 	Color	 	
Upper High Grade	 	UHG	 	2201	 	Red	 	
Upper Low Grade	 	ULG	 	2311	 	Green	 	
Central Barren Zone	 	CBZ	 	2301	 	Blue	 	
Lower Low Grade	 	LLG	 	2321	 	Lt Green	 	
Lower High Grade	 	LHG	 	2401	 	Lt Red	 	
 
16.4.4     GRADE ELEMENTS
 
Palladium (Pd), platinum (Pt), gold (Au), nickel (Ni) and copper (Cu) grades were estimated separately as individual grade elements. A derived 2PGE+Au (Pd + Pt + Au) grade parameter was then calculated by combining the appropriate grade elements for each block. A Palladium Equivalent (PdEq) parameter incorporating Pd, Pt, Au, Ni and Cu grade values was calculated as well.
 
81

 
16.4.5     COMPOSITING
 
Raw sample lengths from the Ahmavaara database ranged from 0.04 m to 21.38 m, with an average sample length of 0.97m. A total of 73% of the samples were one meter in length. For estimation purposes the assays were composited to one-meter length-weighted intervals downhole from the collar. Compositing values were checked against length-weighted accumulation values for each grade element (Table 16.24).
 
Table 16.24: Average composite and accumulation values.

Average Grades	 	Pd (g/t)	 	Pt (g/t)	 	Au (g/t)	 	Ni (%)	 	Cu (%)	 	2PGE+Au (g/t)	 	
1m Composites	 	0.643	 	0.137	 	0.078	 	0.061	 	0.141	 	0.857	 	
Accumulation	 	0.648	 	0.148	 	0.078	 	0.063	 	0.144	 	0.874	 	
 
16.4.6     SPECIFIC GRAVITY
 
Specific gravity values for each lithological unit were taken from Snowden (2004a) as given in Table 16.25.
 
Table 16.25: Specific Gravity Values (after Snowden, 2004).

Unit	 	Rock Code	 	Density (t/m 3 )	 	
Basement Xenolith	 	XENO	 	2.89	 	
Upper Peridotite	 	UP	 	2.95	 	
Hanging Wall	 	HW	 	2.91	 	
Peridotite Marker	 	PM	 	2.92	 	
Ahmavaara Pyroxenite	 	APX	 	2.95	 	
Marginal Upper	 	MU	 	2.93	 	
Marginal Central	 	MC	 	2.96	 	
Marginal Lower	 	ML	 	3.02	 	
Basement	 	BAS	 	2.81	 	
 
16.4.7     SUMMARY STATISTICS
 
Composite grade statistics were compiled for each mineralization zone (Table 16.26). Samples were not de-clustered prior to analysis. Low Coefficient of Variation (COV) values and sensitivity studies carried out on grade-capping and high grade outliers suggested that there is no requirement for grade-capping at this level and none was applied.
 
Probability graphs for each mineralization zone, as shown below, indicate the presence of slightly mixed distributions, suggesting that the defined mineralization zones do not represent totally homogenous sample populations.
 
A visual examination of composite grades also suggests that the defined mineralization zones can be further sub-divided into a high grade western sub-zone and a lower grade eastern sub-zone, with the separation associated with a change in the regional strike of the deposit, which is in agreement with the Snowden’s (2004a) findings. This adds greatly to the complexity of an already complex model, and can have significant impacts on local resource calculations in terms of non-stationarity. For the purposes of this report however, global resource estimates have been carried out within each mineralization zone as a whole, and have not been further subdivided.
 
A strong correlation exists between composite Pd and Pt assay grades across all mineralization zones (Table 16.27). The overall Pd:Pt ratio for the deposit is ~1:6. A weaker correlation was also exhibited between Pd and Ni. No correlation could be demonstrated between Au and the other grade elements.
 
82

 
Figure 16-8:  Ahmavaara Grade Probability Graphs

 
83

 

84

 

85

 

86

 
16.4.8     VARIOGRAPHY
 
Experimental semi-variograms were constructed for each grade element within each mineralization zone, as shown below. Horizontal log-semi-variograms were constructed in order to determine the nominal orientation of maximum grade continuity; plunging semi-variograms within this direction were then examined. The CBZ mineralization zone was found to be essentially isotropic for relatively constrained search ranges, while the UHG, ULG, LLG and LHG mineralization zones displayed a marked anisotropy, plunging -30° → 315°. Estimation semi-variograms are tabulated in Table 16.28
 
Table 16.26: Mineralization zone composite grade statistics

UHG	 	AU	 	CU	 	NI	 	PD	 	PT	 	
Count	 	4857	 	4857	 	4857	 	4857	 	4857	 	
Minimum	 	0.005	 	0.0025	 	0.0025	 	0.005	 	0.005	 	
Maximum	 	3.88	 	5.95	 	2.49	 	31.34	 	13.19	 	
Mean	 	0.1454	 	0.173	 	0.08855	 	1.427	 	0.37139	 	
Median	 	0.1	 	0.14	 	0.07	 	1.17	 	0.28	 	
SD	 	0.1986	 	0.1881	 	0.08473	 	1.2634	 	0.37862	 	
COV	 	1.36531	 	1.08732	 	0.95689	 	0.88533	 	1.01945	 	

ULG	 	AU	 	CU	 	NI	 	PD	 	PT	 	
Count	 	4362	 	4362	 	4362	 	4362	 	4362	 	
Minimum	 	0.005	 	0.0025	 	0.0025	 	0.005	 	0.005	 	
Maximum	 	4.13	 	8.19	 	2.33	 	16.3	 	4.15	 	
Mean	 	0.09127	 	0.1947	 	0.06566	 	0.88215	 	0.1875	 	
Median	 	0.06	 	0.15	 	0.05	 	0.67	 	0.14	 	
SD	 	0.1353	 	0.2763	 	0.09465	 	0.89217	 	0.1875	 	
COV	 	1.4828	 	1.41945	 	1.44152	 	1.01136	 	0.9998	 	

LLG	 	AU	 	CU	 	NI	 	PD	 	PT	 	
Count	 	2846	 	2846	 	2846	 	2846	 	2846	 	
Minimum	 	0.005	 	0.0025	 	0.0025	 	0.005	 	0.005	 	
Maximum	 	2.25	 	5.98	 	2.51	 	11.95	 	2.19	 	
Mean	 	0.08508	 	0.1961	 	0.06127	 	0.67942	 	0.1225	 	
Median	 	0.06	 	0.16	 	0.04	 	0.5	 	0.08	 	
SD	 	0.09851	 	0.2222	 	0.08309	 	0.7046	 	0.1312	 	
COV	 	1.15787	 	1.13322	 	1.35599	 	1.03705	 	1.0714	 	

LHG	 	AU	 	CU	 	NI	 	PD	 	PT	 	
Count	 	4903	 	4903	 	4903	 	4903	 	4903	 	
Minimum	 	0.005	 	0.0025	 	0.0025	 	0.005	 	0.005	 	
Maximum	 	19.41	 	12.47	 	3.16	 	69.54	 	7.42	 	
Mean	 	0.17593	 	0.3701	 	0.1466	 	1.3589	 	0.2237	 	
Median	 	0.12	 	0.27	 	0.08	 	0.9	 	0.15	 	
SD	 	0.46472	 	0.48845	 	0.2438	 	2.1503	 	0.3276	 	
COV	 	2.64151	 	1.3198	 	1.66352	 	1.58239	 	1.46448	 	
 
87

 
Table 16.27: Grade element correlation and Pd ratios (1m composites)

 	 	 	 	Pd:Pt	 	Pd:Au	 	Pd:Ni%	 	Pd:Cu%	 	Ni%:Cu%	 	
ALL	 	R 2	 	0.75	 	0.12	 	0.55	 	0.31	 	0.31	 	
	 	Pd	 	5.6	 	15.4	 	12.9	 	7.9	 	NA	 	
UHG	 	R 2	 	0.86	 	0.09	 	0.26	 	0.22	 	0.29	 	
	 	Pd	 	3.6	 	21.1	 	29.3	 	14.1	 	NA	 	
ULG	 	R 2	 	0.82	 	0.26	 	0.67	 	0.58	 	0.48	 	
	 	Pd	 	5.3	 	12.9	 	11.5	 	4.22	 	NA	 	
CBZ	 	R 2	 	0.84	 	0.43	 	0.62	 	0.63	 	0.37	 	
	 	Pd	 	5.7	 	11.8	 	10.0	 	4.8	 	NA	 	
LLG	 	R 2	 	0.75	 	0.27	 	0.65	 	0.58	 	0.27	 	
	 	Pd	 	6.2	 	13.8	 	10.5	 	4.2	 	NA	 	
LHG	 	R 2	 	0.78	 	0.07	 	0.60	 	0.25	 	0.23	 	
	 	Pd	 	7.4	 	17.0	 	11.4	 	8.9	 	NA	 	
 
Table 16.28: Ahmavaara estimation semi-variograms.

DOMAIN	 	 	 	Anisotrophy	 	C o %	 	Semi-variogram	 	
UHG
-30° ? 315°	 	Pd	 	1.0	 	0.8	 	0.3	 	38	%	0.6 + SPH(50, 0.8) + SPH(80, 0.2)	 	
	 	Pt	 	1.0	 	1.0	 	0.6	 	7	%	0.01 + SPH(20, 0.11) + SPH(60, 0.023)	 	
	 	Au	 	1.0	 	0.7	 	0.3	 	5	%	0.002 + SPH(20, 0.036) + SPH(30, 0.001)	 	
	 	Ni	 	1.0	 	0.5	 	0.3	 	14	%	0.001 + SPH(30, 0.002) + SPH(199, 0.004)	 	
	 	Cu	 	1.0	 	0.8	 	0.2	 	14	%	0.005 + SPH(50, 0.02) + SPH(99, 0.01)	 	
ULG
-30° ? 315°	 	Pd	 	1.0	 	1.0	 	0.3	 	17	%	0.13 + SPH(70, 0.61) + SPH(99, 0.036)	 	
	 	Pt	 	1.0	 	0.9	 	0.3	 	3	%	0.001 + SPH(50, 0.028) + SPH(99, 0.004)	 	
	 	Au	 	1.0	 	0.8	 	0.4	 	6	%	0.001 + SPH(90, 0.01) + SPH(199, 0.007)	 	
	 	Ni	 	1.0	 	0.4	 	0.1	 	11	%	0.001 + SPH(60, 0.005) + SPH(99, 0.003)	 	
	 	Cu	 	1.0	 	0.5	 	0.1	 	7	%	0.005 + SPH(30, 0.04) + SPH(99, 0.0314)	 	
CBZ
Isotropic	 	Pd	 	1.0	 	1.0	 	1.0	 	46	%	0.12 + SPH(15, 0.14) + SPH(99, 0.0034)	 	
	 	Pt	 	1.0	 	1.0	 	1.0	 	21	%	0.002 + SPH(12, 0.006) + SPH(50, 0.0017)	 	
	 	Au	 	1.0	 	1.0	 	1.0	 	47	%	0.002 + SPH(10, 0.001) + SPH(65, 0.0013)	 	
	 	Ni	 	1.0	 	1.0	 	1.0	 	47	%	0.002 + SPH(9, 0.0021) + SPH(99, 0.0002)	 	
	 	Cu	 	1.0	 	1.0	 	1.0	 	22	%	0.004 + SPH(11, 0.01) + SPH(30, 0.0039)	 	
LLG
-30° ? 315°	 	Pd	 	1.0	 	0.5	 	0.2	 	1	%	0.005 + SPH(30, 0.3) + SPH(50, 0.19)	 	
	 	Pt	 	1.0	 	0.4	 	0.2	 	6	%	0.001 + SPH(35, 0.007) + SPH(50, 0.009)	 	
	 	Au	 	1.0	 	1.0	 	0.2	 	4	%	0.0004 + SPH(90, 0.008) + SPH(99, 0.0011)	 	
	 	Ni	 	1.0	 	0.5	 	0.3	 	1	%	0.0001 + SPH(40, 0.002) + SPH(60, 0.0048)	 	
	 	Cu	 	1.0	 	0.7	 	0.2	 	10	%	0.005 + SPH(35, 0.034) + SPH(70, 0.01)	 	
LHG
-30° ? 315°	 	Pd	 	1.0	 	0.7	 	0.3	 	44	%	0.7 + SPH(50, 0.7) + SPH(99, 0.2)	 	
	 	Pt	 	1.0	 	0.5	 	0.3	 	7	%	0.007 + SPH(40, 0.08) + SPH(60, 0.02)	 	
	 	Au	 	1.0	 	0.7	 	0.4	 	28	%	0.06 + SPH(40, 0.1) + SPH(99, 0.056)	 	
	 	Ni	 	1.0	 	0.6	 	0.4	 	37	%	0.022 + SPH(20, 0.035) + SPH(99, 0.0024)	 	
	 	Cu	 	1.0	 	0.7	 	0.2	 	42	%	0.1 + SPH(10, 0.11) + SPH(80, 0.029)	 	
 
88

 
Figure 16-9:  Ahmavaara Experimental Semivariogram Models

 
89

 

90

 

16.4.9     BLOCK MODEL
 
The drilling grid used at Ahmavaara is nominally at a 50 m x 50 m spacing, with some overlap due to a change in orientation of the deposit. A 30 m x 30 m x 8 m (L x W x H) estimation block size was selected as a compromise between (potential) mining selectivity, the narrow width of the mineralized zones, the existing drilling grid and semi-variogram ranges. The block model framework contains 160,000 blocks covering the defined area of mineralization (Table 16.29).
 
91

 
Table 16.29: NAP 2006 block model

 	 	Min	 	Max	 	Block Size	 	Count Blocks	 	
Easting	 	29820	 	32220	 	30	 	80	 	
Northing	 	10000	 	11500	 	30	 	50	 	
Elevation	 	168	 	-152	 	8	 	40	 	
Volume = 3072 M m 3	 	 	 	 	 	 	 	
 
Block model mineralization zone parameters were constructed from a percent Gemcom™ (version 5.45) model, allowing for a high degree of accuracy in terms of volume estimates. Bulk specific gravity values were defined by the appropriate lithological unit and assigned based on the block centroid.
 
16.4.10                  CUT-OFF GRADES
 
Due to the polymetallic nature of the deposit the resource inventory was tabulated relative to a 2PGE+Au cut-off for comparative purposes with the Snowden estimates. The resource inventory listing is the arithmetic accumulation of individual grade element estimates from blocks that meet the 2PGE+Au cut-off value.
 
Current NAP operating practice is to also express resources in terms of a NSR cutoff. A resource inventory has therefore also been calculated relative to a NSR palladium equivalent (PdEq). PdEq grades were calculated from individual grade element values as follows:
 
PdEq = Pd + (3.49 x Pt) + (1.96 x Au) + (19.21 x Ni) + (5.56 x Cu)
 
PdEq parameters were derived from the data given in Table 16.30.
 
Table 16.30: PdEq NSR parameters* - Ahmavaara.

Element	 	Metal Price
$ US/lb or oz	 	Concentrate
Recovery	 	Net 
Revenue %	 	Pd Eq
Factor	 	Payable
Metal
$US/tonne/%
or gram	 	
Pd	 	277.00	 	75	%	22	%	1.00	 	5.71	 	
Pt	 	1057.00	 	67	%	16	%	3.49	 	19.91	 	
Au	 	546.00	 	77	%	5	%	2.09	 	11.19	 	
Ni	 	9.04	 	65	%	33	%	19.23	 	109.72	 	
Cu	 	2.56	 	75	%	23	%	5.56	 	31.76	 	

* Confidential smelter terms: Smelter recovery/payables, treatment and refining charges, penalties, freight costs, etc., used to determine NSR/net revenues per unit are not shown because of their proprietary nature.
 
16.4.11                  ESTIMATION
 
The resource estimate was based on 506 diamond drillholes. Estimates were made from 3D block models using commercial Gemcom™ software, version 5.45.
 
Separate in-situ global resource estimates were produced for the defined grade elements for 30 m x 30 m x 8 m blocks using Ordinary Kriging on composite grades. Estimation data were constrained within demarcated mineralization zones. No dilution factor was applied to the estimates, as the mining method has not been selected and the selective mining unit size is indeterminate.
 
The kriging neighbourhood was examined for the Pd one-meter composite data. For the CBZ zone, a horizontal search ellipse with semi-axes of 80 m x 80 m x 20 m was selected. For the anisotropic zones, a rotated search ellipse with semi-axes of 80 m x 40 m x 20 m was used, with the major semi-axis plunging -30° at Azimuth 315°. Block discretization was set at 4 x 4 x 4, and a minimum of five samples and a maximum of twenty samples were used for block estimation.
 
92

 
Estimation results were further divided into a high-grade western resource zone and lower-grade eastern resource zone. The demarcation line separating the western and eastern resource zones was selected after consultation with the GFAP geologists, and corresponds to a change in the regional strike of the Ahmavaara deposit (Figure 16-6). The zones selected do NOT correspond with the East and West Ahmavaara resources reported by Snowden (2004a), which appear to have been based on different criteria.
 
Figure 16-10: Eastern & western resource zones (Source: F.H. Brown 2006).

 
NAP 2006 Global Resource Inventory – Ahmavaara Deposit
 
The Global Resource Inventory for the Ahmavaara deposit, based on a NSR palladium equivalent estimated at a cut-off of 1.0 g/t PdEq produced a Measured + Indicated total of 153.013 million tonnes with an average PdEq grade of 3.877 g/t and an Inferred total of 34.757 million tonnes at an average PdEq grade of 4.015 g/t
 
The 2006 Global Resource Inventory for the Konttijärvi and Ahmavaara Deposits (1.0 g/t PdEq cut-off) as prepared by F.H. Brown (2006) is shown in Table 16.11 and for convenience the Ahmavaara data are separately summarized in Table 16.31 below.
 
Table 16.31: NAP 2006 PdEq resource inventory – Ahmavaara Deposit

 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	PdEq	 	
CAT	 	M t	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	38.197	 	0.707	 	0.141	 	0.086	 	0.061	 	0.161	 	3.425	 	4.206	 	
IND	 	114.816	 	0.851	 	0.175	 	0.104	 	0.071	 	0.181	 	4.027	 	14.867	 	
INF	 	34.757	 	0.867	 	0.187	 	0.103	 	0.070	 	0.170	 	4.015	 	4.486	 	
1.0 g/t PdEq cutoff	 	NAP standard is to report grades to 3 decimal places.	 	
 
The distribution of mineralization within the lithological units was also examined (Table 16.32), as this directly affects the metallurgical characteristics of the mineralization. A complete resource inventory by cut-off grade for 2PGE+Au and PdEq estimates is given in Appendix 1.
 
93

 
Table 16.32: Lithological unit mineralization – 1.0 g/t 2PGE+Au cutoff (Measured + Indicated + Inferred)

 	 	 	 	PM	 	APX	 	MU	 	MC	 	ML	 	BAS	 	
M Tonnes	 	 	 	4	 	12	 	26	 	20	 	18	 	12	 	
% Mineralization	 	Pd	 	5	%	14	%	27	%	20	%	20	%	14	%	
	 	Pt	 	5	%	17	%	30	%	19	%	17	%	12	%	
	 	Au	 	4	%	12	%	24	%	20	%	24	%	17	%	
	 	Ni	 	3	%	10	%	20	%	21	%	27	%	18	%	
	 	Cu	 	3	%	9	%	21	%	235	%	28	%	17	%	
	 	2PGE+Au	 	5	%	14	%	27	%	20	%	20	%	14	%	
Avg Grade	 	Pd g/t	 	1.301	 	1.325	 	1.173	 	1.135	 	1.263	 	1.306	 	
	 	Pt g/t	 	0.289	 	0.318	 	0.260	 	0.213	 	0.208	 	0.216	 	
	 	Au g/t	 	0.118	 	0.134	 	0.121	 	0.134	 	0.178	 	0.183	 	
	 	Ni %	 	0.076	 	0.076	 	0.069	 	0.096	 	0.133	 	0.134	 	
	 	Cu %	 	0.150	 	0.166	 	0.185	 	0.260	 	0.351	 	0.329	 	
	 	2PGE+Au	 	1.707	 	1.777	 	1.554	 	1.481	 	1.649	 	1.705	 	
 
16.4.12                  CLASSIFICATION
 
Global mineral resources were classified in accordance with guidelines established by the Canadian Institute of Mining, Metallurgy and Petroleum (CIM, 2005) and outlined in Section 16.1.13 of this report.
 
Kriging efficiency (Krige, 1996) and the geological continuity of the Pd estimate were used as the primary criteria for classification. Kriging efficiency was calculated for each Pd block estimate using the OK estimation variance and the block variance. Kriging efficiency (KE) and the estimation regression slope (R) were then used to establish the relative confidence of the Pd block estimates, based on the semivariogram parameters and the location of the informing data. Higher kriging efficiencies are associated with better informed blocks, with a value of 100% reflecting perfect theoretical estimation.
 
The following ranges were then used to assign an overall classification level to each block:
 
•                                           Measured: KE >= 75% and R >= 0.8
•                                           Indicated: 30% >= KE < 75% and R >= 0.6
•                                           Inferred: 0% > KE < 30%
 
Well-informed blocks, with a low conditional bias, were therefore classified as Measured. Moderately-informed blocks were classified as Indicated. Poorly-informed blocks with less than adequate sample information or a poor spatial distribution of sample information were classified as Inferred. Blocks with a zero or negative KE were deleted from the estimation inventory (Figure 16-7). Only Pd results were used to classify the resource, as Pd is the primary grade element of economic interest.
 
16.4.13                  MODEL VALIDATION
 
Five methods were used to test the validity of the OK model:
 
•                                           Visual inspection of rock codes, composite grades and block grades in multiple north-south sections. No systematic bias was evident in the model estimates.
 
•                                           Global comparison of average composite grades per block and model block grades for classified blocks (Table 16.33). Results for all grade elements indicate that the model provides a reasonable smoothed global estimate.
 
94

 
•                                           Comparison of model block estimates with a Nearest Neighbour estimate. The average grade of the model estimate and the NN estimate were compared at a nominal cutoff grade of 0.01. The mean grades were within 12% for all grade elements, indicating that the model retains a minor conditional bias, primarily due to the mixing of high-grade and low-grade populations and the wider drillhole spacing (Table 16.34). Results are sufficient for global estimates.
 
•                                           OK block estimates, NN estimates and composite block averages were also compared to average blasthole grades from the 2004 trial mine sampling program (GFAP, 2004). The two blocks affected straddle the UHG and ULG contact. The sample size is too small to make any significant conclusion, but the results do demonstrate the impact of change-of-support on the resource estimate (Table 16.35).
 
•                                           As a further test of the reasonableness of the model, the average informing distance of the closest sample used for estimation of the Pd grade element was also examined (Table 16.36). The distances used are consistent with industry practice for classification.
 
Figure 16-11: Block classification (Section 30855, looking east)
(Source: F.H. Brown 2006) .

 
Table 16.33: Comparison of composite averages and model estimates

Grade Element	 	Composite Avg	 	Model Estimate	 	
Pd	 	 	0.815	 	0.752	 	
Pt	 	 	0.169	 	0.163	 	
Au	 	 	0.096	 	0.095	 	
Ni	 	 	0.070	 	0.068	 	
Cu	 	 	0.168	 	0.168	 	
2PGE+Au	 	 	1.080	 	1.010	 	
 
95

 
Table 16.34: Comparison of OK and NN block model estimates.

Grade Element	 	OK	 	NN	 	VAR	 	
Pd	 	0.769	 	0.691	 	-11.3	%	
Pt	 	0.159	 	0.145	 	-9.6	%	
Au	 	0.093	 	0.085	 	-9.4	%	
Ni	 	0.064	 	0.061	 	-4.9	%	
Cu	 	0.164	 	0.152	 	-7.9	%	
 
Table 16.35: Comparison of blasthole averages and NAP block model estimates

Grade Element	 	Blasthole Avg	 	Comp. Avg	 	OK	 	NN	 	
Pd	 	1.803	 	1.600	 	0.886	 	0.516	 	
Pt	 	0.333	 	0.306	 	0.133	 	0.104	 	
Au	 	0.245	 	0.163	 	0.099	 	0.140	 	
Ni	 	0.205	 	0.152	 	0.095	 	0.050	 	
Cu	 	0.322	 	0.319	 	0.259	 	0.167	 	
2PGE+Au	 	2.381	 	2.069	 	1.118	 	0.760	 	
 
Table 16.36: Average informing distance of the closest sample for Pd

 	 	MINIMUM	 	AVERAGE	 	MAXIMUM	 	
Measured	 	1.0	 	9.7	 	22.7	 	
Indicated	 	2.5	 	19.7	 	46.0	 	
Inferred	 	8.4	 	31.3	 	61.1	 	
 
16.4.14  AHMAVAARA MINERAL RESERVES ESTIMATES
 
An optimized pit design was produced by AMC Consultants for the Snowden 2004 resource (Lycopodium, 2005) for the Ahmavaara deposit. Because the deposit was uneconomic at the time, no reserve meeting NI43-101 standards could be declared. A “Pit Design Inventory” was listed for the optimized pit however, and is tabulated in Table 16.37. The “Pit Design Inventory” incorporates an additional 2.5% dilution at zero grade to account for edge effects (AMC, 2003), as well as 5% dilution included in the Uniform Conditioning “recoverable resource” model.
 
Table 16.37: GFAP Feasibility study pit design constrained resource inventory

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	2PGE+Au	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
Snowden 2004
5 x 5 x 4	 	MEAS	 	19.1	 	1.52	 	0.33	 	0.15	 	0.13	 	0.27	 	2.01	 	1.23	 	
	 	IND	 	21.6	 	1.19	 	0.26	 	0.14	 	0.09	 	0.24	 	1.58	 	1.10	 	
	 	INF	 	NA	(1)	 	 	 	 	 	 	 	 	 	 	 	 	 	 	

(1) 1.0 g/t 2PGE+Au cutoff	(1) Inferred resources not in production schedule.	 	
 
The AMC pit design was applied to the NAP 2006 Ahmavaara resource model for comparative purposes only, and the results are shown for a 2PGE+Au resource in Table 16.38 and a PdEq resource in Table 16.39. No dilution factors were applied to the NAP model, and the resource inventory represents the total tonnage contained within the pit design.
 
96

 
Table 16.38: NAP 2006 pit design 2PGE+Au resource inventory

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	2PGE+Au	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
OK (1)
30 x 30 x 8	 	MEAS	 	10.105	 	1.355	 	0.274	 	0.156	 	0.113	 	0.270	 	1.785	 	0.580	 	
	 	IND	 	32.473	 	1.404	 	0.307	 	0.162	 	0.107	 	0.240	 	1.873	 	1.955	 	
	 	INF	 	5.319	 	1.321	 	0.312	 	0.140	 	0.091	 	0.191	 	1.774	 	0.303	 	

(1) 1.0 g/t 2PGE+Au cutoff                   NAP standard is to report grades to 3 decimal places.
 
Table 16.39: NAP 2006 pit design PdEq resource inventory

 	 	 	 	Tonnes	 	Pd	 	Pt	 	Au	 	Ni	 	Cu	 	PdEq	 	
SOURCE	 	CAT	 	Mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
OK (1)
30 x 30 x 8	 	MEAS	 	25.198	 	0.796	 	0.160	 	0.094	 	0.068	 	0.172	 	3.806	 	3.083	 	
	 	IND	 	50.637	 	1.068	 	0.231	 	0.125	 	0.083	 	0.194	 	4.795	 	7.806	 	
	 	INF	 	9.191	 	0.961	 	0.222	 	0.105	 	0.069	 	0.154	 	4.129	 	1.220	 	

(1) 1.0 g/t PdEq cutoff                           NAP standard is to report grades to 3 decimal places.
 
16.4.15  CONCLUSIONS
 
This report represents an exploratory, global resource model of the complex Ahmavaara deposit. In comparison with the Snowden (2004a) resource inventory (50 m x 50 m x 4 m blocks), the estimation strategy used produced a resource containing 8% more Measured + Indicated tonnes, with a 2PGE+Au grade 3% higher (Table 16.40).
 
The 2005 Feasibility Study utilized a “recoverable resources” model, based on a 5 m x 5 m x 4 m selective mining unit, derived from the 50 m x 50 m x 4 m large panel OK estimate. The NAP estimate reflects the impact of additional diamond drilling results and an expanded block model, as well as a more conservative approach in estimation. The model therefore serves to validate the previous GFAP resource estimates, although additional work will be required to quantify grade trends and sub-zone based variography on the resource estimate, including in-fill drilling to reduce the conditional bias.
 
It must be kept in mind that the Ahmavaara and Konttijärvi deposits are unique deposits in a geologically complex environment for which no production experience or knowledge base is available. Until a sufficient amount of mining has taken place to permit reconciliation of the resource model with actual production, the validity of any model cannot be proved. It should also be noted that a definitive mineralization model has not yet been established. For this reason NAP estimation strategy has emphasized the conservative use of restrictive search ellipses and kriging efficiency for classification.
 
It is however possible to validate the resource models against standard industry practice. Sufficient information is available regarding the Ahmavaara deposit to make preliminary financial decisions, and the approach taken in terms of diamond drilling and sampling is suitable for resource estimation applied to open-pit resources.
 
97

 
Table 16.40: Comparison of Measured and Indicated 2PGE+Au resources (1.0 g/t cut-off)

 	 	SNOWDEN 2004	 	SNOWDEN 2004	 	NAP 2006	 	
ELEMENT	 	“Recoverable Resource” 5 x 5 x 4	 	large panel 50 x 50 x 4	 	30 X 30 X 8	 	
2PGE+Au (M t)	 	54.7	 	61.8	 	67.395	 	
2PGE+Au (g/t)	 	1.89	 	1.69	 	1.739	 	
2PGE+Au (M oz)	 	3.32	 	3.37	 	3.768	 	
Pd (g/t)	 	1.43	 	1.28	 	1.308	 	
Pt (g/t)	 	0.30	 	0.28	 	0.274	 	
Au (g/t)	 	0.16	 	0.14	 	0.156	 	
Ni (%)	 	0.10	 	0.08	 	0.104	 	
Cu (%)	 	0.28	 	0.25	 	0.250	 	
Volume (M m 3 )	 	2120	 	2120	 	3072	 	
 
98

 
17.0        OTHER RELEVANT DATA AND INFORMATION
 
There are no other data considered relevant to this report that have not been included.
 
99

 
18.0        INTERPRETATION AND CONCLUSIONS
 
Based on the substantial amount of data generated by the extensive technical and metallurgical work programs conducted on both the Konttijärvi and Ahmavaara deposits the following interpretations and conclusions can be reached.
 
•                   The current NAP resource estimates contained in this report are intermediate between the Snowden “recoverable resource” model and the Snowden OK block model, reflecting differences that can be attributed to changes in the geological model, block model layout and volumes, updated diamond drilling results, different density models and the estimation approach taken.
 
•                   For the NAP Konttijärvi estimate, in-situ resources were estimated using a conservative short range estimation strategy to minimize conditional bias. The Snowden models made use of empirical mineralization envelopes, which constrain the estimation process, while the NAP model was allowed to continue on all blocks within the Pd search ellipse.
 
•                   The Suhanko deposit is a Greenfield project – it is a unique deposit that is mineralogically complex for which no production experience or knowledge base is available. Until a sufficient amount of mining has taken place to permit reconciliation of the resource model with actual production, the validity of any model or estimate cannot be proved.
 
•                   It should also be noted that a definitive mineralization model for the Ahmavaara deposit has not yet been established. For this reason the NAP estimation strategy emphasized the conservative use of restrictive search ellipses and kriging efficiency for classification.
 
•                   Sufficient information is available regarding the Suhanko deposits to make preliminary financial decisions, and the approach taken in terms of diamond drilling and sampling is suitable for resource estimation applied to open-pit resources.
 
•                   Sufficient drilling information is presently available to justify undertaking preliminary financial analysis of the project based on the current resource estimates for both the Konttijärvi and Ahmavaara deposits
 
•                   The most recent studies essentially verified previous work by Snowden used in the 2005 Lycopodium Feasibility Study.
 
•                   The current resource estimates prepared for NAP did not use SMU or uniform conditioning but instead were prepared using alternative geological and mineralization domains.
 
•                   There is limited potential to increase known resources in area immediately around Konttijärvi and Ahmavaara, however, the Arctic Platinum Project includes several other mineralized prospects i.e. Narkaus (Nutturalampi, Kuohunki, and Siika Kama), Little Suhanko and Suhanko North, Tuumasuo etc that warrant further investigation
 
•                   The Ahmavaara deposit contains significant Inferred resources that should be upgraded to at least Indicated status
 
Based on the project results to date, it can be concluded that the Suhanko Project represents a potential mining situation of sufficient merit to warrant more advanced, feasibility level engineering studies.
 
100

 
19.0        RECOMMENDATIONS
 
19.1        INTRODUCTION
 
The recommendations regarding the Suhanko Project arising from Section 1 of this report, the “Technical Report and Mineral Resource Estimate”, are outlined below. Further recommendations arising from the “Preliminary Economic Assessment” portion of this report are contained in Section 32.0 titled “Recommendations for Future Work”.
 
19.2        RECOMMENDATIONS
 
P&E feels that the Konttijärvi and Ahmavaara are of sufficient merit and have advanced to the stage where the following recommendations can be made:
 
1                                           It is recommended that a program of infill diamond drilling be conducted in order to further define and upgrade current resource categories’ (i.e. Inferred to Indicated) especially near the upper portions of known mineralized zones.
 
2                                           It is further recommended that GFAP proceed with a preliminary economic analysis (“PEA”) using the current resource estimate.
 
3                                           Upon completion of the recommended detailed in-fill drilling program the project database should be updated and used to revise current resource estimates.
 
4                                           It is recommended that should positive results be gained from the PEA then a bankable feasibility study should be conducted in order to determine if an acceptable financial return can be achieved for the project. Several options are being explored by Aker Kvaerner that are designed to lower operating costs for the processing of Suhanko mineralization and possibly reducing capital investment. The Aker Kvaerner flow sheet option appears to offer some advantages over the strategies that have been presented in past studies and serious consideration should be given to utilizing this process model in any higher level studies.
 
Assuming a decision to proceed with a feasibility level study as recommended above the following is recommended by Aker Kvaerner:
 
5                                           Metallurgical performance is perhaps one of the most important aspects of the success of any project; therefore a metallurgical test program has been developed jointly by Aker Kvaerner and SGS Lakefield. It is recommended that this process be considered for use in a possible feasibility Study in order to optimize the metal recoveries and reagent consumptions. This work would utilize the mineralized samples that were used for the trade-off study testwork. Other mineralized samples have also been shipped to SGS Lakefield and preliminary testing should be designed to help evaluate the possible use of high-pressure grinding rolls as an alternative to SAG milling.
 
6                                           A capacity trade-off study should be carried out for the selected process flow sheet to determine if a larger capacity will increase the financial return for the project.
 
101

 
19.3        PROPOSED BUDGET
 
19.3.1     PROPOSED 2007 EXPLORATION PROGRAM
 
The following work programs and budgets are proposed for 2007

1)             Continued infill at drilling at Ahmavaara	$	 700,000	 	
2)             Additional Trail mining and bulk sampling of Konttijärvi	$	400,000	 	
3)             Pilot plant for Konttijärvi and Ahmavaara at an estimated cost of	$	600,000	 	
4)             Resource modeling, pit optimization and engineering studies	$	400,000	 	
5)             Complete environmental submission for Narkaus project(s)	$	200,000	 	
 	 	 	 	
Infill Drilling	 	 	 	
 	 	 	 	
Diamond Drilling	 	 	 	
5000m @ $130/meter	 	$	650,000	 	
Assays	 	 	 	
500 samples @ $50/sample	 	$	25,000	 	
Miscellaneous field expenses	 	$	25,000	 	
Subtotal DDH	 	$	700,000	 	
 	 	 	 	
Resource Modeling, Pit Optimization and Engineering Studies	 	 	 	
 	 	 	 	
Resource modeling and pit optimization	 	$	200,000	 	
Engineering	 	$	200,000	 	
 	 	 	 	 	
Subtotal Engineering	 	$	400,000	 	
 	 	 	 	 	
Trial Mining/Process Plant	 	 	 	
 	 	 	 	
Trail Mining and Bulk Sample from Konttijärvi	 	$	400,000	 	
Pilot plant for processing material from both Konttijärvi and Ahmavaara.	 	$	600,000	 	
Contingency 10%	 	$	200,000	 	
 	 	 	 	 	
Subtotal Trial Mining and Processing	 	$	1,200,000	 	
 	 	 	 	
TOTAL 2007 BUDGET	 	$	2,300,000	 	

102

 
20.0        REFERENCES
 
Alapieti, T.T. and Huhtelin, T.A. 2005. The Kemi intrusion and associated chromitite deposit; in Early Palaeoproterozoic (2.5–2.4) Tornio–Näränkävaara layered intrusion belt and related chrome and platinum group element mineralization, northern Finland, Geological Survey of Finland, Guide 51a, p.13-32.
 
Alapieti, T.T. 2005. Classification of the Fennoscandian early Palaeoproterozoic (2.5–2.4 Ga) layered intrusions and related PGE deposits; Extended Abstracts, 10th International Platinum Symposium, August 8–11, 2005, Oulu, Finland, p.21-24.
 
Alapieti, T.T. & Lahtinen, J.J. 2002. Platinum-Group Element Mineralization in Layered Intrusions of Northern Finland and the Kola Peninsula, Russia. In, The Geology, Geochemistry, Mineralogy and Mineral Beneficiation of Platinum-Group Elements. L.J. Cabri., (ed.) Canadian Institute of Mining, Metallurgy and Petroleum, Special Volume 54, p. 507-546.
 
APP (2002). Suhanko BFS Volume 3: geology and resource estimation. Unpublished report dated September 2002.
 
Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia: The Joint Ore Reserves Committee (Jorc) 2004:The 2004 Australasian Code for Reporting Exploration Results, Mineral Resources and Ore Reserves (The Jorc Code)
 
BMGS, 2004: A report on the geological process at the Konttijärvi and Ahmavaara trial open pits, Suhanko Project, Finland. Unpublished report to Gold Fields Arctic Platinum Oy by BM Geological Services, 49pp
 
Brown, F.H., 2006a, Summary Review of the Konttijärvi Geochemical Database, Rovaniemi, Finland; Prepared for North American Palladium, 8p.
 
Brown, F.H., 2006b, Summary Review of the Ahmavaara Geochemical Database, Rovaniemi, Finland; Prepared for North American Palladium, 5p.
 
Carter, L.M. and Martin, G.J., 2005, Determination of the Degree of Sulphide Oxidation in Samples from the Arctic Platinum Project; Mineralogical Report No. MIN 1004/115, SGS Lakefield Research Ltd., 13p.
 
Easton, R.M. and Hart, T.R. 2005: 10th International Platinum Symposium: insights relevant to Ni-Cu-PGE mineral exploration in Ontario; in Summary of Field Work and Other Activities 2005, Ontario Geological Survey, Open File Report 6172, p.14-1 to 14-14.
 
Evans, C. and Guerney, P., 2002, Suhanko Project Mineralogy; JKTech-JKMRC Commercial Division, MLA Bureau, MLA Job No. D83, JKTech Job No. 02283, 25p.
 
Evans, D.M., 2000, Report on Geological Interpretations at Ahmavaara and Konttijärvi Projects, Finland; For the Arctic Platinum Partnership, Gold Fields International Inc. and Outokumpu Oy, 32p.
 
Hayward, C. and Robb, L., 2005, Mineralogical and Metal Deportment Study of Cu-Ni-PGE Mineralization at the Konttijärvi and Ahmavaara Deposits. Suhanko Project; A Study Commissioned by Gold Fields Arctic Platinum Oy, Finland, 96p.
 
Iljina, M., 1994, The Portimo Layered Igneous Complex with Emphasis on Diverse Sulphide and Platinum-Group Deposits; Acta Univ. Oul. A 258, 158p.
 
103

 
Iljina, M.J. and Lee, C.A. 2005: PGE deposits in the marginal series of layered intrusions; in Exploration for platinum-group element deposits, Mineralogical Association of Canada, Short Course Volume 35, p.75-96.
 
Liipo, J., 2001, Mineralogy and Geochemistry of the Ahmavaara and Konttijärvi Laboratory Test Feeds; Outokumpu Research Oy, Report # 01042-ORC-T, 31p.
 
Lycopodium (2005). Arctic Platinum Partnership Suhanko Project Feasibility Study. Prepared by Lycopodium Pty., dated May 2005
 
Marion, R., 2006, Final Report for Qemscan Analysis of Drill Core Composite Samples-Prepared for North American Palladium Ltd; Project# CALR 11341-001, SGS Mineral Services, Lakefield Site, 10p.
 
Martin, C., McKenzie, S., 2006, An Investigation into the Recovery of Base and Precious Metals from Suhanko Ores: Prepared for North American Palladium Ltd.; Project No. 11018-003, SGS Lakefield Research Ltd., 27p.
 
McKay, N., 2002, QemSCAN Analysis of Mineral Assemblage and Sulphide Liberation in Four Ground Ore Samples: Submitted by Arctic Platinum; Project No. 8901-233 MI5006-NOV01, Lakefield Research Limited, 24p.
 
Mungall, J.E. 2005. Magmatic geochemistry of the platinum-group elements; in Exploration for platinum-group element deposits, Mineralogical Association of Canada, Short Course Volume 35, p.1-34.
 
Naumanen, K., 2005, Sulphide Oxidation Program-Ahmavaara and Konttijärvi: October 2005-February 2005; Prepared for Goldfields Arctic Platinum Oy, 40p.
 
Richards, J.M. and Martin, G.J., 2001, The Mineralogical Examination of Four Samples from Ahmavaara and Konttijärvi Sections, Arctic Platinum Partnership, Finland; with Emphasis on PGE and Gold Deportment; Mineralogical Report: No. MIN 1001/132, Lakefield Research Africa (Ptyl Limited), 17p.
 
Rickard, J.H., 2004, Report on the Platinum Group Mineralogy of All Ore Types from the Lac des Îles Mine; Prepared for North American Palladium Ltd., 26p.
 
Rickard, J.H. 2006a, Mineralogical and Geochemical Study of the Konttijärvi and Ahmavaara Deposits, Arctic Platinum Project, Finland; Prepared for North American Palladium Ltd., 26 p.
 
Rickard, J.H. 2006b, Sulphide Mineralogy of the Konttijärvi and Ahmavaara Deposits, Arctic Platinum Project, Finland; Prepared for North American Palladium Ltd., 8p
 
Robb, L. 2004: Cu-Ni-PGE Mineralization at Konttijarvi: A Contribution to Ore Genesis and Ore Body Modelling. GFAP Internal Report, April 2004
 
Smee and Associates Consulting Limited, February 2003. Best Practices in Mineral Exploration and results of an audit at GTK Laboratories, Rovaniemi, Finland. Unpublished report for APP,  41pp.
 
Snowden, 2004a: Konttijärvi PGE deposit resource estimate. Snowden Mining Industry Consultants, August 2004. Internal Report No. 4729, (Unpublished) 280pp
 
Snowden, 2004b: Ahmavaara PGE Deposit Resource Estimate. Snowden Mining Industry Consultants, August 2004. Internal Report No. 4729, (Unpublished).
 
104

 
Snowden, 2004: Ahmavaara and Konttijärvi analysis of assay quality control data. Snowden Mining Industry Consultants, September 2004. Internal Report No 4729, (Unpublished), 280pp
 
Snowden, 2002: Analysis of QAQC data from the Suhanko and Narkaus projects, Arctic Platinum Partnership. Snowden Mining Industry Consultants, September 2002. Internal Report No. 4155, (Unpublished). 232pp
 
Saltikoff B., Puustinen, K., and Tontti, M., 2006. Metallogenic Zones and Metallic Mineral Deposits in Finland, Explanation to the Metallogenic Map of Finland. Geological Survey of Finland Special Paper 35.
 
105

 
21.0        CERTIFICATES OF AUTHORS
 
CERTIFICATE of AUTHOR
 
I, Fred H. Brown, do hereby certify that:
 
1.                                       I am a self-employed consultant with a business address of Main St. Lynden, WA, 98264 USA and a postal address of PO Box 332, Lynden WA 98264, USA.
 
2.                                        I graduated with a Bachelor of Science degree in Geology from New Mexico State University in 1987. In addition, I have obtained a Graduate Diploma in Engineering (Mining) in 1997 from the University of the Witwatersrand and a Master of Science in Engineering (Civil) from the University of the Witwatersrand in 2005.
 
3.                                        I am registered with the South African Council for Natural Scientific Professions as a Professional Geological Scientist (registration number 400008/04) and the American Institute of Professional Geologists as a Certified Professional Geologist (certificate number 11015). I am also a member of the Society of Economic Geologists.
 
4.                                        I have worked as a geologist continuously since my graduation from university in 1987.
 
5.                                        I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person”.
 
6                                           I am responsible for the preparation of Sections 16.0, and portions of Sections 5.0, 13.0, and 18.0 of this report titled “Technical Report, Mineral Resource Estimate, and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” and dated October 29, 2007.
 
7                                           I visited the Konttijärvi and Ahmavaara property for one day on May 4, 2006.
 
8.                                        I have not had prior involvement with the property that is the subject of this report.
 
9.                                        As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading;
 
10. I am independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-101.
 
11. I have read National Instrument 43-101 and Form 43-101F1, and this report has been prepared in compliance with that instrument.

DATED this 29th Day of October, 2007	

{SIGNED and SEALED}	

 	 	
Fred H. Brown. CPG, Pr. Sci. Nat.	
 
106

 
DR. WAYNE D. EWERT, P.GEO.
 
CERTIFICATE of AUTHOR
 
I, Wayne D. Ewert, P. Geo., residing at 10 Langford Court, Brampton, Ontario, L6W 4K4 do hereby certify that:
 
1.                I am a principal of P & E Mining Consultants Inc. and currently contracted as a consultant by North American Palladium Ltd.
 
2.                I graduated with an Honours Bachelor of Science degree in Geology from the University of Waterloo in 1970 and with a PhD degree in Geology from Carleton University in 1977.
 
3.                I am a member of the Geological Association of Canada, of the Canadian Institute of Mining and Metallurgy and a P. Geo., Registered in the Province of British Columbia (APEGBC No. 18965) and the Province of Ontario (APGO No. 0866).
 
4.                I have worked as a geologist for a total of 37 years since obtaining my B.Sc. degree.
 
5.                I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:

 	•    Principal, P&E Mining Consultants Inc.,	2004 – Present	
 	•    Vice-President, A.C.A. Howe International Limited,	1992 – 2004	
 	•    Canadian Manager, New Projects, Gold Fields Canadian Mining Limited,	1987 – 1992	
 	•    Regional Manager, Gold Fields Canadian Mining Limited,	1986 – 1987	
 	•    Supervising Project Geologist, Getty Mines Ltd.,	1982 – 1986	
 	•    Supervising Project Geologist III, Cominco Ltd.,	1976 – 1982	
 
6.                I am responsible for Sections 1 through 14, 17, 20 as well as the overall structuring of the technical report titled “Technical Report, Mineral Resource Estimate, and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” and dated October 29, 2007.
 
7.                I have not visited the Konttijärvi or Ahmavaara properties and have not had any prior involvement with the Suhanko Project.
 
8.                As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading;
 
9.                I am independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-101.
 
10.          I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

DATED this 29th Day of October, 2007	

{SIGNED and SEALED}	

 	 	
Dr. Wayne D. Ewert P. Geo.	
 
107

 
EUGENE J. PURITCH, P. ENG.
 
CERTIFICATE OF AUTHOR
 
I, Eugene J. Puritch, P. Eng., residing at 44 Turtlecreek Blvd., Brampton, Ontario, L6W 3X7, do hereby certify that:
 
1                                       I am President of P&E Mining Consultants Inc.
 
2.                                    I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as obtaining an additional year of undergraduate education in Mine Engineering at Queen’s University. In addition I have also met the Professional Engineers of Ontario Academic Requirement Committee’s Examination requirement for Bachelor’s Degree in Engineering Equivalency. I have practiced my profession continuously since 1978. My summarized career experience is as follows:

•    Mining Technologist - H.B.M.&S. and Inco Ltd.	1978-1980	
•    Open Pit Mine Engineer – Cassiar Asbestos/Brinco Ltd	1981-1983	
•    Pit Engineer/Drill & Blast Supervisor – Detour Lake Mine	1984-1986	
•    Self-Employed Mining Consultant – Timmins Area	1987-1988	
•    Mine Designer/Resource Estimator – Dynatec/CMD/Bharti	1989-1995	
•    Self-Employed Mining Consultant/Resource-Reserve Estimator	1995-2004	
•    President – P & E Mining Consultants Inc.	2004-Present	
 
3.                                    I am a mining consultant currently licensed by the Professional Engineers of Ontario (License No. 100014010) and registered with the Ontario Association of Certified Engineering Technicians and Technologists as a Senior Engineering Technologist. I am also a member of the National and Toronto CIM.
 
4.                                        I am the co-author of Sections 11-13, 16-19 and 22 of this technical report titled “Technical Report, Mineral Resource Estimate, and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” and dated October 29, 2007.
 
5.                                    I have not visited the Konttijärvi or Ahmavaara properties.
 
6.                                    As of the date of this certificate, to the best of my knowledge, information, and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
 
7                                       I am an independent of the issuer applying the test in section 1.4 of NI 43-101.
 
8.                                    I have had no prior involvement with the Suhanko Project that is the subject of this Technical Report.
 
9.                                    I have read NI 43-101 and Form 43-101F1 and the Report has been prepared in compliance therewith.
 
10.                              I am a “qualified person” for the purposes of NI 43-101 due to my experience and current affiliation with a professional organization (Professional Engineers of Ontario) as defined in NI 43-101.

DATED this 29th day of October, 2007	

{SIGNED and SEALED}	

 	 	
Eugene Puritch, P.Eng.	
 
108

 
JASON RICKARD, P.GEO
 
CERTIFICATE of AUTHOR
 
I, Jason Rickard, P.Geo., residing at 417 Strand Ave., Thunder Bay, Ontario, P7B 5A8, do hereby certify that:
 
1.                                        As of the effective date of this report, I was employed by, and carried out this assignment for, North American Palladium Ltd., Metals Exploration Division, 710 Norah Cres., Thunder Bay, ON, P7C 4T8, tel. (807) 623-8005, fax (807) 623-8074, e-mail jrickard@napalladium.com.
 
2.                                        I am a graduate of Carleton University and hold the following academic qualifications:

H.BSc. (Geology) Carleton University	1998	
M.Sc (Geology) Carleton University	2000	
 
3.                                        I am a registered Professional Geoscientist with the Association of Professional Geoscientists of Ontario (Registration Number 1017) and have practiced my profession continuously since 2000. My experience includes mineral exploration, advanced exploration and petrology mostly within nickel, copper, platinum-group element deposit types.
 
4.                                        I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” as defined by NI43-101.
 
5.                                        I am responsible for section 23.0 – “Mineralogy Deportment” of this technical report titled “Technical Report, Mineral Resource Estimate, and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” and dated October 29, 2007.
 
6.                                        I have not visited the Konttijärvi or Ahmavaara deposits but have reviewed data and drill core from the property.
 
7.                                        I have had no prior involvement with the Suhanko Project that is the subject of this Technical Report.
 
8.                                        As of the date of this certificate, to the best of my knowledge, information, and belief, section 23.0 – “Mineralogy and Deportment” of the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
 
9.                                        I am not independent of the issuer applying the test in section 1.4 of NI 43-101.
 
10.                                  I have read National Instrument 43-101 and Form 43-101F1, and the section 23.0 – “Mineralogy Deportment” of the Technical Report has been prepared in compliance with that instrument and form.

DATED this 29th Day of October, 2007	

{SIGNED and SEALED}	

 	 	
Jason Rickard P. Geo .	
 
109

 
DR. DAVID J. KING, P.ENG.
 
CERTIFICATE of AUTHOR
 
I, David J. King, P. Eng., residing at unit 210, 102 Bronte Road, Oakville, Ontario, L6L 6J5 do hereby certify that:
 
1.                I am an employee of Aker Kvaerner E&C. which is currently contracted as a consultant by North American Palladium Ltd.
 
2.                I graduated in Chemical Engineering from the University of Manchester with an Honours Bachelor of Science degree in 1964, a Masters degree in 1966 and a PhD degree in 1968.
 
3.                I am a member of the Canadian Institute of Mining and Metallurgy and a P. Eng. in the Province of Ontario (PEO No. 23824501).
 
4.                I have worked as a chemical engineer for a total of 42 years since obtaining my B. Sc degree.
 
5.                I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.
 
6.                I have supervised the preparation of Sections 24 to 32 of the technical report titled “Technical Report, Mineral Resource Estimate and Preliminary Assessment (Scoping Study) on the Suhanko Project, Northern Finland” and dated October 29, 2007.
 
7.                I have not visited the Konttijärvi and Ahmavaara properties.
 
8.                As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all the scientific and technical information that is required to be disclosed to make the report not misleading.
 
9.                I am independent of the issuer applying all of the tests in section 1.4 of the National Instrument 43-101.
 
10. I have read National Instrument 433-101 and Form 43-101F1 and the Technical Report has been prepared in compliance with the instrument and form.

DATED this 29th Day of October, 2007	

{SIGNED and SEALED}	

 	 	
Dr. David J. King, P.Eng.	
 
110

 
SECTION 2:         ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT AND PRODUCTION PROPERTIES - PRELIMINARY ECONOMIC ASSESSMENT OF THE SUHANKO PROJECT
 
The reader is advised that the following section of this report contains a preliminary economic assessment of the Suhanko Project that is preliminary in nature, and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary assessment will be realized, in whole or in part.
 
111

 
22.0        MINING OPERATIONS
 
22.1        INTRODUCTION AND DESIGN PARAMETERS
 
At this stage of the exploration of the Konttijärvi and Ahmavaara properties, the open pit mining areas within approximately 3 km of each other have provided indications of sufficient nickel, copper, gold, platinum and palladium mineral concentrations to justify computer modeling and preparation of a preliminary economic assessment. The in-situ mineral resources of the Konttijärvi and Ahmavaara deposits were described and tabulated in Section 16.0 of this report.
 
The block models for the two pits were assessed for potential mining recovery using the Whittle 4X pit optimizing software. This is the mining industry’s standard optimization software package, used by many mine engineering consulting firms and many large mining companies. The result of the optimizing process is a pit shell that contains all the mineralized material that is potentially economically mineable according to the physical and economic input design parameters.
 
Open Pit Optimization is an iterative process due to the initial design parameters being selected on the basis of assumptions regarding mining rate, fleet size, process recoveries, waste/ore ratio and metal prices and, as a result, the operating costs and revenues are, as a result, preliminary in nature. Continuing study and testwork since the initial resource assessments were made have resulted in refinements to some input data. The design parameters used in the current pit optimizations for design purposes are set out in Table 22.1 and Table 22.2.
 
Physical Pit Design Parameters
 
The block dimensions for the Konttijärvi deposit are 15 m x 15 m x 8 m and those for the Ahmavaara deposit are 30 m x 30 m x 8 m. The bulk density for the mineable portion of the resources at Konttijärvi average 2.91 t/m³ while the waste rock is 2.85 t/m³ and overburden is 1.85 t/m³. The bulk density for the mineable portion of the resources at Ahmavaara average 2.92 t/m³ while the waste rock is 2.85 t/m³ and overburden is 1.85 t/m³.
 
In the detail design stage for both pits, the bench geometry was set to give a 50 degree inter-ramp slope angle with batter angles set at 75. This bench configuration resulted in a 13.7m wide berm every 24 vertical metres of final bench height
 
112

 
Table 22.1
Konttijärvi Pit Design Parameters (all currency USD)

Physical Parameters	 	 	 	 	 	 	
Block Dimensions	 	 	 	15m x 15m x 8m	 	 	
Bulk Densities	 	Ore	 	2.91 t/m³	 	 	
 	 	Waste Rock	 	2.85 t/m³	 	 	
 	 	Overburden	 	1.85 t/m³	 	 	
Haulage Ramps	 	Width	 	26m (20 m roadway, plus safety berm & ditch)	 	 	
 	 	Gradient	 	10%	 	 	
Wall Slopes	 	All rock	 	50° inter-ramp angle	 	 	
Economic Parameters	 	 	 	 	 	 	
Mining Costs	 	Ore & Waste	 	$1.84 USD /t Mined	 	 	
 	 	Overburden	 	$1.23 USD/t Mined	 	 	
 	 	Processing Cost	 	$5.51 USD/t Milled	 	 	
 	 	General & Admin. Costs	 	$0.67 USD/t Milled	 	 	
 	 	Concentrate Freight	 	$48.90 USD/t of Concentrate	 	 	
 	 	 	 	$215 USD/t of Copper	 	 	
 	 	Smelter Charges	 	Concentrate	 	 	
Product Recoveries to Concentrate	 	 	 	 	 	 	
 	 	Nickel	 	65%	 	 	
 	 	Copper	 	73%	 	 	
 	 	Gold	 	75%	 	 	
 	 	Platinum	 	65%	 	 	
 	 	Palladium	 	69%	 	 	
 	 	Conc. Ratio	 	71.43:1	 	 	
Payable Quantities in Concentrate	 	 	 	 	 	 	
 	 	Nickel	 	90%	 	 	
 	 	Copper	 	85%	 	 	
 	 	Gold	 	85%	 	 	
 	 	Platinum	 	89%	 	 	
 	 	Palladium	 	91%	 	 	
Metal Prices	 	 	 	Price	 	Refining	
(18 mo trailing average)	 	Nickel	 	$9.04 USD/lb	 	$0.53 USD/lb	
 	 	Copper	 	$2.56 USD/lb	 	$0.30 USD/lb	
 	 	Gold	 	$546 USD/oz	 	$14.41 USD/oz	
 	 	Platinum	 	$1,057 USD/oz	 	$18.70 USD/oz	
 	 	Palladium	 	$277 USD/oz	 	$16.72 USD/oz	
 
113

 
Table 22.2
Ahmavaara Pit Design Parameters (all currency USD)

Physical Parameters	 	 	 	 	 	 	
Block Dimensions	 	 	 	30 m x 30 m x 8 m	 	 	
Tonnage Factors	 	Ore	 	2.91 t/m³	 	 	
 	 	Waste Rock	 	2.85 t/m³	 	 	
 	 	Overburden	 	1.85 t/m³	 	 	
Haulage Ramps	 	Width	 	26 m (20-m roadway, plus safety berm & ditch)	 	 	
 	 	Gradient	 	10%	 	 	
Wall Slopes	 	All rock	 	50° inter-ramp angle	 	 	
Economic Parameters	 	 	 	 	 	 	
Mining Costs	 	Ore & Waste	 	$1.84 USD /t Mined	 	 	
 	 	Overburden	 	$1.23 USD/t Mined	 	 	
 	 	Processing Cost	 	$5.70 USD/t Milled	 	 	
 	 	General & Admin. Costs	 	$0.67 USD/t Milled	 	 	
 	 	Concentrate Freight	 	$48.90 USD/t of Concentrate	 	 	
 	 	 	 	$215 USD/t of Copper	 	 	
 	 	Smelter Charges	 	Concentrate	 	 	
Product Recoveries to Concentrate	 	 	 	 	 	 	
 	 	Nickel	 	65%	 	 	
 	 	Copper	 	75%	 	 	
 	 	Gold	 	77%	 	 	
 	 	Platinum	 	65%	 	 	
 	 	Palladium	 	69%	 	 	
 	 	Conc. Ratio	 	71.43:1	 	 	
Payable Quantities in Concentrate	 	 	 	 	 	 	
 	 	Nickel	 	90%	 	 	
 	 	Copper	 	85%	 	 	
 	 	Gold	 	85%	 	 	
 	 	Platinum	 	89%	 	 	
 	 	Palladium	 	91%	 	 	
Metal Prices	 	 	 	Price	 	Refining	
(18 mo trailing avg.)	 	Nickel	 	$9.04 USD/lb	 	$0.53 USD/lb	
 	 	Copper	 	$2.56 USD/lb	 	$0.30 USD/lb	
 	 	Gold	 	546 USD/oz	 	$14.41 USD/oz	
 	 	Platinum	 	$1,057 USD/oz	 	$18.70 USD/oz	
 	 	Palladium	 	$277 USD/oz	 	$16.72 USD/oz	
 
114

 
Economic Design Parameters
 
The resource model blocks contained grades for five recoverable minerals: nickel, copper, gold, platinum, and palladium. These grades were combined into one net smelter return (NSR) block value so that the Whittle program was optimizing dollars rather than grades. This was a three-step process. First, the recoverable portion of each mineral was calculated, then the payable value of each, and finally the five mineral values were added together for one block value. From the NSR values have been deducted the costs of concentrate transportation, smelting, and refining. They are based on the plant recoveries, metal prices, and payable quantities shown in Table 22.1 and Table 22.2 above.
 
22.2        CUT-OFF GRADE
 
In the pit optimization process, the pit shell is expanded only if the resource block being assessed has sufficient value to cover all costs (mining, processing, and administration) for itself and for overlying blocks. This is referred to as the external cut-off grade and will be of the order of USD $10.23/t for Konttijärvi and USD $11.89/t for Ahmavaara, depending on the value of overlying blocks. However, every block within the optimized shell must be excavated, loaded into a truck, and hauled to a discharge point, normally on surface. When a truckload of material reaches the pit rim, it is considered to be economic if its recoverable value exceeds the further costs of processing plus administration. This is referred to as the internal cut-off grade and for the Konttijärvi deposit utilized in the pit optimization was USD $6.18/t and for Ahmavaara USD $6.37/t, being the sum of the estimated costs of processing and administration.
 
22.3        PIT DESIGN
 
The optimum Whittle shell is a useful design tool but, while mathematically correct, it does not produce a practical pit design. The shell is a series of concentric, inverted cones that may shrink down to one single block, whereas a practical minimum bench width is about 30 m. It may also create square corners or small step-outs that are not conducive to a stable wall design. Finally, it follows the average wall slope(s) defined in the design parameters without the necessary access ramp.
 
The pit shell is imported into the Gemcom software package as a three-dimensional surface that is contoured to match the pit bench intervals. The final pit is designed interactively, from the bottom up, using those contours as a guide for creating the toe and crest lines with Gemcom’s pit design tool utility while smoothing the walls and inserting the haulage ramp. This process tends to leave a little of the optimized material within the walls and to take additional waste rock above the ramp. It was repeated for each of the two deposits considered in this study. This pit design was developed for the purpose of maximizing grade for the first 10 years of production.
 
The Konttijärvi deposit mineralization has a strike of 105 o and dips 30 o to 35 o to the northeast. The 10 year design pit contains 17.5 Mt of potentially mineable mineralization with 6% dilution added at grades of 0.032% Ni, 0.040% Cu, 0.029 g/t Au, 0.089 g/t Pt and 0.302 g/t Pd resulting in final production diluted grades of 0.072% Ni, 0.114% Cu, 0.108 g/t Au, 0.386 g/t Pt and 1.346g/t Pd and a mine gate (inclusive of concentrate shipping, smelting, refining, penalties and marketing costs) NSR value of USD $23.44/t. The overall strip ratio is 1.22:1
 
The Ahmavaara deposit mineralization is divided into west and east zones, both of which have a southeast strike and a northeast dip. Ahmavaara West has a variable
 
115

 
dip between –70 o to –5 o while Ahmavaara East dips from –25 o to –10 o . The 10 year pit design contains 57.5 Mt of potentially mineable mineralization with 6% dilution added at grades of 0.032% Ni, 0.093% Cu, 0.049 g/t Au, 0.082 g/t Pt and 0.440 g/t Pd resulting in final production diluted grades of 0.086% Ni, 0.203% Cu, 0.125 g/t Au, 0.229 g/t Pt and 1.085 g/t Pd and a mine gate (inclusive of concentrate shipping, smelting, refining, penalties and marketing costs) NSR value of USD $222.12/t. The overall strip ore ratio is 2.01:1
 
Plans indicating the general configuration of the Konttijärvi and Ahmavaara Pits are provided in Figures 22-1 and 22-2.
 
The Konttijärvi Pit as designed has dimensions of approximately 500 m wide, 900 m long and up to 160 m deep. The 26 m wide main ramp is and descends at a -10% gradient from elevation 150 m to the main floor at 88 m. Two sub-pits are accessed to the 40 m and 16 m elevations by ramps on which the last 24 vertical metres are driven at -15% and with a width of only 18m for single lane traffic. Some robbing of high grade from these floors was designed for a final depth 8 m below the final ramp limits. Typical cross-sections through the Konttijärvi pit are provided in Figures 22-3 and 22-4.
 
The Ahmavaara Pit as designed has dimensions of approximately 500 m wide, 1,800 m long and up to 225 m deep. The 26 m wide main ramp descends at a -10% gradient from elevation 140 m to the main floor at 28m. Two sub-pits are accessed to the 8 m and -72 m elevations by ramps on which the last 24 vertical metres are driven at -15% and a width of only 18m for single lane traffic. Some robbing of high grade from these floors was designed for a final depth 8 m below the final ramp limits. Typical cross-sections through the Ahmavaara pit are provided in Figures 22-5, 22-6 and 22-7.
 
During the pre-stripping period (a portion of Year 1), waste rock that is non-acid generating could be used for tailings dam and road construction. All potentially acid generating (PAG) waste rock from the two pits will be stored near the pits and hauled back into the Konttijärvi pit after production ceases.
 
116

 

117

 

118

 

Figure 22.3: Konttijärvi Deposit Cross Section 54350 E (Looking Grid East)

 
Figure 22.4:  Konttijärvi Deposit Cross Section 54000E (Looking Grid East)
 
119

 

Figure 22.5:  Ahmavaara Deposit Cross Section 30450E (Looking Grid East)

 
Figure 22.6:  Ahmavaara Deposit Cross Section 30900E (Looking Grid East)
 
120

 

Figure 22.7:  Ahmavaara Deposit Cross Section 31575E (Looking Grid East)
 
22.4         DILUTION
 
In order to mine the mineralized materal, some waste rock must also be taken as dilution. The amount of this external dilution depends on a number of factors, largely related to the size of the mining equipment and the selective mining unit (SMU). It was assumed that the average external dilution would be the width of the burden of the  drill and blast pattern (5m burden by 6m spacing) on both the hanging wall and footwall mineralized limits. This was calculated by assessing the average mineable zone width at 24m vertical intervals. Due to the flat dipping nature of these deposits and 8 m active mining benches, a resultant external dilution of 6.0% was calculated at wall rock grades determined from the block model. The average grades of the diluting material are shown in Table 22.3 below.
 
Table 22.3
Estimated Grades of Diluting Material

Diluting Grade	 	Konttijärvi	 	Ahmavaara	 	
Ni (%)	 	0.032	 	0.032	 	
Cu (%)	 	0.040	 	0.093	 	
Au (g/t)	 	0.029	 	0.049	 	
Pt (g/t)	 	0.089	 	0.082	 	
Pd (g/t)	 	0.302	 	0.440	 	
 
121

 
22.5         MINEABLE PORTION OF THE MINERAL RESOURCES
 
The estimated diluted, potentially mineable portion of the mineral resources contained within the Konttijärvi and Ahmavaara pit area design perimeters are set out in Table 22.4 together with the associated waste rock quantities. The mineable portion of the mineral resource total is 75,000,000 t, which at the proposed milling rate of 21,500 t/d, would provide a project life of 10.0 years.
 
Table 22.4:  Estimated Diluted Potentially Mineable Portion of the Mineral Resources for the First 10 Years of Mine  Life

May 28/07	 	 	 	 	 	 	 	 	 	 	 	 	 	NSR	 	Waste & OVB	 	 	 	Strip	 	
Konttijärvi	 	Tonnes	 	Ni%	 	Cu%	 	Au g/t	 	Pt g/t	 	Pd g/t	 	$	 US/t	 	Tonnes	 	Total Tonnes	 	Ratio	 	
Measured	 	12,445,000	 	0.075	 	0.113	 	0.104	 	0.390	 	1.346	 	$	23.75	 	 	 	 	 	 	 	
Indicated	 	4,698,000	 	0.065	 	0.116	 	0.117	 	0.361	 	1.298	 	$	22.19	 	 	 	 	 	 	 	
Inferred	 	357,000	 	0.056	 	0.127	 	0.157	 	0.543	 	1.987	 	$	29.14	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	21,300,000	 	38,800,000	 	1.22	 	
Ahmavarra	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	17,771,000	 	0.079	 	0.197	 	0.108	 	0.187	 	0.933	 	$	21.67	 	 	 	 	 	 	 	
Indicated	 	35,370,000	 	0.090	 	0.210	 	0.134	 	0.248	 	1.166	 	$	25.57	 	 	 	 	 	 	 	
Inferred	 	4,359,000	 	0.076	 	0.171	 	0.119	 	0.239	 	1.043	 	$	22.30	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	113,400,000	 	170,900,000	 	2.01	 	
Combined	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Measured	 	30,216,000	 	0.078	 	0.162	 	0.106	 	0.271	 	1.103	 	$	22.53	 	 	 	 	 	 	 	
Indicated	 	40,068,000	 	0.087	 	0.199	 	0.132	 	0.261	 	1.181	 	$	25.17	 	 	 	 	 	 	 	
Inferred	 	4,716,000	 	0.075	 	0.168	 	0.122	 	0.262	 	1.114	 	$	22.82	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	134,700,000	 	209,700,000	 	1.8	 	
 
Note: Mineral resources that are not mineral reserves do not have demonstrated economic viability. The reader is cautioned that inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that any value from such resources will be realized either in whole or in part.
 
At the metal prices utilized in the November 30, 2006 18-month trailing average metal price estimate, it was seen that nickel provides 33% total NSR value, while copper contributes 20%, gold 5%, platinum 19% and palladium 23%.
 
22.6         ADDITIONAL MINEABLE RESOURCES
 
The purpose of this study was to evaluate the Konttijärvi and Ahmavaara pits for a period of 10 years to determine the economic viability of the project. There are additional mineable resources beyond the 10 years that were not subject to mine design or production scheduling because it was deemed that the first 10 years of operations would be adequate to determine project viability.
 
Table 22.5 shows the total combined estimated diluted, potentially mineable portion of the mineral resources.
 
22.7         MINE PLANNING BASIS
 
The mine planning process began with the development of preliminary capital and operating expenditure estimates for 5.0 and 7.5 Mt/a processing rates based on pit optimizations. The results of the assessment indicated that a processing rate of 7.5 Mt/a offered an effective mining rate and lower unit costs. The results were then used to further delineate the ultimate pit limits, haul roads, crusher and waste rock dump locations, haul distances, and to identify cost saving opportunities such as:
 
•       the interactive development of Konttijärvi and Ahmavaara pits
•       the sequencing of production in order to reduce pre-stripping and increase revenues in the initial years of mining
 
122

 
•       the use of waste rock in tailings starter dam construction and mine/plant site roads
•       sizing the Owner-Operators’ mine equipment fleet based on typical annual total material production requirements.
 
In consideration of the above, the proposed mining operations are based on a 21,500 t/d, 350 d/a processing rate over a 10 year period. The mine planning was thus, done on an initial 10 year basis using the assumption that the first phase of mining would be considered a “starter” pit operation.
 
Table 22.5:  Total Mineable Portion of the Mineral Resources Contained within the Ultimate Pit Shells, Ahmavaara and Konttijärvi Deposits

Category	 	Tonnes	 	Pd g/t	 	Pt g/t	 	Au g/t	 	Ni%	 	Cu%	 	
Ahmavarra Deposit	 	
Measured	 	27,230,000	 	0.900	 	0.179	 	0.108	 	0.077	 	0.199	 	
Indicated	 	87,602,000	 	1.017	 	0.211	 	0.123	 	0.084	 	0.210	 	
Sub - Total, M&I	 	114,832,000	 	0.989	 	0.203	 	0.119	 	0.082	 	0.207	 	
Inferred	 	22,168,000	 	0.988	 	0.225	 	0.108	 	0.074	 	0.168	 	
Waste & OVB	 	269,800,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	406,800,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.97	 	 	 	 	 	 	 	 	 	 	 	
Konttijärvi Deposit	 	
Measured	 	18,844,000	 	1.342	 	0.387	 	0.101	 	0.069	 	0.121	 	
Indicated	 	18,826,000	 	1.262	 	0.353	 	0.102	 	0.054	 	0.131	 	
Sub -Total, M&I	 	37,670,000	 	1.302	 	0.370	 	0.101	 	0.062	 	0.126	 	
Inferred	 	3,230,000	 	1.411	 	0.390	 	0.109	 	0.040	 	0.159	 	
Waste & OVB	 	76,300,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	117,200,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.87	 	 	 	 	 	 	 	 	 	 	 	
Combined Ahmavaara Plus Konttijärvi Deposits	 	
Measured	 	46,074,000	 	1.081	 	0.264	 	0.105	 	0.074	 	0.167	 	
Indicated	 	106,428,000	 	1.060	 	0.236	 	0.119	 	0.079	 	0.196	 	
Sub - Total, M&I	 	152,502,000	 	1.067	 	0.245	 	0.115	 	0.077	 	0.187	 	
Inferred	 	25,398,000	 	1.042	 	0.246	 	0.108	 	0.070	 	0.167	 	
Waste & OVB	 	346,100,000	 	 	 	 	 	 	 	 	 	 	 	
Total Material	 	524,000,000	 	 	 	 	 	 	 	 	 	 	 	
Strip Ratio	 	1.95	 	 	 	 	 	 	 	 	 	 	 	
 
Note: Mineral resources that are not mineral reserves do not have demonstrated economic viability.
 
22.8         MINE PRODUCTION SCHEDULE
 
The mine planning process began with the development of preliminary capital and operating expenditure estimates based on 3.5, 5.0 and 7.5 Mt/a processing rates based on pit optimizations. The result of the assessment indicated that a processing rate of 7.5 Mt/a offered a more effective mining rate and lower unit costs.
 
Table 22.6 shows the sources of material scheduled for each year of the mine life. As indicated in Table 22.6, production in the first two years of the mine life comes from the Konttijärvi mine, year three from both pits and the remaining 7 years thereafter from Ahmavaara.
 
123

 
Table 22.6
Production Schedule

Year	 	Pit	 	Ore	 	Ni	 	Cu	 	Pt	 	Pd	 	Au	 	NSR	 	Waste	 	Overburden	 	Total	 	Waste/Ore	 	 	 	
 	 	 	 	(t)	 	(%)	 	(%)	 	(g/t)	 	(g/t)	 	(g/t)	 	(USD/t)	 	(t)	 	(t)	 	(t)	 	Ratio	 	t/d	 	
Pre-strip	 	Konttijarvi	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	1 139 134	 	2 360 866	 	3 500 000	 	 	 	70 000	 	
 	 	Konttijarvi	 	7 432 838	 	0.076	 	0.108	 	0.371	 	1.289	 	0.098	 	22.85	 	9 236 575	 	1 047 211	 	17 716 625	 	1.38	 	 	 	
 	 	Ahmavaara	 	67 162	 	0.105	 	0.159	 	0.303	 	1.207	 	0.148	 	27.12	 	432 527	 	11 533 687	 	12 033 376	 	178.17	 	 	 	
1	 	 	 	7 500 000	 	0.076	 	0.109	 	0.371	 	1.289	 	0.099	 	22.89	 	9 669 102	 	12 580 898	 	29 750 000	 	2.97	 	85 000	 	
 	 	Konttijarvi	 	7 500 000	 	0.069	 	0.106	 	0.361	 	1.252	 	0.098	 	21.63	 	6 977 223	 	 	 	14 477 223	 	0.93	 	 	 	
 	 	Ahmavaara	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	11 940 920	 	3 331 857	 	15 272 777	 	 	 	 	 	
2	 	 	 	7 500 000	 	0.069	 	0.106	 	0.361	 	1.252	 	0.098	 	21.63	 	18 918 143	 	3 331 857	 	29 750 000	 	2.97	 	85 000	 	
 	 	Konttijarvi	 	2 560 823	 	0.072	 	0.153	 	0.503	 	1.788	 	0.166	 	30.48	 	573 486	 	 	 	3 134 309	 	0.22	 	 	 	
 	 	Ahmavaara	 	4 939 177	 	0.077	 	0.150	 	0.266	 	1.110	 	0.112	 	22.14	 	20 651 232	 	1 025 282	 	26 615 691	 	4.39	 	 	 	
3	 	 	 	7 500 000	 	0.075	 	0.151	 	0.347	 	1.342	 	0.131	 	24.99	 	21 224 718	 	1 025 282	 	29 750 000	 	2.97	 	85 000	 	
4	 	Ahmavaara	 	7 500 000	 	0.078	 	0.163	 	0.276	 	1.160	 	0.125	 	23.24	 	22 250 000	 	 	 	29 750 000	 	2.97	 	85 000	 	
5	 	Ahmavaara	 	7 500 000	 	0.077	 	0.172	 	0.250	 	1.093	 	0.119	 	22.53	 	17 000 000	 	 	 	24 500 000	 	2.27	 	70 000	 	
6	 	Ahmavaara	 	7 500 000	 	0.073	 	0.173	 	0.228	 	1.030	 	0.106	 	21.18	 	8 250 000	 	 	 	15 750 000	 	1.10	 	45 000	 	
7	 	Ahmavaara	 	7 500 000	 	0.077	 	0.191	 	0.194	 	0.958	 	0.101	 	21.07	 	6 500 000	 	 	 	14 000 000	 	0.87	 	40 000	 	
8	 	Ahmavaara	 	7 500 000	 	0.095	 	0.230	 	0.174	 	0.971	 	0.111	 	23.97	 	4 750 000	 	 	 	12 250 000	 	0.63	 	35 000	 	
9	 	Ahmavaara	 	7 500 000	 	0.105	 	0.241	 	0.212	 	1.109	 	0.124	 	27.18	 	4 750 000	 	 	 	12 250 000	 	0.63	 	35 000	 	
10	 	Ahmavaara	 	7 500 000	 	0.109	 	0.302	 	0.247	 	1.281	 	0.213	 	32.14	 	975 000	 	 	 	8 475 000	 	0.13	 	24 214	 	
Total	 	 	 	75 000 000	 	0.083	 	0.184	 	0.266	 	1.148	 	0.123	 	24.08	 	115 426 097	 	19 298 903	 	209 725 000	 	1.80	 	 	 	
 
124

 
22.9         PROPOSED MINE OPERATIONS
 
The conceptual mining operation is envisioned to be conducted with conventional open pit mining equipment.
 
Blasting, Loading and Haulage
 
Blast holes will be drilled using conventional drills. An allowance has been included to allow some pre-shearing of ultimate pit walls. A hydraulic shovel will be used to excavate blasted waste rock and load 180t haul trucks. The waste rock and overburden will be hauled to the nearby waste rock and overburden storage areas for the Konttijärvi and Ahmavaara pits as shown in Figure 22-7.
 
A large wheel loader will be used to excavate blasted mineralized material and load 180t haul trucks. The material will be hauled to the gyratory crusher. Another wheel loader will be available to serve as a back-up unit to the main production loader and to reclaim material from the stockpile located at the crusher. Haul roads will be maintained using a large grader, and road dust will be controlled using a water truck.
 
Pit Dewatering
 
Intercepted run-off and groundwater will be collected in located in the flows of the open pit mines sumps and pumped to the concentrator. It is assumed that no additional measures will be required to dewater the pits. The pits will be allowed to flood at closure.
 
Potentially Acid Generating Waste Rock
 
Once the Konttijärvi pit has been exhausted at the end of year three of the 10 year mine production plan, it will be backfilled with any potentially acid generating waste rock from Ahmavaara and from a stockpile from Konttijärvi. All potentially acid generating waste rock will be sub aqueously stored in the exhausted Konttijärvi pit. In a scenario whereby the two open pits are operated for an extended mine life in excess of 20 years, potentially acid generating intermediate waste rock storage areas will need to be identified in order that this material will be sub aqueously deposited in the exhausted pits prior to mine closure.
 
Mine Closure
 
The exhausted open pits will be allowed to flood and fill up to form pit lakes. Overburden from their storage areas will be placed in a layer over the reduced final slopes of the waste rock storage areas. The overburden layer will be vegetated with native grasses and trees.
 
Mine Equipment Maintenance
 
A mine shop garage will be constructed between the Konttijärvi and Ahmavaara pits. The garage will include six service bays as well as a welding shop, maintenance planning offices, a critical-parts warehouse, and a maintenance dry. The mine equipment re-fuelling station is to be located adjacent to the shop. A fuel truck will also be used to re-fuel equipment in the field. A cold storage building and yard areas for equipment parking and lay down will also be developed.
 
125

 

126

 
22.10       MINE EQUIPMENT SELECTION
 
Mill Feed requirements are met by concurrent operations in the Konttijärvi and Ahmavaara open pits. The material from each of the pits is truck hauled to the crusher located adjacent to the mill. Waste rock is truck hauled to waste rock dumps located within one kilometre of the top of the pit ramps. The haulage truck fleet size was estimated based on a combined average ore production rate, a stripping ratio of 1.8 and average haul distances based on the preliminary pit designs. The haul truck cycle times based on using Cat 789C haul trucks or equivalent are summarized in Table 22.7.
 
Table 22.7
Estimated Haulage Truck Cycle Times

Mill Feed Konttijärvi Pit	 	Mill FeedAhmavaara Pit	
 	 	Waste Stripping	 	 	 	Waste Stripping	
20.6 minutes/load	 	19.4 minutes/load	 	25.5 minutes/load	 	222.3 minutes/load	
 
The haul truck speeds were selected based on performance information presented in the Caterpillar Handbook Version 35. The speed of haul trucks traveling down the pit ramp and on the pit floor was limited to 30 km/h (18 mph). It is estimated that fourteen CAT 789C trucks or equivalent will be required. These haul trucks have a nominal payload capacity of 180 t (198 tons); gross diesel power of 1,417 kW (1900 hp); and an operating width of 7.67 m (25 ft 2 in.’). Haul trucks available from other manufacturers such as, but not limited to, Komatsu could also be considered. The key pieces of equipment are listed in Table 22.8.
 
Table 22.8
Key Pit Equipment

Pit Equipment	 	Quantity	
 	 	 	
Production blasthole drills	 	2	
 	 	 	
Face shovel	 	2	
 	 	 	
Front end loader	 	2	
 	 	 	
Haul trucks	 	14	
 	 	 	
Road grader	 	2	
 	 	 	
RT Dozer	 	2	
 	 	 	
Dozer	 	2	
 
127

 
22.11       MINE LABOUR REQUIREMENTS PER DAY
 
The estimated direct and indirect mine labour requirements are summarized in Table 22.9 while a breakdown of the mine direct labour requirements is shown in Table 22.10.
 
Table 22.9
Estimated Mine Direct and Indirect Labour Summary per Day

Area	 	Indirect Labour	 	Direct Labour	 	Total	 	
Pit operations	 	8	 	56	 	64	 	
Maintenance	 	3	 	18	 	21	 	
Mine Dry	 	2	 	 	 	2	 	
Total	 	13	 	74	 	87	 	
 
Table 22.10
Estimated Mine Operations Direct Labour Requirements per Day

Mine Operations Area	 	Direct Labour Requirements	
Drilling	 	4	
Blasting	 	6	
Shovel	 	4	
Front End Loader	 	2	
Haul Truck	 	28	
RT Dozers/Dozers	 	8	
Other equipment	 	4	
Total	 	56	
 
22.12       OTHER ON-SITE SUPPLIERS AND CONTRACTORS
 
Blasting Agent and Explosives Supplier
 
An on-site supplier will supply blasting agents, explosives and blasting accessories to the mine and will load blast holes. The suppliers’ facilities and explosive and detonator magazines will be located in secure areas on-site.
 
Site Security
 
A security contractor will be used to guard access and patrol the mine site.
 
128

 
23.0         MINERALOGY DEPORTMENT
 
23.1         INTRODUCTION
 
A mineralogical study of the different lithologies that host the PGE, copper, and nickel mineralization at Ahmavaara and Konttijärvi was undertaken. The purpose of this study was to identify the various PGE mineral associations, sulphide textures and grain interlocking textures. Of particular interest were the host sulphide and silicate minerals and the relative amount of PGEs locked up in silicates. These are factors which may have an impact on metal recoveries. Other factors that could have an impact on metal recoveries were also investigated, including the grain size of the PGE, the distribution of the PGE within the rocks and their host minerals, the types of sulphide present and the fine grained intergrowth of these sulphides with gangue minerals.
 
The ore was evaluated by analysis of the in-situ PGE distribution within the different rock types from the Konttijärvi and Ahmavaara deposits. For simplicity, these two deposits will be referred to here as Suhanko. Also, the use of PGE will be used to refer to the platinum group minerals and electrum (Au-Ag alloy).
 
Data from various historical studies was also used to generate some of the observations contained in this section. Some of the historic information was edited and put into a format that was useable to generate comparative graphs and figures between the various datasets. Care was taken, as much as possible, so as not to compare “apples to oranges”.
 
The following is meant to provide a reasonable scoping-level assessment of the mineralogy and deportment of the Konttijärvi and Ahmavaara deposits. In addition, some comparisons to the Lac des Îles deposit have been included. More detailed accounts of the studies and a list of the samples used are provided in Rickard (2006a, 2006b).
 
23.2         PGE MINERALOGY
 
23.2.1      BACKGROUND
 
Some historic work has been completed on the Suhanko deposits that involved analysis of the PGE minerals. An academic study completed by Iljina (1994) was part of a PhD thesis investigating all of the PGE deposits associated with the Portimo Complex (Narkaus and Suhanko intrusions). A PGE deportment study by Richards and Martin (2001) of Lakefield Research was completed that focused on the measurement of PGE grain liberation after the rock is crushed and ground. An important study was done by Hayward and Robb (2005) whose main focus was measurement of the mineralogical variability within the Konttijärvi and Ahmavaara deposits.
 
All of these studies produced useful data, but there were limited amounts of PGE identified and there is a discrepancy between the PGE deportment studies and the mineralogy studies regarding the amount of PGE hosted by silicates at Suhanko. The PGE deportment study (Richards and Martin, 2001) suggests that only small amounts of PGE are associated with silicates, whereas Hayward and Robb (2005) indicate that the majority of PGE are associated with silicates.
 
129

 
23.2.2      SAMPLING AND METHODS
 
The PGEs are very small and difficult to locate using a petrographic microscope, so the slides were scanned using backscattered electron imaging (within a Scanning Electron Microscope(SEM) to highlight the PGEs. The PGEs were measured and counted; the host silicates and sulphides identified and pictures were taken of the textures. A quantitative analysis was collected for each PGE and the surrounding host minerals. When large clusters of PGEs were encountered, a representative number of quantitative analyses were collected and the remainder of the minerals were qualitatively identified using the Energy Dispersive Spectra (EDS) spectra.
 
23.2.3      PGE DISTRIBUTION AND GRAIN SIZE
 
The PGE minerals within the Suhanko deposits are typically subhedral, but may be elongated or somewhat amorphous. It is common to find the grains elongated along sulphide-silicate grain boundaries, cleavage planes or grain fractures. In terms of distribution, the PGEs are typically found in three ways:
 
a)              completely enclosed by sulphide,
b)              along sulphide-silicate grain boundaries, or
c)              completely enclosed by silicates (Figure 23-1).
 
It is also common to find PGEs scattered as “clusters” within sulphides and silicates [Figure 23-1 (D)].
 
To properly estimate the volume of PGE associated with silicates or sulphides, one must consider the area of the PGEs analyzed. This study shows that the majority of PGEs (in terms of total area) are associated with sulphides and smaller amounts are completely enclosed by silicates. This is consistent with Richards and Martin (2001), but is contrary to Hayward and Robb (2005). Iljina (1994) did not calculate the area of the PGEs. However, compilation of the grain count data from Iljina (1994) indicates that more PGE were associated with sulphides than silicates.
 
130

 

Figure 23-1:  PGE Minerals within the Suhanko Deposits
 
When considering the grain size of PGEs, the long dimension of each grain was used. Data from this study shows that the average grain size of the PGEs is 12.06µm (n=272), with a range of 0.81µm to 118.0µm. Analyses of Suhanko rocks done by Richards and Martin (2001) have shown that the PGEs range in size from 0.5µm to 19.2µm with an average size of 3.42µm (n=68). Hayward and Robb (2005) found that the PGE grains are typically in the range of <1 to 5µm, with some from 20-100µm. The Richards and Martin (2001) data suggests that the PGEs are quite small compared to the other studies. This discrepancy could be due to the fact that the Richards and Martin (2001) analyses were done on crushed rocks, possibly breaking the larger PGEs.
 
In comparison of the Konttijärvi and Ahmavaara deposits, it was found that the distribution of the PGEs within silicates and sulphides is quite similar. At Konttijärvi, 24.72% of the PGEs are enclosed by silicates and 75.28% are associated with sulphide. At Ahmavaara, 31.98% of the PGEs are enclosed by silicates and 68.02% are associated with sulphide. The average grain size of the PGEs within each of the deposits is virtually identical; 12.1µm (n=123) for Konttijärvi and 11.96µm (n=150) for Ahmavaara. The difference in PGE grain size between stratigraphic units within each deposit was also examined, but the size of the dataset is too small to draw any firm conclusions. However, it seems that the PGEs within the basement rocks of both deposits have the smallest average grain size; 7.43µm (n=56) for Konttijärvi and 5.42µm (n=15) for Ahmavaara.
 
131

 
23.2.4      PGE TYPES
 
A large number of different types of PGE have been documented in this study and by other workers. Iljina (1994) found that the PGEs could be “roughly classified into
 
•               Elements
 
•               alloys of platinum-group elements only or of PGE and base metals
 
•               sulphides
 
•               sulpharsenides, and
 
•               compounds of PGE with one or more of the following elements: arsenic, antimony, tellurium, tin, silver, lead, selenium and mercury.
 
The abundance of each of the PGE species has been calculated by estimating the total area of each of the PGE species. The results are different from Richards and Martin (2001), but the current dataset is more than three times larger so it may be more accurate. The current study suggests that the Pd-tellurides (26.55%) and Pd-arsenides (15.34%) are the most abundant PGE, followed by Pt-arsenides (11.47%) and Pd-sulphides (10.47%).
 
Both studies show that the Pd-arsenides and Pd-sulphides are the largest sized PGE in the Suhanko rocks.
 
23.2.5      PGE – MINERAL ASSOCIATIONS
 
It was found in this study that the PGE minerals are preferentially associated with specific sulphide and silicate mineral species. The PGEs have a strong affinity for chalcopyrite with lesser amounts associated with pyrrhotite, pentlandite, and pyrite, and only a few attached to sphalerite, galena and millerite (Figure 23-2 Section a). In terms of silicate minerals, the PGEs have a strong correlation with amphibole, with lesser numbers associated with feldspar, chlorite, epidote, and biotite, and a minor amount with quartz, pyroxene and other unknown minerals (Figure 23-2 Section b).
 
132

 

Figure 23-2:  PGE Mineral Associations
 
23.3         SULPHIDE MINERALOGY
 
23.3.1      SULPHIDE MINERALS
 
Examination of samples from the current study and results from other reports (Evans and Guerney, 2002; McKay, 2002; Martin and McKenzie, 2006) indicate that the main sulphide minerals identified within the Suhanko deposits are pyrrhotite, chalcopyrite, pentlandite, pyrite, bornite, sphalerite, violarite, millerite, covellite, and the main oxides are magnetite, ilmenite, chromite and rutile. This is consistent with QemScan analyses by McKay (2002) and Marion (2006) that show the most abundant sulphides
 
133

 
are pyrrhotite and chalcopyrite, followed by pyrite, pentlandite and other phases; however there is variation between stratigraphic units.
 
Electron microprobe analyses of sulphide minerals were completed by Hayward and Robb (2005). The average Ni content of pentlandite grains from the Konttijärvi and Ahmavaara deposits were found to be 34.37 wt% (n=12, S.D.=1.65 wt%) and 34.04 wt% (n=29, S.D.=2.36 wt%), respectively. The average Ni content of pyrrhotite grains from the aforementioned deposits were 0.44 wt% (n=12, S.D.=0.21 wt%) and 0.51 wt% (n=24, S.D.=0.25 wt%), respectively.
 
Some of the sulphide textures that may be of interest from a liberation stand point include:
 
pentlandite (Pn) flames and lamellae in pyrrhotite (Po) (Figure 23-3 Section a); Cp=chalcopyrite)
 
medium-grained pentlandite (Figure 23-3 Section b)
 
fine suphide disseminations in silicates (Figure 23-3 Sections c and d)
 
sulphide interlocked with magnetite identified as “Mt” (Figure 23-3 Section e)
 
pyrite (Py) intergrown with Ni-Cu-bearing sulphides (Figure 23-3 Section f). As noted by Evans (2000), the intergrowth of the sulphide minerals with the fine grained gangue minerals may have a deleterious effect in terms of flotation.
 
23.3.2      SULPHIDE LIBERATION
 
Sulphide liberation data has been generated from a number of different studies (McKay, 2002; Liipo, 2001; Martin and McKenzie, 2006). Each of these studies was completed on drill core composite samples collected from both the Konttijärvi and Ahmavaara deposits. The main sulphide minerals were classified based on their degree of liberation from gangue minerals.
 
The study by McKay (2002) investigated 4 composite samples and found that pentlandite exhibited the poorest sulphide liberation in all of the samples, while the results for pyrrhotite and chalcopyrite were much better. In terms of combined sulphide liberation, it was found that the Konttijärvi Pyroxenite sample had the lowest degree of liberation (69.8%), while the Ahmavaara Gabbro had the highest (81.8%).
 
A study by Liipo (2001) looked at 2 test composites. It was found that there was better liberation of the sulphides within the Ahmavaara samples compared to the Konttijärvi samples, which may have been due to the coarser grained nature of the sulphides in the Ahmavaara rocks.
 
134

 

Figures 23-3:  Sulphide Mineral Textures
 
Martin and McKenzie (2006) reported the results of testing conducted on two composite samples. They found that the flotation process is highly selective for chalcopyrite and that “more than 80% of the chalcopyrite in the fast floating concentrate is highly (>80%) liberated, while 79% of the chalcopyrite in the tails is present as low-grade locks (<30% liberated)”. It was also found that the flotation is less selective for pentlandite and there may be a problem floating fine, liberated pentlandite. This was demonstrated by the fact that 42% of the pentlandite in the tails was >80% liberated. Of note, 81% of the pentlandite in the “fast floating” concentrate was liberated.
 
23.3.3      SULPHIDE OXIDATION
 
An investigation into the degree of sulphide oxidation within the Konttijärvi and Ahmavaara deposits was completed by Carter and Martin (2005) and the results were presented and summarized by Naumanen (2005). The study found that the amount of sulphide alteration was low and the majority of the samples showed no effects of
 
135

 
oxidation, except in close proximity to faults and shears. Some minor sulphide alteration was identified within both of the Ahmavaara and Konttijärvi deposits at the bottom of the pyroxenite units and at the top of the respective marginal series units (alteration products were bornite ± millerite).
 
23.4         SGS LAKEFIELD - BULK MODAL ANALYSES (BMA)
 
Due to the heterogenic nature of the Ahmavaara and Konttijärvi deposits it is necessary to understand the variability of the minerals between each of the major stratigraphic units as minerals such as talc can be problematic within a flotation circuit. The purpose of Bulk Modal Analysis (BMA) of each of the major stratigraphic units is to determine if certain rock units have elevated amounts of problematic minerals.
 
The results of recent QemScan analyses (Marion, 2006) and previous QemScan analyses (McKay, 2002) are presented together in Table 23.1. The McKay (2002) analyses may be more representative due to the size and number of the samples used to make the composite samples. Comparison of the BMA data shows that the three gabbroic units at Ahmavaara are all quite similar in terms of composition
 
The basement rocks underlying each of the deposits are dissimilar in appearance, and it is now confirmed that they are also dissimilar in composition. Almost one-quarter (23.2%) of the Ahmavaara basement is composed of quartz, whereas the Konttijärvi basement contains only 4.3% quartz. There is also a notable contrast in the biotite content, in which the Ahmavaara basement contains 15.1% biotite compared to 6.7% biotite in the Konttijärvi basement. This may be of interest as biotite has been found to be host to some of the PGEs.
 
Despite the similarities in appearance between the peridotite marker units in each of the deposits, it seems that there is actually a significant difference in the talc content between the two units. The data shows that the Ahmavaara peridotite marker contains 50.3% talc while the Konttijärvi peridotite marker contains only 24.1% talc.
 
A contrast in talc content is also seen between the pyroxenite units. However, the newer dataset (Marion, 2006) divides the Mg-rich micas into “talc” and “Mg mica clays”, whereas the older dataset (McKay, 2002) has no field for “Mg mica clays”. It is possible that the all the Mg-rich micas were grouped together into the “talc” field within the older dataset, which would account for the contrast in talc content between the Konttijärvi pyroxenite and Ahmavaara pyroxenite composite samples.
 
136

 
Table 23.1:QemScan Analyses

 	 	SGS Lakefield Report: Marion (2006)	 	Lakefield Research Report: McKay (2002)	 	
Deposit	 	Ahmavaara	 	Ahmavaara	 	Ahmavaara	 	Ahmavaara	 	Konttijarvi	 	Konttijarvi	 	Konttijarvi	 	Ahmavaara	 	Ahmavaara	 	KonttiKonttijarvi	 	
 	 	 	 	 	 	 	 	 	 	Marginal	 	 	 	 	 	 	 	 	 	 	 	
 	 	 	 	 	 	 	 	 	 	Series/	 	 	 	 	 	 	 	 	 	 	 	
 	 	.	 	 	 	Lower	 	Central	 	Transition	 	Peridotite	 	 	 	 	 	Peridotite	 	 	 	
Stratigraphic Unit	 	Pyroxenite	 	Upper Gabbro	 	Gabbro	 	Gabbro	 	Zone	 	Marker	 	Basement	 	Basement	 	Marker	 	Pyroxenite	 	
Mineral Msss(%)	 	AHMA-06-OO1	 	Ai-IMA-O6-002	 	A1IM#06-003	 	AHMA-06-004	 	KoNr-o6-o1	 	KoNr-oG-002	 	KONT-O6-003	 	 	 	 	 	 	 	
Pentlandite	 	0.1	 	0.3	 	0.2	 	0.1	 	0.3	 	0.2	 	0.1	 	0.1	 	0.2	 	0.1	 	
Chalcopyrite	 	0.6	 	0.7	 	0.7	 	0.7	 	0.5	 	0.2	 	0.5	 	0.8	 	0.6	 	0.4	 	
Bornite	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0	 	0	 	0	 	
2ndary Cu-Sulfide	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.1	 	0.1	 	0	 	
Pyrite	 	0.0	 	0.8	 	0.3	 	0.7	 	0.3	 	0.1	 	0.2	 	0.2	 	0	 	0	 	
Pyrrhotite	 	0.0	 	0.9	 	0.0	 	1.6	 	0.4	 	0.3	 	0.1	 	1.7	 	1.4	 	0.2	 	
Other_Sulfides & PGM	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0	 	0	 	0	 	
 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	
Quartz	 	0.6	 	2.9	 	4.4	 	6.4	 	3.6	 	0.0	 	4.3	 	23.2	 	0	 	0	 	
Feldspar	 	0.0	 	23.1	 	24.9	 	19.6	 	31.7	 	0.0	 	47.7	 	35.3	 	0	 	0	 	
Orthopyroxene	 	1.8	 	0.5	 	0.2	 	0.3	 	0.7	 	4.6	 	0.1	 	 	 	 	 	 	 	
Diopside/Salite	 	1.2	 	2.0	 	4.1	 	4.5	 	1.9	 	0.8	 	1.5	 	 	 	 	 	 	 	
Ca-Amphiboles	 	37.6	 	27.6	 	33.3	 	33.7	 	34.9	 	28.9	 	14.4	 	 	 	 	 	 	 	
Ampbiboles	 	0.0	 	4.1	 	6.8	 	7.7	 	0.8	 	0.0	 	2.2	 	6.4	 	6.7	 	26.9	 	
Epidote/Zoisite	 	0.0	 	5.2	 	1.6	 	1.3	 	1.9	 	0.0	 	2.0	 	0.5	 	0	 	1.3	 	
Carbonates	 	0.1	 	1.0	 	0.1	 	0.6	 	0.0	 	1.4	 	0.0	 	0.5	 	6.7	 	0.7	 	
Talc	 	1.4	 	0.2	 	0.1	 	0.2	 	0.4	 	24.1	 	0.1	 	0	 	50.3	 	12.7	 	
Chlorite	 	45.8	 	17.9	 	8.6	 	12.0	 	11.1	 	33.5	 	12.1	 	15.2	 	26.6	 	53.2	 	
Biotite	 	0.2	 	2.8	 	4.2	 	2.2	 	5.4	 	1.3	 	6.7	 	15.1	 	5	 	3.9	 	
Muscovite	 	0.0	 	1.2	 	0.4	 	0.4	 	0.2	 	0.0	 	0.6	 	 	 	 	 	 	 	
Mg Mica Clays	 	10.2	 	8.3	 	8.4	 	7.2	 	5.6	 	2.9	 	6.6	 	 	 	 	 	 	 	
Fe-Oxides	 	0.1	 	0.6	 	0.7	 	0.6	 	0.3	 	1.4	 	0.5	 	0.5	 	2.1	 	0.5	 	
Other/Accesory	 	0.3	 	0.1	 	0.2	 	0.2	 	0.1	 	0.2	 	0.3	 	0.4	 	0.1	 	0.1	 	
Other	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	0.0	 	 	 	 	 	 	 	
Total	 	100.0	 	100.0	 	100.0	 	100.0	 	100.0	 	100.0	 	100.0	 	100.1	 	99.7	 	100	 	
 
137

 
23.5        WHOLE ROCK GEOCHEMISTRY
 
There is quite an extensive database of whole rock geochemical data for both the Konttijärvi and Ahmavaara deposits. The database is composed of geochemical analyses of drill core samples taken from each of the deposits and is representative of each of the major stratigraphic units (see Brown, 2006a and 2006b).
 
Of importance is the MgO content of individual units. As expected, the peridotite marker units in each of the deposits contain the greatest amount of MgO. The peridotite marker unit at Konttijärvi has a mean MgO content of 26.98% (n=213, max: 32.13% MgO, min: 20.54% MgO), while the Ahmavaara peridotite marker has a mean MgO content of 24.67% (n=207, max: 30.24% MgO, min: 11.71% MgO). The Konttijärvi pyroxenite and Ahmavaara pyroxenite units also contain significant MgO contents which are 21.08% MgO (n=69, max: 25.24% MgO, min: 13.82% MgO) and 20.20% MgO (n=110, max: 25.27% MgO, min: 5.30% MgO) respectively.
 
23.6        COMPARISON TO THE LAC DES ÎLES DEPOSIT
 
23.6.1     SILICATES AND SULPHIDES
 
It is important to compare the Suhanko deposits to the Lac des Îles deposit as they are all low-grade PGE-Cu-Ni deposits hosted by mafic to ultramafic rocks. Though many of the rock types are similar between the two deposits, there are significant differences. Most important is the gneissic basement ore found at Suhanko. This rock type is harder than any of the LDI rocks and contains much more quartz than anything found at LDI. Table 23.2 shows the summarized results of QemScan Bulk Mineral Analyses of rock from the Suhanko and LDI deposits. The talc and biotite contents are significantly higher at Suhanko, while the feldspar content is much higher at LDI. There are also minor differences in the amphibole, chlorite and quartz content between the two deposits.
 
Table 23.2:  QemScan Bulk Mineral Analyses

Mineral (%)	 	Suhanko (Avg. of 10 Composites)	 	LDI (Mill Head, n=64)	 	
Feldspar	 	16.2	 	33.1	 	
Amphibole	 	32.2	 	26.1	 	
Chlorite	 	21.0	 	16.1	 	
Quartz	 	4.0	 	9.4	 	
Clay Minerals	 	6.3	 	6.4	 	
Pyroxene	 	3.1	 	3.8	 	
Epidote/Zoisite	 	1.2	 	3.2	 	
Talc	 	8.0	 	0.7	 	
Biotite/Phlogopite	 	4.2	 	0.2	 	
Oxides	 	0.7	 	0.2	 	
Carbonate	 	1.0	 	0.2	 	
Other	 	0.5	 	0.2	 	
Sulphide	 	1.6	 	0.4	 	
Total	 	99.9	 	100	 	
 
Overall, the Suhanko deposits contain slightly more sulphide than the LDI deposit. The sulphides at Lac des Îles are typically found as fine grained disseminations and medium to coarse grained blebs, however there are occasional instances of thin veinlets and rare occurrences of semi-massive stringers. The sulphide textures at Suhanko are somewhat similar, but sulphide veins and veinlets are very common and
 
138

 
semi-massive and massive sulphides are fairly significant towards the base of the Ahmavaara deposit. The types of sulphide found are similar, but there is a considerable difference in the relative abundance of pyrite and pyrrhotite. The Suhanko deposits contain much more pyrrhotite relative to pyrite (Figure 23-4 Section a), while the opposite is true for LDI (Figure 23-4 Section b).

 
Figure 23-4: Sulphide Content
 
23.6.2     PGE MINERALS
 
Much of the LDI mineralogical data used for comparison is taken from Rickard (2004). It is difficult to compare the PGE distribution between the Suhanko deposits and the LDI deposit because an estimate of the volume of PGEs (grain area measurements) associated with silicates versus sulphides is not available for LDI. The best comparison that can be made is with PGE grain counts, which do not take into account the size of the PGEs. There is a greater proportion of PGEs associated with sulphides at Suhanko compared to LDI. However, this contrast is likely much less drastic in terms of volume as the PGEs associated with sulphide at LDI are larger than those associated with silicates.
 
The PGEs in each deposit are associated with similar types of sulphide, but there is a much greater proportion of PGEs at LDI that are associated with pyrite compared to Suhanko. Also, there are more PGEs associated with pyrrhotite at Suhanko compared to LDI (Figures 23-5 Sections a) and b). This may be due to the contrasting pyrite and pyrrhotite contents seen between LDI and Suhanko.
 
139

 

Figure 23-5: Suhanko PGE & LDI Sulphide Associations
 
In terms of PGE distribution within silicates, it seems that the PGEs within both deposits are often linked to amphibole, chlorite and epidote. There is also a strong association to feldspar within Suhanko that is seen to a lesser degree at LDI. The only significant difference between the two deposits is that PGEs are often associated with biotite at Suhanko, which is not seen at all within LDI. Biotite is quite common at Suhanko, but is a relatively rare phase at LDI.
 
On average the PGEs at LDI are smaller than those at Suhanko, but the difference is small and they could be considered quite similar. The average grain size of the PGEs at Suhanko is 12.06µm (n=272), with a range of 0.81µm to 118.0µm; while the average grain size at LDI is 6.68µm (n=3933) with a range of <1 to 200µm. The disparity in grain size may be attributed to the difference in the size of the dataset.
 
The types of PGE found at each of the deposits are similar, but there are some differences. For example, there are no Rh-, Ir- or Ru-bearing PGEs at LDI, but there are a small number present at Suhanko. Again, it is difficult to compare estimated PGE volumes due to the lack of grain area measurements at LDI, but it seems that the most common PGEs within both deposits are Pd tellurides.
 
23.6.3     GEOCHEMISTRY
 
The MgO content of LDI ore is relatively similar to ore from Suhanko, but the Suhanko rocks are more MgO-rich. The average MgO content of the Ahmavaara and Konttijärvi composite samples used in the pilot plant test are 14.3% and 15.3%, respectively. It is important to compare these to the LDI mill feed. The average MgO content of 281 mill feed samples from September 1, 2005 to June 27, 2006 was 9.10% (max: 11.31% MgO, min: 6.79% MgO). Further, 64 samples taken from LDI’s Roby Zone and High Grade Zones (including the Offset High Grade Zone) returned an average MgO content of 16.1%. The most MgO-rich rocks at both deposits are the ultramafics. The peridotitic rocks at Suhanko contain up to 32.13% MgO, which is somewhat higher than the pyroxenite at LDI that has up to 27.5% MgO.
 
23.7        SUMMARY POINTS
 
23.7.1     KONTTIJÄRVI AND AHMAVAARA (SUHANKO)
 
•                   PGEs are found i) completely enclosed by sulphide, ii) along sulphide-silicate grain boundaries, ii) completely enclosed by silicates iv) scattered as “clusters” within sulphides and silicates.
 
140

 
•                   The majority of PGEs are associated with sulphides, with a smaller proportion completely enclosed by silicates.
 
•                   The PGE distribution (association with sulphides/silicates) is similar between the Konttijärvi and Ahmavaara deposits.
 
•                   Average grain size of the PGEs is 12.06µm (n=272), with a range of 0.81µm to 118.0µm.
 
•                   The larger PGEs, on average, are associated with sulphide.
 
•                   PGEs within the basement rocks of both deposits have the smallest average grain size; 7.43µm (n=56) for Konttijärvi and 5.42µm (n=15) for Ahmavaara.
 
•                   PGE/electrum abundance: Pd-tellurides (26.55%), Electrum (22.61%), Pd-arsenides (15.34%), Pt-arsenides (11.47%) and Pd-sulphides (10.47%), Unknowns (5.97%), Pd-stannides (3.67%), Pd-antimonides (2.91%), Pt-tellurides (0.83%), Rh-minerals (0.08%), PdHg (0.06%), Pd-bismutharsenides (0.04%).
 
•                   The Pd-arsenides and Pd-sulphides are the largest PGEs in grain size.
 
•                   PGEs have a strong correlation for chalcopyrite with lesser amounts associated with pyrrhotite, pentlandite, and pyrite, and only a few attached to sphalerite, galena and millerite.
 
•                   PGEs have a strong correlation with amphibole, with lesser numbers associated with feldspar, chlorite, epidote, and biotite, and a minor amount with quartz and pyroxene.
 
•                   The three gabbroic units at Ahmavaara are all quite similar in terms of composition.
 
•                   Ahmavaara Peridotite marker contains 50.3% talc while the Konttijärvi Peridotite marker contains only 24.1% talc.
 
•                   Ahmavaara Basement contains 23.2% quartz, whereas the Konttijärvi Basement contains only 4.3% quartz.
 
•                   Ahmavaara Basement contains 15.1% biotite compared to 6.7% biotite in the Konttijärvi Basement.
 
•                   The Ahmavaara deposit is much more sulphide rich in comparison to the Konttijärvi deposit
 
•                   Pyrrhotite is the most abundant sulphide within the Suhanko deposits, followed by chalcopyrite, pyrite, pentlandite, and other sulphides (bornite, sphalerite, violarite, millerite and covellite).
 
•                   Chalcopyrite is the most abundant sulphide within the Konttijärvi and Ahmavaara Pyroxenite units.
 
•                   Chalcopyrite is also the most abundant sulphide within the Konttijärvi Marginal Series/Transition Zone and in the Konttijärvi Basement.
 
•                   The intergrowth of the sulphide minerals with the fine grained gangue minerals may have a deleterious effect in terms of flotation (Evans, 2000).
 
141

 
•                   In terms of combined sulphide liberation, it was found that a Konttijärvi Pyroxenite composite sample had the lowest degree of liberation (69.8%), while an Ahmavaara Gabbro composite sample had the highest (81.8%) (McKay, 2002).
 
•                   There was better liberation of the sulphides within Ahmavaara composite samples compared to Konttijärvi composite samples, which may have been due to the coarser grained nature of the sulphides in the Ahmavaara rocks (Liipo, 2001). Sulphides are most commonly intergrown or attached to amphibole, feldspar and chlorite with lesser amounts associated with biotite and quartz.
 
•                   Sulphides are most commonly intergrown or attached to amphibole, feldspar and chlorite with lesser amounts associated with biotite and quartz
 
•                   No significant alteration/oxidation of the sulphide minerals is present.
 
23.7.2     COMPARISON TO THE LAC DES ÎLES DEPOSIT
 
The following is a point form summary comparison of the geology and mineralogy of the Konttijärvi and Ahmavaara deposits (Suhanko) with the Lac des Îles deposit:
 
Similarities
 
•                   The mineralization is predominantly hosted by mafic to ultramafic rocks.
 
•                   PGE mineralization is for the most part spatially associated with sulphide mineralization.
 
•                   On average the PGEs at LDI are smaller (6.68µm) than those at Suhanko (12.06µm), but the difference is small and may be considered quite similar.
 
•                   PGE textures and general distribution are similar (associated with both sulphides and silicates).
 
•                   PGEs associated with sulphides tend to be larger.
 
•                   The gangue minerals are similar; however there are important differences in the abundances of each.
 
Differences
 
•                   The ore-bearing rocks at Suhanko are predominantly layered igneous rocks and granite gneisses; the ore-bearing rocks at LDI are predominantly igneous breccia, pyroxenite, and pegmatitic gabbroic rocks.
 
•                   Overall, the Suhanko deposits are more sulphide-rich, especially the Ahmavaara deposit.
 
•                   Higher pyrite, lower pyrrhotite content at LDI.
 
•                   More PGEs associated with pyrite at LDI.
 
•                   More biotite in Suhanko rocks.
 
•                   PGEs associated with biotite at Suhanko.
 
•                   Slightly greater proportion of PGEs is associated with sulphide at Suhanko compared to LDI.
 
142

 
•                   Some PGE species are different between the deposits.
 
•                   MgO contents are relatively similar, but the Suhanko rock are more MgO-rich (9.10%, average for 281 LDI mill feed samples; 14.3-15.3% MgO for composites from the Suhanko pilot plant test).
 
143

 
24.0        METALLURGICAL TESTWORK
 
24.1        PREVIOUS TESTWORK
 
Extensive testwork was carried out on the APP prior to the involvement of Aker Kvaerner. The main conclusions from this test work are summarized below.
 
•                   It has been determined that the mineralized samples from the Ahmavaara deposit have a higher fast-floating component as compared to Konttijärvi samples; which resulted in higher  mass pull and higher recoveries with lower concentrate grades. On the other hand, the Konttijärvi samples displayed a higher concentrate grade but produced lower mass pull and overall recoveries.
 
•                   It was determined that blending talcose with non-talcose material for both deposits did not affect recovery. However, overall blending of both deposits resulted in lower recoveries . Recovery of 2PGE+Au was about 5% lower for the overall talcose and non-talcose composites compared to the individual composites.
 
•                   Highest PGE+Au and base metal losses occurred in the +75 micron-size range. Almost all the different sample blends indicated that, as the particle size was decreased, the resulting effect was an increase in recovery.
 
•                   The optimum target grind for the flotation circuit was determined to have a product size of 80% passing 75 microns. At this grind size, a rougher/scavenger recovery of 85% PGE+Au was obtained. Also, at this grind size, cleaner grade and recovery were improved.
 
•                   Fast-floating PGE+Au minerals are recoverable at a coarse grind without slimes and MgO entrainment. Flash flotation was recommended for the recovery of these coarse particles (approximately 200 micron size).
 
•                   Bulk cleaning of the rougher/scavenger concentrate provides higher overall recoveries of the metals than cleaning the major component of the concentrate separately, that is, producing separate copper and nickel concentrates.
 
•                   Mineralogical examination of the scavenger tails, at a product size of 80% passing 75 microns, showed that 92% of the PGE were locked in silicate grain edges and would require further grinding for liberation.
 
•                   Open circuiting of the first cleaner tails would be required in order to reduce talc content.
 
•                   Kinetics curves for mass pull versus recovery suggested that 15- to 20-minute flotation times for secondary flotation was sufficient.
 
•                   Reagent optimization and pH should be further investigated. Testwork performed thus far indicates that pH 7.0 appears to be optimum.
 
24.2        RECENT TESTWORK
 
24.2.1     TRADE-OFF STUDY
 
In June 2006, a Trade-off Study was started by Aker Kvaerner in conjunction with SGS Lakefield on samples obtained in 2005 from the Suhanko Project.
 
Under the direction of Aker Kvaerner, SGS Lakefield was contracted by GFAP to conduct a series of bench tests for the purpose of examining several flow sheet options set forth by Aker Kvaerner in the hopes of improving metallurgical results obtained from previous testwork that was carried out prior to June of 2006. Several options were explored that included:
 
144

 
•                   flash flotation
•                   two-stage grinding – fine grind
•                   sequential copper/nickel flotation
•       bulk copper/nickel flotation
•                   magnetic separation.
 
Each option above was tested on samples of both the Ahmavaara and Konttijärvi orebodies. For both ores, test conditions were similar.
 
Due to the fact that the samples tested were taken from a master composite collected back in 2005, a series of flotation tests were performed in order to establish sample integrity. It was determined after a series of rougher kinetics flotation tests that mass pulls, recovery and grade were similar with those tests conducted back in 2005, indicating, at the time, that sample degradation or aging did not inhibit recovery or concentrate grade. Subsequently, during the locked-cycle testwork, it became more apparent that the sample may have been affected by degradation. Refer to Figure 24-1.

 
Figure 24-1: Mass Pull vs Recovery
 
During the initial phase of the recent testwork program, a total of 18 flotation tests were performed. Grind times along with optimum grind size were established from previous work. Grind times of 25 minutes were used on a 2 kg charge at a pulp density of 65% solids. With this grind time, a K 80 of 70 microns was obtained. Reagent addition rates were taken from the 2005 testwork program. Although reagent optimization was not achieved in 2005, the testwork did enable the selection of a reagent scheme that resulted in good metal recoveries.
 
Summary of Results
 
From the recent testwork program, several factors became apparent that include the following:
 
•                                           Due to 0the higher sulphide content in the Ahmavaara mineralization, the mass pulls for the final concentrate were over 3% while for Konttijärvi they were below 1.5%.
 
•                                           Mass pulls in the rougher circuit of approximately 22 to 30% resulted in optimum recoveries. Mass pulls in excess of 30% did not improve metal recoveries.
 
•                                           The grade recovery relationship is typical for both sets of mineralized samples since the grade can be improved at the expense of metal recovery. Conversely, recovery can be improved at the expense of grade.
 
•                                           Both ores exhibit rapid flotation kinetics, 70 to 80% of metal recovered within the first 10 minutes.
 
•                                           Rougher kinetics were the same for both coarse and fine grind.
 
•                                           Fifty percent (50%) of the copper was recovered in the first 2 minutes of rougher flotation.
 
145

 
•                                           Froth stability improves with the removal of coarse fast floating copper.
 
•                                           Higher metal recoveries were achieved with the Ahmavaara sample. A 7-10% increase for Pd, Cu, and Pt were observed while nickel recoveries were 15% higher than Konttijärvi.
 
•                                           Higher concentrate grades were obtained with the Konttijärvi material due to its lower sulphide content.
 
•                                           The MgO values in the final concentrate were lower for Ahmavaara than Konttijärvi, (3% versus 7%.)
 
•                                           CMC additions were 30% higher for the Ahmavaara sample.
 
•                                           High-grade copper concentrate were easily obtainable, > (25% Cu) for both samples.
 
•                                           For the Ahmavaara material, PGE values in the copper concentrate account for 54% of the total value of the concentrate while for Konttijärvi this amount increases to 75%.
 
•                                           The bulk concentrate produced was low grade, approximately 9% Cu+Ni with 40 to 73 g/t PGEs for both samples.
 
•                                           Nickel depression in the copper concentrate was more effective with the Ahmavaara concentrate (< 1% recovery versus < 5% recovery for Konttijärvi, 0.37% Ni in copper concentrate for Ahmavaara versus 1.46% Ni for Konttijärvi).
 
•                                           Magnetic separation preformed on a bulk concentrate from Konttijärvi resulted in approximately 45% of the total metal value retained with the magnetic pyrrhotite.
 
•                                           Magnetic separation preformed on the copper concentrate cleaner tail from Ahmavaara resulted in approximately 30% of the metal value retained with the magnetic pyrrhotite.
 
•                                           A Higher amount of pyrrhotite is observed in Ahmavaara sample as compared to the Konttijärvi sample.
 
Discussion of Results
 
Rougher Kinetics
 
A series of rougher kinetics tests were performed on both samples in order to evaluate the flotation times required and determine optimum mass pulls. Reagent addition for each sample was kept the same except for CMC additions. As a result of the higher talc content in the Ahmavaara sample, approximately 25% more depressant was added. A series of graphs that depicts the effects of mass pull and recovery versus flotation times are presented in figures 24-2 and 24-3.

 
Figure 24-2: Konttijärvi Mineralization
 
146

 

Figure 24-3 Ahmavaara Mineralization
 
From the graphs, it can be determined that mass pull and recovery begin to level off after 20 minutes. If the flotation times were to be extended past 20 minutes, higher mass pulls may be achieved however more gangue material would be recovered. A general rule of thumb worth noting for establishing rougher flotation times is that the flotation process should be terminated at the point at which incremental concentrate grades drop below the head grade. At this point, the material is no longer worth floating to the rougher concentrate. Further flotation will result in higher cleaning losses due to the difficulty of cleaning low-grade material. Ideal flotation residence times for both ores can then be determined to be 20 minutes, with a rougher mass pulls of approximately 23% for Konttijärvi and 30% for Ahmavaara.
 
From Figures 24-2 and 24-3, it also becomes evident that, for the Konttijärvi sample, more than 70% of each metal was recovered to the rougher concentrate within 10 minutes. This number increases dramatically to better than 84% recovery of each metal to the rougher concentrate for the Ahmavaara sample for the same time period.
 
The results confirm that the flotation kinetics for the Ahmavaara sample are much faster than the Konttijärvi sample. This is due in part to the greater abundance of floatable silicates in the sample and, as a result, recoveries and mass pull for Ahmavaara are much higher. Also, from mineralogical examination conducted in 2005, it was found that Ahmavaara material contained coarser grain copper mineralogy and a larger proportion of sulphide - hosted nickel, which also happens to be coarse grained thus helping speed up flotation kinetics.
 
Flash Flotation
 
A series of flash flotation tests were carried out on both samples. The principle behind flash flotation is to remove quick floating coarse liberated sulphide material early in the initial phase of grinding to prevent the material from becoming over ground, thus decreasing the potential for metal losses further down stream. As a result, a high-grade concentrate can be produced without slimes and MgO entrainment. By incorporating flash flotation in the flow sheet, it does become possible to produce a very clean concentrate, however, unfortunately, it does not enhance overall metal recovery.
 
The feed to flash flotation was ground for 10 minutes at a pulp density of 65% solids. This resulted in a feed material with a K 80 of 85 microns. The total residence time in the cell was two minutes. All tails from flash flotation were collected and reground for 15 minutes and further subjected to flotation. As a result, two products of interest were produced, a flash flotation concentrate and a bulk concentrate.
 
147

 
With both samples, a copper-PGE concentrate was obtained. The remaining bulk copper/nickel concentrate had significantly lower copper and PGE values while the nickel grade remained similar for both products (Tables 24.1 and 24.2.).
 
Table 24.1
Metallurgical Balance - Konttijärvi

Konttijärvi	 	Cumulative Grade	 	Cumulative Recovery	 	
Product	 	Pd(g/t)	 	Cu (%)	 	Ni (%)	 	Pt(g/t)	 	Pd (%)	 	Cu (%)	 	Ni (%)	 	Pt (%)	 	
Flash Copper Concentrate	 	122.173	 	22.031	 	2.604	 	26.038	 	26.87	 	43.07	 	17.11	 	19.30	 	
Bulk Concentrate	 	55.784	 	7.256	 	1.601	 	14.982	 	56.37	 	49.32	 	56.12	 	59.69	 	
Flash + Bulk Conconcentrate	 	57.995	 	7.717	 	1.639	 	15.384	 	83.25	 	92.38	 	73.23	 	78.99	 	
 
Table 24.2
Metallurgical Balance - Ahmavaara

Ahmavaara	 	Cumulative Grade	 	Cumulative Recovery	 	
Product	 	Pd(g/t)	 	Cu (%)	 	Ni (%)	 	Pt(g/t)	 	Pd (%)	 	Cu (%)	 	Ni (%)	 	Pt (%)	 	
Flash Copper Concentrate	 	98.500	 	24.000	 	2.806	 	15.694	 	29.75	 	47.01	 	12.82	 	20.31	 	
Bulk Concentrate	 	33.626	 	5.910	 	2.132	 	6.608	 	55.94	 	48.58	 	74.16	 	55.15	 	
Flash + Bulk Conconcentrate	 	35.463	 	6.411	 	2.156	 	6.871	 	85.68	 	95.59	 	86.98	 	75.46	 	
 
The mass pulls for both samples were well below 1.0% for the flash concentrates and both concentrates had Mg0 values of 3% and lower. As mentioned earlier, the residence time in the flash cell was two minutes.
 
As shown in Figures 24-4 and 24-5, it can be seen that the metal distribution for copper is well over 40% in the flash concentrate for both samples and approximately 25% for the PGEs. This test is consistent with other rougher kinetic tests in that copper flotation is very rapid and close to half of the copper in the concentrate is recovered within the first two minutes. The PGEs for both samples respond quite similarly and their recoveries are almost identical, thus perhaps reflecting the similar PGE grain size distribution. Also of interest is the fact that the copper, and a portion of the PGEs, can be recovered using a relatively coarse grind. However, from the total metal recovery for the duration of the float test, it can be seen that overall recovery is not improved using the flash flotation approach.
 
Since there is no metallurgical benefit to using flash flotation and due to the cost to implement a flash cell along with the difficulty of operating the cell, it can be safely omitted from the flow sheet as a viable option.

 
Figure 24-4: Flash Flotation Bulk Rougher Float
 
148

 

Figure 24-5: Metal distribution in Flash and Bulk Concentrate
 
Two-Stage Grinding/Fine Grinding
 
A series of flotation tests were conducted to determine if two-stage grinding/fine grinding offered any improvement in recovery and grade. The rationale behind this flow sheet was to remove the coarser liberated fast floating metal from the circuit as quickly as possible while still producing a high-grade concentrate and then to further subject the middlings/tails to further grinding to enhance particle liberation for improved metal recovery. The two-stage grinding approach would also eliminate the potential of over grinding the sulphides which, when ground to a fine size fraction, become difficult to float. Refer to Figure 24-6.

 
Figure 24-5: Two-stage Grinding Results
 
The above graphs indicate that 2-stage grinding offers no metallurgical benefit. Mass pulls, metal recoveries and concentrate grade were similar when compared to single-stage grinding. Results indicate that, after the first stage of grinding, metal recoveries have almost reached their maximum with roughly 6 to 8% additional recovery after the second stage of grinding. At this point, it is most probably the extra residence time that is responsible for the additional recovery after the second stage of grinding and not the regrinding.
 
Sequential Copper/Nickel Circuits
 
The purpose of this flow sheet was to produce separate copper and bulk concentrates that can be readily accepted by both copper and nickel smelters.
 
The thought behind the flow sheet was to produce a separate copper concentrate that is generated from the initial rougher flotation. This copper concentrate would be pH adjusted in order to depress the pyrrhotite and nickel. The tails generated from the nickel depression would then be combined with the final copper/nickel bulk concentrate. After nickel depression, it was hoped that the final copper concentrate  would contain low nickel and PGE values. As a result of the nickel depression, the bulk copper/nickel concentrate would now contain high nickel and PGE values with low to moderate copper values which would make it attractive to a nickel smelter.
 
149

 
The bulk copper/nickel circuit would also be used as a buffer to collect the copper, nickel and PGEs that were not initially floated in the copper circuit or lost in the copper cleaning circuit. By doing this, the mass pull in the copper circuit could be low thus enabling the production of a high-grade copper concentrate.
 
Several different tests were performed in order to optimize this circuit; Figures 24-7 to 24-11 are a series of graphs that highlight the best results achieved.

 
Figure 24-7: Percentage Mass Pull versus Recovery for the Ahmavaara Sample

 
Figure 24-8: Mass Pull Versus Grade for the Ahmavaara Sample

 
Figure 24-9: Mass Pull versus Recovery for the Konttijärvi Sample

 
Figure 24-10: Percentage Mass Pull versus Grade for the Konttijärvi Sample
 
150

 

Figure 24-11: Metal Distribution in Each Concentrate for Both Samples
 
Figures 24-7 to 24-11 illustrate several readily apparent factors. Figure24-11 shows the copper concentrate for the Konttijärvi samples contains approximately 75% of the PGE value while for the Ahmavaara ore this number increases to 50%. If a copper smelter offers a high payout on the PGEs, this does not become an issue, however, if the payout is low then the production of a separate copper concentrate becomes financially undesirable.
 
Copper distribution in the copper concentrate is high for both samples which is to be expected, however, for the Konttijärvi ore, nickel separation in the copper circuit was not good resulting in higher than acceptable nickel grade (1.4% Ni). This was not an issue for the Ahmavaara sample. Copper smelter terms and conditions should be investigated further to determine the quality of concentrate that would be acceptable by the smelter.
 
The remaining bulk concentrate generated, combined with the copper cleaner tails, was low in grade with a combined copper/nickel value of 5 to 8% and combined Pd and Pt values of 30 to 40 g/t, making it unattractive for a nickel smelter.
 
As a result of the testwork performed above, it was decided to continue with a copper pre float after the initial roughing stage and then combine this concentrate with the final bulk copper/nickel concentrate at the end, thus abandoning the separate copper/nickel flow sheet option.
 
Magnetic Separation
 
The purpose behind magnetic separation was to clean the bulk concentrate by the removal of magnetic sulphur (pyrrohite) in the hopes of further upgrading the concentrate. A wet high intensity magnetic separator utilizing various currents was employed. The results are presented in figures 24-12 and 24-13.
 
Figures 24-12 and 24-13 indicate that magnetic separation option is less then desirable, with a large amount of valuable metal reporting to the magnetic portion. Due to the limited number of tests performed, ideal conditions were not achieved. The potential for further pyrrhotite rejection may be increased with finer grinding of the concentrate. This will be looked at in the next phase of testwork.
 
151

 

Figure 24-12: Results of Magnetic Separation on the Copper Cleaner Tails for the Ahmavaara Copper Concentrate

 
Figure 24-13: Results from Magnetic Separation on a Bulk Concentrate
 
24.2.2     SCOPING STUDY
 
Immediately upon completion of the Trade-off Study testwork, a second test program was developed for the Scoping Study. The onus on this phase of the testwork was three-fold; first to finalize flow sheet development, secondly to target reagent optimization and finally to perform a series of locked cycle tests on both ores using the best set of conditions developed during bench testing.
 
A total of 15 flotation bench tests were carried out along with 3 locked-cycle tests. Of these, 13 bench tests and 1 locked-cycle test were performed on the Ahmavaara samples while two bench tests and 2 locked-cycle tests were carried out on the Konttijärvi sample. Throughout the course of the test program, it was understood that Ahmavaara would be the first material to be milled therefore most of emphasis was placed on the Ahmavaara material. It should be noted that for both the Ahmavaara and Konttijärvi samples frothing issues developed during the second cycle of the locked-cycle test program which affected the metallurgical performance of both. As a result overall recoveries and grades were less then expected. It is suspected that oxidation of the samples may have had some influence on flotation behaviour.
 
Once the final flow sheet for the Ahmavaara material was established and a logical reagent scheme developed, repeat tests were carried out on the Konttijärvi sample to see if the set conditions derived for the Ahmavaara material were also favourable for optimum grade and recovery on the Konttijärvi ore.
 
The first 4 tests were designed to look at both gravity and magnetic separation. In test # 1a, the final copper concentrate from the copper pre-float was subjected to gravity separation in the hopes of separating the palladium from the chalcopyrite.
 
From the results, it was observed that 8% of the total palladium was associated with the + 38 micron material with virtually no upgrading occurring in this size range. The – 38 micron material consisted of approximately 20% of the total palladium with slight
 
152

 
upgrading occurring in the coarser fractions (concentrate 1). Please see Figure 24-14 below.
 
The results indicate that roughly 30 % of the total palladium is associated with the copper concentrate with very little upgrading possible, thus gravity separation of the copper concentrate could not be justified.

 
Figure 24-14 Gravity Separation of Cu Cone
 
The next series of tests examined the possibility of magnetic separation as a means of rejecting pyrrhotite. Tests were conducted with and without regrind on both the bulk concentrate and combined bulk cleaner tails. Results indicate that a high metal value was associated with the magnetic portion of the pyrrhotite. Fine grain inclusions of palladium and nickel are associated within the magnetic portion of the sulphur, thus rejection of this material resulted in lower recoveries of both palladium and nickel. Please see charts in Table 24.3
 
Table 24.3: Metal Recoveries
 
Metal Recoveries - Non Magnetics

Test #	 	Cu	 	Ni(T)	 	Pd	 	
1a	 	6.37	 	4.86	 	10.38	 	
2a	 	11.71	 	30.06	 	17.70	 	
3	 	4.06	 	2.63	 	4.38	 	
5	 	2.6	 	9.21	 	3.74	 	
 
Metal Recoveries - Magnetics

Test #	 	Cu	 	Ni(T)	 	Pd	 	
1a	 	3.49	 	9.88	 	8.83	 	
2a	 	6.84	 	17.61	 	11.91	 	
3	 	3.54	 	6.18	 	4.49	 	
5	 	0.48	 	5.32	 	0.66	 	
 
Regrind
 
A number of tests were conducted which explored the possible benefit of regrinding the bulk concentrate prior to cleaning in order to improve both grade and recovery. Using similar conditions as in the regrind tests several other tests were carried out which did not involve concentrate regrinding prior to cleaning. Therefore a comparison was drawn between the use of regrind versus no regrind. It should be noted that identical reagent conditions were not maintained also mass pulls varied slightly for each test.
 
153

 
The testwork indicates that pyrrhotite rejection is increased as the cleaner concentrate was reground, resulting in lower metal recoveries for palladium and nickel. This can be attributed to perhaps over grinding of the sulphur or oxidation of the pulp when subjected to regrinding. The ORP was not measured thus this cannot be confirmed. The graphs in Figure 24-15 indicate the influence an metal recovery of regrinding the concentrate on metal recovery.

 
Figure 24-15 Influence of Regrinding on Metal Recovery
 
It is interesting to note that regrinding the concentrate prior to cleaning did not improve the grade of both copper and nickel, in the concentrate however palladium grade was significantly improved as a result of regrinding as illustrated in the Figure 24-16.

 
Figure 24-16 Grade % of Metals in Concentrate after Regrinding
 
Copper Pre-float
 
A series of flotation tests were conducted which examined the effect of a copper pre-float on the final grade and metal recoveries. Operating parameters such as reagent additions and mass pulls varied slightly for each test. For each test the bulk
 
154

 
concentrate was reground. The metal recovered in the copper pre float was subjected to a single stage of cleaning with no regrind. The remaining material was recovered in the scavenger stage and subjected to four stages of cleaning. The copper concentrate and bulk concentrate where then combined together in order to calculate overall recoveries and grades. For the bulk flotation tests, rougher/scavenger concentrate was collected, reground and cleaned four times.
 
From the graphs below, it becomes apparent that a copper pre float had a positive impact on recoveries; however it resulted in a lower grade concentrate. It also became evident, when conducting the tests, that the removal of the copper within the first few minutes of flotation improved both froth stability and movement later on in the bulk cleaning stages. The lower grades achieved with the copper pre float are a result of the improved froth characteristics in the bulk cleaning stage and can be improved with increased silicate depressants. See graphs in Figure 24-17.

 
Figure 24-17 Copper Pre-float Impact on Recoveries
 
Due to the limited amount of testwork that was carried out on the Konttijärvi sample it is difficult to assess the benefit or negative effect of regrinding the concentrate or utilizing a copper pre float. A detailed metallurgical test program needs to be carried out in order to properly evaluate and establish optimum operating parameters.
 
Locked-Cycle Tests
 
Due to the limited amount of material that was available, only one locked-cycle test (LCT) was carried out for the Ahmavaara sample while two locked-cycle tests were performed on the Konttijärvi sample. Due to a time constraint and limited amount of sample, test conditions for the Konttijärvi LCT were taken from what was believed to be the best set of conditions established from Ahmavaara bench test work. It is well recognized that both the Ahmavaara and Konttijärvi samples differ in their mineralogical makeup but, as mentioned, time constraints and limited sample did not allow for a detailed evaluation on the Konttijärvi material.
 
For both samples the flow sheet was similar, that is a copper pre float was utilized with no regrinding of the bulk concentrate. Figure 24-18 shows the process flow sheet that was used for the LCT on both samples.
 
During the locked-cycle testwork, the designed reagent scheme was changed after the 2 nd cycle for each sample in LCT numbers 1 and 2. The reason was due to
 
155

 
frothing issues that arose which resulted in flocculation occurring in the concentrate. The froth formed a thick gel which became very non selective thus silicate depression was difficult. Collector additions were reduced considerably and lime was introduced into the grinding circuit. This did improve froth conditions somewhat.
 
For both materials, copper sulphate was added in order to activate the sulphur in the hopes of increasing recovery, however, after the 2 nd cycle, due to the formation of gel in the froth, it was abandoned and replaced with sodium sulphide. It was believed that the copper sulphate reacted with the CMC to form a gel which resulted in a froth that was non selective. The addition of the sodium sulphide only marginally improved the froth characteristics.
 
As a result of the frothing issues, mass pulls varied greatly between the 2 nd and 3 rd locked-cycle test for the Konttijärvi sample and was high for the Ahmavaara sample. It should be noted that cycle stability was only reached after the 3 rd cycle of a 6 cycle test. Due to the difficulties faced during the locked-cycle program, results obtained do not reflect true recoveries and grades that may have otherwise have been attained. It is suspected that sample aging or oxidation of the samples was responsible for the poor performance. Table 24.4 summarizes the tests results
 
Table 24.4: Locked- cycle Program Test Results

 	 	Combined Concentrate Grade	 	% Distribution	 	
LCT	 	wt
%	 	% Cu	 	 %
Ni	 	Pd g/t	 	Pt g/t	 	Au
g/t	 	% S	 	%
MgO	 	Cu	 	Ni	 	Pd	 	Pt	 	Au	 	S	 	
Ahmavaara	 	2.97	 	8.76	 	3.11	 	49.17	 	9.60	 	5.40	 	29.86	 	3.30	 	86.14	 	74.30	 	76.07	 	59.20	 	82.00	 	67.64	 	
Konttijärvi 1	 	2.13	 	7.95	 	1.49	 	65.22	 	15.63	 	5.62	 	14.82	 	13.90	 	81.51	 	54.37	 	66.65	 	54.21	 	56.92	 	78.33	 	
Konttijärvi 2	 	0.77	 	16.67	 	3.28	 	136.80	 	26.96	 	12.42	 	26.68	 	3.10	 	68.12	 	50.00	 	53.89	 	36.74	 	63.08	 	52.34	 	
 
It was hoped that, for this phase of testwork, reagent optimization would have been established; however, the results and observations from the three locked-cycle tests indicate that more bench testing still needs to be carried out on both materials in order to achieve reagent optimization.
 
High-Pressure Grinding Roll Crusher
 
A short program of testwork was carried out on a small high-pressure grinding roll (HPGR) crusher unit at SGS Lakefield. The results of the tests were then sent to KHD, who developed a preliminary flow sheet, equipment sizing and equipment costs for the circuit.
 
156

 

Figure 24-18 Process Flow Sheet Used for LCT on Ore
 
157

 
25.0        CAPACITY TRADE-OFF EVALUATION
 
25.1        INTRODUCTION
 
As part of the Scoping Study, GFAP requested that Aker Kvaerner carry out a capacity trade-off study to evaluate 5 and 7.5 Mt/a operations to pick the optimum capacity for executing the rest of the study. This evaluation was carried out at the beginning of the Scoping Study in mid November using the preliminary estimates that were available at that time. These values are marginally different to those developed at the end of the Scoping Study.
 
The preliminary process plant designs were produced based on producing a bulk concentrate using the Aker Kvaerner flow sheet. Then, order-of-magnitude capital and operating cost estimates were developed for each option.
 
P&E developed preliminary mine production plans for each option.
 
The above information was then used as a basis for producing preliminary financial models for each case so that they could be compared.
 
25.2        CAPITAL COST ESTIMATES
 
The capital cost estimates for the two cases are shown in Table 25.1.
 
Table 25.1:  Capital Cost Estimates

 	 	5 Mt/a (MUSD)	 	7.5 Mt/a (MUSD)	 	
Mining	 	11.8	 	13.8	 	
Process Facilities	 	99.5	 	126.8	 	
Infrastructure	 	101.4	 	109.5	 	
Indirect Costs	 	52.3	 	60.4	 	
Contingency	 	53.0	 	62.1	 	
Sub-total	 	317.9	 	372.8	 	
Owner’s Costs	 	32.2	 	32.2	 	
Total Project Cost	.	350.1	 	404.9	 	

25.3        MINE PRODUCTION SCHEDULES AND OPERATING COSTS
 
The mine production schedules prepared by P&E for this evaluation are presented in Table and 25.3.
 
158

 
Table 25.2 :  5 Mt/a Mine Production Schedule

 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	OVER	 	 	 	STRIP/	 	 	 	
YEAR	 	PIT	 	Mill Feed	 	Ni	 	Cu	 	Pt	 	Pd	 	Au	 	NSR	 	WASTE	 	BURDEN	 	TOTAL	 	WASTE	 	TONNES	 	
 	 	(t)	 	(%)	 	(%)	 	(g/t)	 	(g/t)	 	(g/t)	 	USD/t	 	(t)	 	(t)	 	(t)	 	Ratio	 	Per Day	 	 	 	
1	 	Kont-Ahma	 	5 000 000	 	0.079	 	0.113	 	0.399	 	1.381	 	0.105	 	26.28	 	6 898 721	 	5 601 279	 	17 500 000	 	2.50	 	50 000	 	
2	 	Kont-Ahma	 	5 000 000	 	0.074	 	0.122	 	0.381	 	1.344	 	0.103	 	25.40	 	6 853 226	 	5 646 774	 	17 500 000	 	2.50	 	50 000	 	
3	 	Kont-Ahma	 	5 000 000	 	0.078	 	0.170	 	0.343	 	1.328	 	0.137	 	25.56	 	12 500 000	 	 	 	17 500 000	 	2.50	 	50 000	 	
4	 	Ahmavarra	 	5 000 000	 	0.083	 	0.182	 	0.251	 	1.088	 	0.111	 	22.09	 	12 500 000	 	 	 	17 500 000	 	2.50	 	50 000	 	
5	 	Ahmavarra	 	5 000 000	 	0.092	 	0.196	 	0.266	 	1.124	 	0.124	 	24.14	 	12 500 000	 	 	 	17 500 000	 	2.50	 	50 000	 	
6	 	Ahmavarra	 	5 000 000	 	0.098	 	0.215	 	0.268	 	1.180	 	0.128	 	25.83	 	10 750 000	 	 	 	15 750 000	 	2.15	 	45 000	 	
7	 	Ahmavarra	 	5 000 000	 	0.102	 	0.225	 	0.263	 	1.231	 	0.121	 	26.67	 	5 500 000	 	 	 	10 500 000	 	1.10	 	30 000	 	
8	 	Ahmavarra	 	5 000 000	 	0.101	 	0.239	 	0.218	 	1.161	 	0.119	 	25.66	 	5 500 000	 	 	 	10 500 000	 	1.10	 	30 000	 	
9	 	Ahmavarra	 	5 000 000	 	0.137	 	0.311	 	0.243	 	1.266	 	0.144	 	33.10	 	2 000 000	 	 	 	7 000 000	 	0.40	 	20 000	 	
10	 	Ahmavarra	 	5 000 000	 	0.127	 	0.331	 	0.274	 	1.390	 	0.190	 	34.56	 	775 000	 	 	 	5 775 000	 	0.16	 	16 500	 	
11	 	Ahmavarra	 	1 624 716	 	0.150	 	0.407	 	0.271	 	1.619	 	0.270	 	41.54	 	43 537	 	 	 	1 668 254	 	0.03	 	14 634	 	
TOTAL	 	 	 	51 624 716	 	0.099	 	0.217	 	0.290	 	1.261	 	0.133	 	27.39	 	75 820 485	 	11 248 053	 	138 693 254	 	1.69	 	 	 	
 
Note: Kont-Ahma indicates ore from both Konttijärvi and Ahmavaara are treated during the year.
 
159

 
Table 25.3  “7.5 Mt/a Production Schedule

 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	 	OVER	 	 	 	STRIP/	 	 	 	
YEAR	 	PIT	 	MILL FEED	 	Ni	 	Cu	 	Pt	 	Pd	 	Au	 	NSR	 	WASTE	 	BURDEN	 	TOTAL	 	WASTE	 	TONNES	 	
 	 	 	 	(t)	 	(%)	 	(%)	 	(g/t)	 	(g/t)	 	(g/t)	 	USD/t	 	(t)	 	(t)	 	(t)	 	Ratio	 	Per Day	 	
1	 	Kont-Ahma	 	7 500 000	 	0.073	 	0.108	 	0.360	 	1.250	 	0.095	 	23.74	 	9 911 731	 	8 838 270	 	26 250 000	 	2.50	 	75 000	 	
2	 	Kont-Ahma	 	7 500 000	 	0.065	 	0.110	 	0.345	 	1.208	 	0.089	 	22.39	 	8 591 256	 	10 158 744	 	26 250 000	 	2.50	 	75 000	 	
3	 	Kont-Ahma	 	7 500 000	 	0.073	 	0.161	 	0.341	 	1.302	 	0.128	 	24.58	 	18 750 001	 	 	 	26 250 000	 	2.50	 	75 000	 	
4	 	Ahmavarra	 	7 500 000	 	0.080	 	0.179	 	0.250	 	1.064	 	0.116	 	21.63	 	18 750 000	 	 	 	26 250 000	 	2.50	 	75 000	 	
5	 	Ahmavarra	 	7 500 000	 	0.084	 	0.189	 	0.260	 	1.098	 	0.130	 	22.84	 	18 750 000	 	 	 	26 250 000	 	2.50	 	75 000	 	
6	 	Ahmavarra	 	7 500 000	 	0.083	 	0.199	 	0.240	 	1.073	 	0.119	 	22.43	 	13 500 000	 	 	 	21 000 000	 	1.80	 	60 000	 	
7	 	Ahmavarra	 	7 500 000	 	0.082	 	0.205	 	0.228	 	1.076	 	0.108	 	22.16	 	6 500 000	 	 	 	14 000 000	 	0.87	 	40 000	 	
8	 	Ahmavarra	 	7 500 000	 	0.083	 	0.217	 	0.185	 	0.982	 	0.111	 	21.22	 	5 625 000	 	 	 	13 125 000	 	0.75	 	37 500	 	
9	 	Ahmavarra	 	7 500 000	 	0.106	 	0.247	 	0.214	 	1.102	 	0.120	 	26.00	 	4 750 000	 	 	 	12 250 000	 	0.63	 	35 000	 	
10	 	Ahmavarra	 	7 500 000	 	0.121	 	0.322	 	0.266	 	1.374	 	0.193	 	33.43	 	968 995	 	 	 	8 468 994	 	0.13	 	24 197	 	
11	 	Ahmavarra	 	722 687	 	0.137	 	0.405	 	0.252	 	1.478	 	0.251	 	38.62	 	 	 	 	 	722 687	 	 	 	21 900	 	
TOTAL	 	 	 	75 722 687	 	0.086	 	0.196	 	0.269	 	1.156	 	0.122	 	24.18	 	106 096 981	 	18 997 014	 	200 816 682	 	1.65	 	 	 	
 
Note: Kont-Ahma indicates ore from both Konttijärvi and Ahmavaara are treated during the year.
 
160

 
The mine operating costs are as follows:

 	 	5 Mt/a	 	7.5 Mt/a	 	
 	 	 	 	 	 	
Mineralized material/Waste	 	USD $1.24/t	 	USD $1.19/t	 	
Overburden	 	USD $0.93/t	 	USD $0.89/t	 	
 
25.4        PROCESS OPERATING COST ESTIMATES
 
The process operating cost estimates for the two cases are shown in Table 25.4.
 
Table 25.4  Process Operating Cost Estimates

 	 	5 Mt/a (USD/a)	 	7.5 Mt/a (USD/a)	 	
 	 	 	 	 	 	
Reagents and Consumables	 	12.2	 	16.6	 	
Labor	 	2.8	 	2.8	 	
Electrical Power	 	8.5	 	12.0	 	
Maintenance Supplies	 	3.0	 	3.8	 	
Contingency	 	4.0	 	5.3	 	
Total	 	30.5	 	40.6	 	
 
25.5        FINANCIAL ANALYSIS
 
Preliminary financial models on a pre-tax basis were developed for each case to enable them to be compared. For each case, rates of return were developed for two different sets of metal prices:
 
•                     18-month trailing average
 
•                     24-month trailing average.
 
The metal prices used are shown in Table 25.5.
 
The Internal Rate of Return (IRR) for the two cases is shown in Table 25.6.
 
Table 25.5:Metal Prices

Metals	 	18-Month Trailing (USD)	 	24-Month Trailing (USD)	 	
Copper	 	5,225.9 /t	 	4,718.7/t	 	
Nickel	 	1,8081.0/t	 	1,7154.0/t	 	
Platinum	 	1,027/oz	 	984/oz	 	
Palladium	 	275/oz	 	247/oz	 	
Gold	 	526/oz	 	502/oz	 	
Rhodium	 	4,000/oz	 	4,000/oz	 	
 
Table 25.6: Capacity Trade-off IRR

 	 	24-month Trailing average	 	18-month Trailing average	 	
5 Mt/a	 	6.7	%	11.3	%	
7.5 Mt/a	 	9.8	%	13.6	%	

161

 
25.6        COMPARISON OF THE OPTIONS
 
The 7.5 Mt/a option showed a significantly better IRR and only a modest increase in capital cost, so the decision was made to select this as the plant capacity for the Scoping Study.
 
162

 
26.0        PROCESS DESIGN
 
26.1        BASE CASE – SAG MILL
 
For the purposes of this Scoping Study, Aker Kvaerner has evaluated the following circuit at a throughput of 7.5 Mt/a.
 
The circuit is presented in Figure 26-1 and consists of:
 
•          primary crushing
•          SAG milling
•          ball milling
•          copper flotation
•          copper concentrate column cleaning
•          copper/nickel rougher and scavenger flotation
•          copper/nickel concentrate regrinding
•          copper/nickel concentrate cleaning
•          bulk concentrate thickening and filtration.
 
The preliminary plant design is based on the design criteria derived in the recent testwork carried out by SGS Lakefield. The main design criteria are as follows:

•	Annual feed rate	 	7.5 Mt/a	
•	Feed composition	 	 	
 	•	Copper	 	0.22%	
 	•	Nickel	 	0.09%	
 	•	Platinum	 	0.36 g/t	
 	•	Palladium	 	1.49 g/t	
 	•	Gold	 	0.15 g/t	
 	•	Rhodium	 	0.023 g/t	
•	Plant availability	 	90%	
•	Mass pull to final concentrate	 	2%	
 
A preliminary layout of the concentrator is shown in Figure 26-2
 
The ROM material is transported by trucks to a 62 by 75 in. primary gyratory crusher where it is crushed to minus 165 mm. The ore is then conveyed to a 24 200 t live-crushed Material stockpile from where it is reclaimed by belt feeders and conveyed to the grinding circuit. The grinding circuit consists of a 10.4-m dia. by 6.7-m long, open circuit SAG mill followed by a 7.9-m dia. by 12.2-m long ball mill in closed circuit with a cyclone cluster. A HP 500 cone crusher is included in the grinding circuit to break circulating mill oversize. The feed is ground to 80% passing 75 microns and the cyclone overflow passes directly to the copper flotation circuit.
 
The copper rougher flotation circuit consists of a conditioning tank followed by two 130 m(3) tank cells in rougher/scavenger configuration. The concentrate from both of these two tank cells is high in copper and is combined to form a copper rougher concentrate. The copper rougher concentrate is fed to a cleaning circuit consisting of two column cells in series. The concentrate from the first column goes to the second cleaner column for final cleaning while the first cleaner column tail, together with the second cleaner tail, is sent to the head of the combined Cu/Ni cleaning circuit.
 
The second cleaner copper concentrate is combined with the copper/nickel final concentrate and thickened, filtered and then packed for shipment to a nickel smelter. The tails from the second cleaner column go to the head of the Cu/Ni cleaner circuit.
 
The tails from the copper-rougher-scavenger circuit are sent to two parallel lines in the Cu/Ni rougher/scavenger circuit. Each of these lines comprises six 130-m³ cells in series. Concentrate from all the Cu/Ni rougher and scavenger cells are combined and
 
163

 
sent to three vertimills to reduce the material to a P 80 of approximately 20 microns. The vertimills operate in closed circuit with cyclones and the overflow from these units is sent to the first stage of Cu/Ni cleaning, which is done in a series of eighteen 38-m³ tank cells. Tails from the Cu/Ni scavenger cells are sent to tailings. The concentrate from the first nine cells in the first cleaner train are sent to the second Cu/Ni cleaners while the concentrates from the remaining nine cells are recycled back to the head of the first Cu/Ni cleaner cells.
 
The second Cu/Ni cleaners consist of nine 38-m³ tank cells. Tails from the second Cu/Ni cleaners are returned to the head of the first cleaners. The second Cu/Ni cleaner concentrate is fed to the third Cu/Ni cleaners which consist of two 19-m³ columns in series. The tails from the third Cu/Ni cleaners are sent back to the head of the second cleaners for additional metal recovery. The third Cu/Ni cleaner concentrate is sent to a single fourth 11-m³ column for final upgrading. This concentrate, together with the copper concentrate, constitutes the final concentrate which is fed to a 17 m dia. conventional filter tank before filtering for shipment. The fourth cleaner tail is recycled back to the head of the third Cu/Ni cleaner.
 
26.2        ALTERNATIVE CASE - HIGH-PRESSURE ROLLS CRUSHING
 
In an effort to reduce operating costs, particularly power costs, the option of using a high-pressure rolls crusher in place of the SAG mill in the grinding circuit has been evaluated. This circuit is shown in Figure 26-3.
 
In this instance, the primary crusher is the same size as in the SAG mill case. The material is crushed to 100% passing 165 mm in open circuit in the gyratory crusher. The crushed rock is fed to a double-deck vibrating screen with a lower-deck aperture of 50 mm. The oversize is fed to a HP 800 cone crusher, the product of which discharges back to the same screen as a closed circuit loop.
 
The minus 50-mm material is choke fed via a surge box into the throat of a RP-10-170/140 high-pressure rolls crusher. About 10% of the feed material, which runs through the edges of the rolls, is returned to the rolls crusher feed box for regrinding. The remaining crushed material from the center of the rolls (80% minus 13 mm) constitutes the feed to a 26 ft dia. by 36 ft long ball mill. This mill is equipped with a 11 mm slotted discharge trommel which is used to separate large oversize material in the mill discharge. This oversize material is also returned to the rolls crusher. A belt magnet is used to remove grinding balls so as to protect the rolls.
 
The minus 11-mm material discharges into a pump box whence it is sent to a nest of hydrocyclones. The cyclone overflow feeds the flotation circuit and the cyclone underflow is returned to the ball mill for regrinding.
 
A preliminary costing of the two options (SAG milling and high-pressure rolls crushing) indicates that an annual operating cost saving of USD $1,540,00 can be achieved if the high-pressure rolls crushing option is used. Capital costs would decrease by USD $30,090,000.
 
164

 

 

 
27.0        INFRASTRUCTURE
 
The infrastructure areas for the Suhanko Project consist of the following:
 
•        site development
•        main access road
•        site roads
•        tailings storage and reclaim
•        power supply
•        main substation distribution
•        fuel storage and distribution
•        water supply (fresh, process and fire)
•        sewage treatment
•        port facility
•        communications
•        mobile equipment
•        ancillary buildings
 
The scope and design criteria for the infrastructure areas are primarily based on the work carried out for the 2005 Lycopodium Feasibility Study. However, Aker Kvaerner made a visit to site in November 2006 to review the infrastructure and meet with various local organizations. As a result, some changes were made to parts of the infrastructure that resulted in some cost revisions
 
Capital cost reductions were made to the main access road, site roads, tailings storage and reclaim power supply following discussion with local organizations.
 
The port facility design criteria were revised for larger capacity ships for shipping concentrate. This necessitated the storage building increase in capacity from 20,000 t to 30,000 t.
 
In a meeting held with the Kemi Port Authority, it was agreed that the location of the dock and storage facility would be relocated to another area within the port site. This revised positioning will allow for easier truck access from the main gate at the port to the storage building. Also the new dock location will accommodate larger ocean vessels.
 
167

 
28.0        CAPITAL COST ESTIMATE
 
28.1        CAPITAL COST SUMMARY
 
An order of magnitude capital cost has been developed for the mining and surface facilities for this preliminary economic assessment. This estimate is in fourth quarter 2007 United States Dollars and is subject to the qualifications and exclusions listed below.
 
The capital costs for the mine, concentrator, and infrastructure for these options are summarized in Table 28.1 and are inclusive of the costs up to and including plant commissioning and start-up. Sustaining capital costs and deferred capital costs are excluded from these estimates as they are included in the cash flow model in the year in which they are incurred.
 
The base currency of the capital cost estimate is in United States Dollars. Other currencies used and conversion rates are as follows.

Currency	 	Conversion Rates	
 	 	 	
Canadian Dollar	 	1.00 CAD = USD $0.8900	
 	 	 	
European EURO	 	1.00 EUR = USD $1.2250	
 	 	 	
Australian Dollar	 	1.00 AUD = USD $0.7000	
 
Table 28-1:Suhanko Project – Project Cost Summary

Project Area Description	 	Cost (USD)	 	
Direct Costs	 	 	 	
Process Mechanical Equipment	 	76,157,000	 	
Factored Commodities Process Plant	 	64,738,000	 	
Infrastructure	 	103,543,000	 	
Mining	 	 	 	
Roads	 	515,000	 	
Mine Operations Complex	 	1,194,800	 	
Mine Maintenance Complex	 	8,749,850	 	
Mine Support Facilities	 	1,354,450	 	
Mine Support Equipment	 	3,687,400	 	
Mining Equipment Fleet (lease down payment)	 	6,654,830	 	
Total Direct Costs	 	266,594,000	 	
Indirect Costs	 	 	 	
Project Indirect Costs	 	62,715,600	 	
Subtotal for Contingency	 	329,308,880	 	
Contingency @ 20%	 	65,862,120	 	
Subtotal Direct and Indirect	 	395,171,000	 	
Owner Costs	 	33,621,000	 	
TOTAL PROJECT CAPITAL COST	 	428,792,000	 	
 
168

 
All costs are expressed in third quarter 2007 USD with no allowance for escalation, interest or financing during construction. Budgetary estimates for mining and mining equipment were provided by P&E Mining Consultants. Processing equipment prices were obtained from the 2005 Lycopodium Feasibility Study and escalated to 2006 prices and in-house historical data from similar projects. Commodity pricing for earthwork, concrete, steel and architectural work has been factored from equipment costs. The cost estimate, based on the designs presented in this Scoping Study, has a predicted accuracy level of ±30%.
 
Table 28.2 represents further breakdown for the plant and infrastructure costs.
 
Table 28-2: Plant and Infrastructure Breakdown
 
Aker Kvaerner Option  2 - Bulk Copper-Nickel Concentrate Capital Cost Estimate
 
(USD)

Description	 	Total Cost	 	
 	 	 	 	 	 	 	 	
Process Mechanical Equipment	 	 	 	 	 	
Area 2200	 	Primary Crushing (Feed Preparation)	 	 	 	4,971,810	 	
Area 2230	 	Grinding (Milling)	 	 	 	33,798,420	 	
Area 2300	 	Copper Flotation	 	 	 	4,925,460	 	
Area 2310	 	Bulk Copper/Nickel Flotation	 	 	 	21,985,350	 	
Area 2320	 	Copper Concentrate Dewatering	 	 	 	—	 	
Area 2330	 	Bulk Copper/Nickel Concentrate Dewatering (Handling)	 	 	 	7,607,580	 	
Area 2340	 	Reagents	 	 	 	2,867,520	 	
 	 	 	 	 	 	
 	 	Process Mechanical Equipment Installed	 	76,157,000	 	
Factored Commodities Process Plant	 	 	 	 	 	 	 	
Earthworks and Civil (Structural)	 	2.0%	 	Mechanical	 	1,524,000	 	
Concrete	 	17.0%	 	Mechanical	 	12,947,000	 	
Structural Steel	 	15.0%	 	Mechanical	 	11,424,000	 	
Buildings, Architectural	 	8.0%	 	Mechanical	 	6,093,000	 	
Building Services	 	5.0%	 	Mechanical	 	3,808,000	 	
Electrical	 	12.0%	 	Mechanical	 	9,139,000	 	
Instrumentation	 	6.0%	 	Mechanical	 	4,570,000	 	
Piping	 	18.0%	 	Mechanical	 	13,709,000	 	
Insulation	 	2.0%	 	Mechanical	 	1,524,000	 	
 	 	SUBTOTAL Factored Commodities	 	64,738,000	 	
Infrastructure	 	 	 	 	 	 	 	
Site Development	 	 	 	 	 	8,196,740	 	
Main Access Road	 	 	 	 	 	4,861,600	 	
Site Road	 	 	 	 	 	5,498,140	 	
Tailings Storage & Reclaim	 	 	 	 	 	16,363,610	 	
Power Supply	 	 	 	 	 	12,697,840	 	
Main Substation Distribution	 	 	 	 	 	14,551,840	 	
Fuel Storage & Distribution	 	 	 	 	 	234,840	 	
Water Supply (Fresh, Process & Fire)	 	 	 	 	 	7,809,460	 	
Sewage Treatment	 	 	 	 	 	1,139,180	 	
Port & other costs	 	 	 	 	 	18,878,870	 	
Communication	 	 	 	 	 	1,814,860	 	
Mobile Equipment	 	 	 	 	 	4,549,510	 	
Ancillary Buildings	 	 	 	 	 	6,946,320	 	
 	 	SUBTOTAL Infrastructure	 	 	 	103,543,000	 	

169

 

Mining	 	 	 	 	 	
Roads	 	 	 	515,000	 	
Mine operations complex	 	 	 	1,194,800	 	
Mine maintenance complex	 	 	 	8,749,850	 	
Mine support facilities	 	 	 	1,354,450	 	
Mining Equipment Fleet (Lease down payment)	 	 	 	6,654,830	 	
Mining Support Equipment	 	 	 	3,687,400	 	
 	 	 	 	 	 	
 	 	SUBTOTAL Mining	 	22,157,000	 	
 	 	TOTAL DIRECT COST	 	266,594,000	 	
Indirect Costs	 	 	 	 	 	
Construction Field Indirects	 	 	 	7,340,000	 	
Construction Support	 	 	 	2,450,000	 	
Spare Parts - Process	 	 	 	1,565,600	 	
Spare Parts - Mining	 	 	 	—	 	
Spare Parts - Port & Other	 	 	 	100,000	 	
First Fills and Inventory	 	 	 	3,000,000	 	
Vendor Supervision	 	 	 	750,000	 	
Freight, Transport and Insurance	 	 	 	9,780,000	 	
Taxes, Duties & Permits	 	Excluded	 	—	 	
Basic Engineering	 	 	 	6,120,000	 	
Detail EPCM	 	 	 	29,340,000	 	
Third Party Engineering	 	 	 	—	 	
EPCM Travel and Related Expenses	 	 	 	1,500,000	 	
Start-up/Commissioning Support	 	 	 	770,000	 	
 	 	SUBTOTAL Indirect Cost	 	62,715,600	 	
 	 	 	 	 	 	
 	 	SUBTOTAL for Contingency	 	329,308,880	 	
 	 	 	 	 	 	
Contingency	 	20%	 	65,862,120	 	
 	 	 	 	 	 	
 	 	Subtotal Direct & Indirect Costs	 	395,171,000	 	
Owner’s Costs	 	 	 	 	 	
Owner’s Costs General	 	 	 	33,621,000	 	
Pre-production	 	 	 	 	 	
Spare Parts	 	 	 	 	 	
Fees/Taxes/Duties	 	 	 	 	 	
Other Owner,s Costs	 	 	 	 	 	
 	 	 	 	 	 	
 	 	SUBTOTAL Owner’s Costs	 	33,621,000	 	
 	 	 	 	 	 	
 	 	PROJECT TOTAL	 	428,792,000	 	
 
28.2        BASIS OF ESTIMATE
 
The capital cost estimate includes the following:
 
•             Lease down payment for mining equipment
•             Direct costs of new equipment for the processing facilities
•             Construction materials and installation labor
•             Temporary buildings and services
•             Construction support services
•             Project infrastructure
•             Spare parts
•             Initial fills (inventory)
•             Freight
•             Vendor supervision
•             Owner’s cost
•             Engineering, procurement and construction management
•             Commissioning and start up
•             Contingency.
 
170

 
28.3        DIRECT COST ELEMENTS
 
The direct costs are all the costs associated with permanent facilities. This includes equipment and material costs, as well as construction and installation costs.
 
Process Equipment
 
Equipment pricing is based on the equipment list and process flow diagrams. Budgetary prices were obtained from the 2005 Lycopodium Feasibility Study for comparable major equipment. In-house cost data was used to cost the remaining equipment, tanks, bins and chutes. Costs for installation of equipment are based on unit man-hour requirements.
 
Other direct costs were factored based on the total installed cost of process equipment for such as:
 
•               earthwork/site work
•               concrete
•               structural steel
•               buildings and architectural
•               electrical
•               instrumentation and controls
•               piping
 
28.4        INDIRECT COST ELEMENTS
 
The indirect costs cover all the costs associated with temporary construction facilities and services, construction support, freight, Vendor representatives, spare parts, initial fills and inventory, Owner’s costs, EPCM, commissioning and start-up assistance.
 
Construction Facilities
 
The costs for construction facilities include all temporary facilities, services and operation, site office operations, security buildings and services, construction warehousing and material management, construction power and utilities, site transportation, medical facilities and services, garbage collection and disposal, and surveying.
 
Spare Parts
 
The cost for spare parts is factored based on equipment costs.
 
Initial Fills (Inventory)
 
The estimated cost for initial fills of reagents is based on 3 months of operating requirements. Budget quotations were obtained for reagent pricing.
 
Freight
 
The freight costs are factored and typically include containerized and break bulk shipping and each are respectively divided into ocean freight and inland freight. For imported equipment, the cost of freight and export packing, ex-works to a local port, is included with the cost of the equipment.
 
Freight insurance is included in the Owner’s cost, project insurance.
 
Vendor Representatives
 
The requirement for Vendor representatives to supervise the installation of equipment or to conduct a checkout of the equipment prior to start-up of the equipment, as deemed necessary for equipment guarantees or warranties, has been included in the estimate. Typically, the cost for this item is inclusive of salary and travel.
 
171

 

 	Taxes and Duties	
 	 	
 	Taxes and duties are excluded.	
 	EPCM	
 	EPCM has been calculated using North American engineering costs.	
 	 	
28.5	CONTINGENCY	
 	 	
 	An overall contingency allowance of 20% has been included in the estimate.	
 	 	
28.6	QUALIFICATIONS AND EXCLUSIONS	
 	 	
 	28.6.1  QUALIFICATIONS	
 	 	
 	All construction work will be executed by local contractors.	
 	 	
 	28.6.2  EXCLUSIONS	
 	 	
 	The following items have been excluded from this report.	
 	 	
 	•   Sunk costs	
 	•   Sustaining capital	
 	•   Deferred capital	
 	•    Working capital	
 	•    Financing and interest during construction	
 	•    Additional exploration drilling	
 	•    Escalation	
 	•    Corporate withholding taxes	
 	•    Legal costs	
 	•    Process royalty fees	
 	•    Metallurgical testing costs	
 	•    VAT, Tax and Import duties.	
 	 	
28.7	PIT CAPITAL COST	
 	 	
 	The estimated mining capital cost for 7.5 Mt/a, prepared by P&E Mining Consultants, is presented in Table 28.3.	
 
172

 
Table 28.3: Estimated Pit Capital Cost – 7.5 Mt/a

 	 	 	 	Unit cost	 	 	 	 	 	 	 	
 	 	 	 	(on-site and	 	Extended	 	Spare Parts	 	Rounded	 	
Item	 	No.	 	assembled)	 	Cost	 	Allowance	 	Total	 	
 	 	 	 	(USD)	 	(USD)	 	(USD)	 	(USD)	 	
Direct Mine Equipment	 	 	 	 	 	 	 	 	 	 	 	
Production drills	 	2	 	1,030,000	 	2,060,000	 	206,000	 	2,266,000	 	
Shovel O&K RH120E	 	2	 	4,480,500	 	8,961,000	 	896,100	 	9,857,100	 	
Front end loader Cat 994F	 	1	 	3,759,500	 	3,759,500	 	375,950	 	4,135,450	 	
Rock trucks Cat 785C	 	14	 	2,858,250	 	40,015,500	 	1,545,000	 	41,560,500	 	
Tractor Cat D10	 	2	 	1,308,100	 	2,616,200	 	183,130	 	2,799,330	 	
Dozer Cat 854	 	2	 	1,823,100	 	3,646,200	 	255,650	 	3,901,850	 	
Grader Cat 16M	 	2	 	921,850	 	1,843,700	 	184,370	 	2,028,070	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	66,548,300	 	
Mine Operations - Support Equipment	 	 	 	 	 	 	 	 	 	 	 	
Water truck	 	1	 	875,500	 	875,500	 	25,750	 	901,250	 	
Fuel truck	 	1	 	824,000	 	824,000	 	25,750	 	849,750	 	
Mechanics field truck	 	2	 	824,000	 	1,648,000	 	25,750	 	1,673,750	 	
Pit pump and pipeline	 	2	 	128,750	 	257,500	 	5,150	 	262,650	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	3,687,400	 	
Mine Operations - Indirect Capital Expenditures	 	 	 	 	 	 	 	 	 	 	 	
Mine office	 	1	 	257,500	 	257,500	 	0	 	257,500	 	
Mine dry	 	1	 	309,000	 	309,000	 	0	 	309,000	 	
Mine dispatch incl. monitor system	 	1	 	309,000	 	309,000	 	10,300	 	319,300	 	
Pick-up trucks	 	4	 	72,100	 	288,400	 	20,600	 	309,000	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	1,194,800	 	
Mine Maintenance Complex (6,000 m 2 )	 	 	 	 	 	 	 	 	 	 	 	
Foundations and reinforced floor	 	1	 	1,442,000	 	1,442,000	 	0	 	1,442,000	 	
Shop building	 	1	 	5,562,000	 	5,562,000	 	0	 	5,562,000	 	
Heat and power	 	1	 	437,750	 	437,750	 	20,600	 	458,350	 	
Shop equipment	 	1	 	1,287,500	 	1,287,500	 	0	 	1,287,500	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	8,749,850	 	
Equipment refueling	 	 	 	 	 	 	 	 	 	 	 	
Main diesel storage tank and piping	 	1	 	257,500	 	257,500	 	0	 	257,500	 	
Refuel station at shop	 	1	 	154,500	 	154,500	 	0	 	154,500	 	
Refuel station at pit	 	1	 	360,500	 	360,500	 	10,300	 	370,800	 	
Gasoline re-fuelling station	 	1	 	25,750	 	25,750	 	0	 	25,750	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	808,550	 	
Geology and Engineering	 	 	 	 	 	 	 	 	 	 	 	
Office	 	1	 	128,750	 	128,750	 	 	 	128,750	 	
Equipment and software	 	1	 	154,500	 	154,500	 	 	 	154,500	 	
Pick-up trucks	 	2	 	72,100	 	144,200	 	 	 	144,200	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	427,450	 	
First Aid and Mine Rescue	 	 	 	 	 	 	 	 	 	 	 	
Nursing station	 	1	 	10,300	 	10,300	 	 	 	10,300	 	
Ambulance	 	1	 	97,850	 	97,850	 	 	 	97,850	 	
Training room and equipment	 	1	 	10,300	 	10,300	 	 	 	10,300	 	
 	 	 	 	 	 	 	 	SUBTOTAL	 	118,450	 	
 	 	 	 	 	 	 	 	TOTAL	 	81,534,800	 	
 
173

 
29.0        OPERATING COST ESTIMATE
 
29.1        MINING
 
P&E Mining Consultants estimated the pit operating costs based on:
 
The estimated labour requirements, and labour costs used in other studies for proposed mine projects in Finland as well as labour cost information provided by Aker Kvaerner.
 
Fuel consumption ranges listed in the Caterpillar handbook version 35 and information from other vendors, and a cost allowance of CAD 1.50/L (USD $1.30/L) for diesel fuel.
 
Typical drilling and blasting material consumable consumptions.
 
Maintenance consumables including parts, lubes, ground engagement tools, and tires taking into account increased demand and prices.
 
The mine being operated by the Owner’s forces.
 
As shown in Table 29.1, the estimated operating cost for 7.5 Mt/a is USD $1.84/t rock. The estimated cost per tonne of ore based on an overall stripping ratio of 1.8 is USD $3.31/t processed.
 
Table 29.1  Estimated Operating Cost – 7.5 Mt/a

Summary	 	Rock 
(USD)	 	Milled 
(USD)	 	
Direct labour	 	0.27	 	0.49	 	
Indirect labour	 	0.16	 	0.29	 	
Diesel Fuel	 	0.68	 	1.22	 	
Drill Steel & Bits	 	0.02	 	0.04	 	
Explosives	 	0.23	 	0.41	 	
Maintenance Consumables	 	0.48	 	0.86	 	
Total	 	1.84	 	3.31	 	
 
29.2        PROCESS PLANT
 
Preliminary process plant operating cost estimates have been developed in third quarter 2006 USD for the Konttijärvi and Ahmavaara ores. The operating costs were estimated by developing costs for labour, reagents and consumables, electrical power and maintenance supplies.
 
The operating costs are summarized below in Table 29.2.
 
174

 
Table 29.2: Process Plant Operating Cost Summary

Cost Area	 	Ahmavaara	 	Konttijarvi	 	
 	 	Annual Costs	 	Costs	 	Annual Costs	 	Costs	 	
 	 	(USD/a)	 	(USD/t ore)	 	(USD/a)	 	(USD/t ore)	 	
Labour	 	4,770,810	 	0.64	 	4,770,810	 	0.64	 	
Reagents and Consumables	 	17,652,862	 	2.35	 	16,421,145	 	2.19	 	
Assay laboratory	 	257,500	 	0.03	 	257,500	 	0.03	 	
Electrical Power	 	13,262,833	 	1.77	 	13,262,833	 	1.77	 	
Maintenance Supplies	 	4,194,459	 	0.56	 	4,194,459	 	0.56	 	
Contigency 10 %	 	4,013,846	 	0.54	 	3,890,675	 	0.52	 	
TOTAL	 	44,152,310	 	5.89	 	42,797,421	 	5.71	 	
 
29.2.1     LABOUR
 
Labour costs are presented in Table 29.3 and were developed by estimating the number of operators and maintenance personnel required for each option and then applying payroll costs for each position. Payroll costs were provided by GFAP for three positions and the cost of the remaining positions were factored from these. The labour costs were the same for both Konttijärvi and Ahmavaara.
 
Table 29.3: Bulk Concentrate for Sale - Labour

Process Labour	 	NumberRequired	 	Shift	 	Cost/Annum	 	Total Labour Costs	 	
Plant Manager	 	1	 	 	 	181,305	 	181,305	 	
Senior Metallurgist	 	1	 	 	 	102,872	 	102,872	 	
Project Metallurgist	 	1	 	 	 	82,014	 	82,014	 	
Operations General Foreman	 	1	 	 	 	82,014	 	82,014	 	
Shift Foreman	 	5	 	x	 	68,135	 	340,673	 	
Crushing Operators	 	5	 	x	 	42,976	 	214,879	 	
Control Room Operators	 	5	 	x	 	59,302	 	296,511	 	
Grinding Operators	 	5	 	x	 	56,745	 	283,724	 	
Flotation Operators	 	5	 	x	 	59,302	 	296,511	 	
Thickener/Filter Operators	 	5	 	x	 	54,296	 	271,482	 	
Reagent Operators	 	5	 	x	 	40,919	 	204,594	 	
Loader operators	 	5	 	x	 	42,955	 	214,776	 	
Holiday Relief (labourers)	 	8	 	 	 	40,919	 	327,350	 	
Chief Assayer	 	1	 	 	 	63,257	 	63,257	 	
Assay Lab Technicians	 	10	 	x	 	40,919	 	409,188	 	
Metallurgical Technicians	 	1	 	 	 	33,679	 	33,679	 	
Maintenance Planners	 	2	 	 	 	50,430	 	100,860	 	
Mechanical Engineer	 	1	 	 	 	50,430	 	50,430	 	
Maintenance General Foreman	 	1	 	 	 	50,430	 	50,430	 	
Chief Electrician	 	1	 	 	 	50,430	 	50,430	 	
Shift Electricians	 	6	 	x	 	42,435	 	254,610	 	
Instrumentation Technicians	 	2	 	 	 	42,435	 	84,870	 	
Maintenance Foreman	 	2	 	 	 	59,341	 	118,683	 	
Mill Wrights	 	12	 	x	 	50,430	 	605,158	 	
Mill Trainer	 	1	 	 	 	50,430	 	50,430	 	
TOTAL	 	92	 	 	 	 	 	4,770,810	 	
 
29.2.2     REAGENTS AND CONSUMABLES
 
Reagents and consumable costs for Konttijärvi and Ahmavaara are presented in Table 29.4 and Table 29.5. The reagent costs were developed based on the consumptions from the latest testwork and the consumptions were based on using in-house data for liners and grinding media. The unit costs for the concentrator reagents and consumables were taken from the 2005 Lycopodium Feasibility Study.
 
175

 
Table 29.4:  Reagents and Consumables - Konttijärvi

 	 	 	 	 	 	Annual	 	
 	 	Consumption	 	Cost	 	Cost	 	
 	 	(t)	 	(USD/t)	 	(USD/a)	 	
 	 	 	 	 	 	 	 	
Grinding Media SAG & Ball Mills	 	5616.7	 	755	 	4,367,857	 	
Grinding Media (Vetimills)	 	702.1	 	880	 	636,376	 	
Mill & Crusher Liners	 	 	 	 	 	3,935,195	 	
Lime	 	82.5	 	220	 	18,695	 	
Coillector	 	2065.0	 	960	 	2,041,849	 	
Frother	 	289.1	 	1500	 	446,654	 	
Depressants	 	3304.0	 	1014	 	3,450,727	 	
Promoter Danafloat	 	1032.5	 	1415	 	1,504,800	 	
Floculents	 	5.7	 	3200	 	18,648	 	
TOTAL	 	 	 	 	 	16,421,145	 	
 
Table 29.5:  Reagents and Consumables - Ahmavaara

 	 	 	 	 	 	Annual	 	
 	 	Consumption	 	Cost	 	Cost	 	
 	 	(t)	 	(USD/t)	 	(USD/a)	 	
 	 	 	 	 	 	 	 	
Grinding Media SAG & Ball Mills	 	5616.7	 	755	 	4,367,857	 	
Grinding Media (Vetimills)	 	702.1	 	880	 	636,376	 	
Mill & Crusher Liners	 	 	 	 	 	3,935,195	 	
Lime	 	82.5	 	220	 	18,695	 	
Coillector	 	2312.8	 	960	 	2,286,872	 	
Frother	 	330.4	 	1500	 	510,462	 	
Depressants	 	4130.0	 	1014	 	4,313,408	 	
Promoter Danafloat	 	1073.8	 	1415	 	1,564,992	 	
Floculents	 	5.7	 	3200	 	18,648	 	
TOTAL	 	 	 	 	 	17,652,862	 	
 
29.2.3     ELECTRICAL POWER
 
Electrical power costs, shown in Table 29.6, were developed based on the power consumption for each item of equipment from the equipment list and then applying factors for utilization and power draw. The unit cost for electrical power was provided by GFAP. The electrical power costs were the same from both Konttijärvi and Ahmavaara.
 
Table 29-6:  Power Costs

 	 	Connected	 	Drawn	 	Annual	 	Annual Power	 	Annual	 	
Area	 	Load	 	Load	 	Utilization	 	Consumption	 	Cost	 	
 	 	(kW)	 	(kW)	 	(%)	 	(MWh)	 	(USD/a)	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Primary Crushing & Storage Area	 	835.1	 	709.8	 	85	 	5285	 	266,743	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Grinding	 	31859.0	 	27080.2	 	85	 	201639	 	10,176,729	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Copper Flotation	 	839.0	 	713.2	 	85	 	5310	 	268,020	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Bulk Cu-Ni Flotation	 	7579.0	 	6442.2	 	85	 	47969	 	2,420,976	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Flotation Concentrate Handling	 	127.0	 	108.0	 	85	 	804	 	40,586	 	
 	 	 	 	 	 	 	 	 	 	 	 	
Reagents	 	281.0	 	238.9	 	85	 	1779	 	89,779	 	
 	 	 	 	 	 	 	 	 	 	 	 	
TOTAL	 	41520.1	 	35292.3	 	 	 	262786	 	13,262,833	 	
TOTAL	 	 	 	 	 	 	 	 	 	13,262,833	 	
USD/t ore	 	 	 	 	 	 	 	 	 	1.77	 	
 
176

 
29.2.4     MAINTENANCE SUPPLIES
 
Maintenance supplies were developed by applying a factor to the total plant direct costs.
 
29.2.5     CONTINGENCY
 
A contingency of 10% has been applied to all the other costs.
 
29.3         GENERAL AND ADMINISTRATION (G&A)
 
The G&A costs cover for management, legal, administration, safety, security and other related expenses. The annual cost is USD $5.15 M.
 
177

 
30.0        FINANCIAL ANALYSIS
 
30.1        INTRODUCTION
 
A preliminary financial model for the project has been developed for comparison of the different options on a pre-tax basis for the first ten years of the mine life. A more detailed financial model will be required at a later stage of the project to determine the true financial return.
 
30.2        METAL RECOVERIES AND MASS PULLS
 
The metal recoveries and concentrate mass pulls used in the analysis for Konttijärvi and Ahmavaara are shown in Table 30.1.
 
Table 30.1:  Metal Recoveries and Mass Pulls

 	 	Konttijärvi	 	Ahmavaara	 	
Mass Pull	 	1.40	 	1.40	 	
Copper	 	73.0	 	75.0	 	
Nickel	 	65.0	 	65.0	 	
Platinum	 	65.0	 	67.0	 	
Palladium	 	69.0	 	75.0	 	
Gold	 	75.0	 	77.0	 	
 
It should be noted that the recoveries used in this study are derived from both the 2005 Lycopodium Feasibility Study and the recent tests conducted at SGS Lakefield. These latter tests were constrained by the availability of fresh ore, which led to lower recoveries than those obtained previously. It is postulated that some, or all, of the sample used became oxidized during storage and that this oxidation interfered with sulphide flotation. For this reason, emphasis was placed on the 2005 Lycopodium Feasibility Study work and some of the locked cycle work, particularly on the Konttijärvi ore, was ignored. It is recommended that a new sample of fresh ore be obtained to confirm the estimated recoveries used in this study and to optimize flotation conditions that will be used in the concentrator.
 
30.3        INFORMATION FROM OWNER
 
The following information that was used in the financial models was provided by GFAP: Table 30.2 represents metal prices for plant capacity 7.5 Mt/a.
 
Table 30-2
Metal Prices for Plant Capacity 7.5 M/ta

 	 	18 Month Trailing	 	24 Month Trailing	 	
Metals	 	(USD)	 	(USD)	 	
Copper	 	7,206/t	 	6,560/t	 	
Nickel	 	33,651/t	 	28,671/t	 	
Platinum	 	1,222/oz	 	1,167/oz	 	
Palladium	 	344/oz	 	325/oz	 	
Gold	 	644/oz	 	613/oz	 	
 
178

 
The 18 and 24 month trailing average prices are dated as of September 30, 2007.
 
•                   Bulk concentrator smelter payable

Copper	 	85%	 	
Nickel	 	90%	 	
Platinum	 	89%	 	
Palladium	 	91%	 	
Gold	 	85%	 	
Rhodium	 	70%	 	
 
•       Bulk concentrate smelting costs USD $215/t of concentrates
 
•       Bulk concentrate refining costs (USD)

Copper	 	661.5/t	 	
Nickel	 	1,168.7/t	 	
Palladium	 	18.7/oz	 	
Platinum	 	16.7/oz	 	
Gold	 	14.4/oz	 	

•      G&A costs	 	USD $5.15 M/a	 	
•      Rebate on road construction	 	USD $3 M	 	
 
30.4        INFORMATION FROM P&E
 
The information in Table 30.3 that was used in the financial models is a summary of what was provided by P&E Mining Consultants.
 
Table 30.3: Ore Grades & Tonnages

 	 	 	 	Konttijärvi	 	Ahmavaara	 	
Copper	 	%	 	0.114	 	0.205	 	
Nickel	 	%	 	0.072	 	0.087	 	
Platinum	 	g/t	 	0.386	 	0.230	 	
Palladium	 	g/t	 	1.346	 	1.088	 	
Gold	 	g/t	 	0.108	 	0.127	 	
Strip ratio	 	 	 	1.4	 	1.4	 	
Total ore	 	t	 	17 493 661	 	57 506 339	 	
 
A more detailed production schedule is shown in Section 22.0, Table 22.6 and this data has been used in the financial model. The above information relates to the first 10 years of operation and there are additional mineable resources as shown in Table 22.5 that could extend the life of the project and have a modest positive impact on the financial return of the project.

Mining Costs	 	 	
 	 	 	
Ore/Waste	USD $1.84/t	 	
Overburden	USD $1.23/t	 	
Mining Equipment Lease Cost /Down Payment	10%	
Balance	 	Over 5 years	
 	 	 	
Sustaining capital	 	5% of Equipment cost/a	
 
179

 
30.5        INFORMATION FROM FEASIBILITY STUDY
 
The information, in Table 30.4, used in the financial models, was taken from the 2005 Lycopodium Feasibility Study.
 
Table 30.4: Information from Feasibility Study

Transportation costs	 	 	
•	Land	USD $13.9/t of concentrate	
•	Sea	USD $35/t of concentrate	
•	Insurance	0.15% of metal value	
Working capital	 	4.6 months	
Tailings sustaining cost	USD $11.1 M	
Rehabilitation	 	USD $24.976 M	
Year 1 throughput	 	87.5%	
Year 1 mine OPEX	 	100%	
Year 1 concentrator OPEX	100%	

30.6        FINANCIAL RESULTS
 
The results of the financial models for the different options are shown in Table 30.5.
 
Table 30-5 : Financial Results Pre-tax Basis, First 10 years of Production

 	 	18 Month Trailing	 	24 Month Trailing	 	
IRR	 	23.9	%	19.1	%	
NPV @ 7.5% (USD)	 	501,108,752	 	337,961,416	 	
NPV @ 10% (USD)	 	359,875,902	 	225,041,891	 	
NPV @ 12.5% (USD)	 	252,143,456	 	139,727,512	 	
 
30.7      SENSITIVITY ANALYSIS
 
A sensitivity analysis has been carried out based on the preliminary financial model for the 18-month trailing average metal prices. The factors that have been investigated are:
 
•       Capital cost
•       Operating cost
•       Commodity prices
•       Copper price
•       Nickel price
•       Platinum price
•       Palladium price
 
The results of the analysis showing the impact on the IRR for the project for variations between 80% and 120% of the values for each of the above factors are shown in Table 30.6 and Figure 30-1. It should be noted in Figure 30-1 that the lines for copper, platinum, and palladium are virtually on top of one another.
 
180

 

Aker Kvaerner	 	179400	
 	 	Gold Fields Arctic Platinum OY
Suhanko Project – Scoping Study, Revision 2	
 	 	Section 12 – Financial Analysis	
 
Table 12-6
Sensitivity Analysis

 	 	80%	 	85%	 	90%	 	95%	 	100%	 	105%	 	110%	 	115%	 	120%	 	
Capital Cost	 	29.2	 	27.7	 	26.3	 	25.1	 	23.9	 	22.9	 	21.9	 	20.9	 	20.0	 	
Copper	 	22.2	 	22.6	 	23.1	 	23.5	 	23.9	 	24.4	 	24.8	 	25.2	 	25.6	 	
Nickel	 	20.1	 	21.1	 	22.0	 	23.0	 	23.9	 	24.9	 	25.8	 	26.7	 	27.6	 	
Platinum	 	22.1	 	22.6	 	23.0	 	23.5	 	23.9	 	24.4	 	24.9	 	25.3	 	25.8	 	
Palladium	 	22.0	 	22.5	 	23.0	 	23.5	 	23.9	 	24.4	 	24.9	 	25.4	 	25.9	 	
Opex	 	27.0	 	26.3	 	25.5	 	24.7	 	23.9	 	23.2	 	22.4	 	21.6	 	20.8	 	
Commodities	 	13.5	 	16.3	 	18.9	 	21.5	 	23.9	 	26.3	 	28.6	 	30.8	 	33.0	 	

 
The results show that the project is most sensitive to the variation in the collective prices for all the commodities with the IRR varying from 13.5% to 33.0%. The capital and operating costs are the next most sensitive factors with the IRR varying from 20.0% to 29.2% for capital cost and 20.8% to 27.0% for operating cost. The nickel price is the most sensitive of the individual metal prices with the IRR varying from 20.1% to 27.6%. The sensitivities for copper, platinum and palladium prices are virtually identical with the IRR varying from 22.0% to 25.9%.
 
181

 
31.0        RISK ASSESSMENT
 
A risk assessment has been carried out on the Suhanko Project by considering the following factors:
 
•               Political/social
•               Permitting/approvals
•           Geological complexity
•           Commodity prices/exchange rates
•               Mineral resources
•           Mining
•           Metallurgy
•           Capital cost/operating cost
•           Saleable concentrate
 
Each of these factors is discussed below.
 
31.1        POLITICAL/SOCIAL FACTORS
 
Finland is a very stable country politically and there is an extensive history of mining in the country, so there should be a good supply of skilled people to work on the construction and operation of the project. The project location is only 60 km south of the town of Rovaniemi which has a population of 34,400. Rovaniemi is the capital of Lapland and is a major regional centre and is serviced by rail, road and air. The port of Kemi is 120 km south of Rovaniemi and it is kept open all year round. Consequently, the risk associated with political/social factors should be low.
 
31.2        PERMITTING/APPROVALS
 
The Mining Lease for Suhanko has gone through the approval process and is only awaiting certification. The Environmental Permit for Suhanko was granted December 7, 2005. There are two outstanding appeals against the permit but are only concerned about the amount of compensation to local land holders. The risks associated with permitting/approvals are negligible.
 
31.3        GEOLOGICAL COMPLEXITY
 
Both the Konttijärvi and Ahmavaara deposits are unique mineralizing systems in a geologically complex environment for which no experience and knowledge base is available. Until a sufficient amount of mining has taken place to permit reconciliation of the resource model with actual production, the validity of any resource model cannot be proved. It should also be noted that a definitive mineralization model has not yet been established for either of the deposits.
 
For this reason, the NAP resource estimation strategy emphasized the conservative use of restrictive search ellipses combined with kriging efficiency for classification of the resources.
 
More importantly, the Suhanko Project consists of two distinct orebodies, Ahmavaara and Konttijärvi. While they have many similarities, they exhibit some unique characteristics, example, higher sulphide content in Ahmavaara and Konttijärvi together are comprised of 11 different ore bearing lithologies that exhibit considerable variances in MgO, talc content and hardness.
 
There is, however, sufficient information available regarding the deposits to make preliminary financial decisions, and the approach taken in terms of diamond drilling and sampling is suitable for resource estimation applied to open-pit resources at the
 
182

 
scoping level. The risk associated with geological complexity is considered to be moderate to high due to the ore variability.
 
31.4        COMMODITY PRICES/EXCHANGE RATES
 
The financial return for the Suhanko Project is very dependent upon the prices of copper, nickel, platinum, palladium and, to a lesser extent, on gold. The prices of these metals are currently all high and are helping to give the project a positive rate of return. However, metal prices are very variable and so this presents a significant risk to the financial viability of the project. Care will have to be given as to what metal prices will be used in the definitive feasibility study.
 
The exchange rate of the Euro against the US dollar has been quite variable with the Euro continuing to strengthen. As the project capital and operating cost estimates are presented in US dollars but many of the costs would be incurred in Euros, this presents a significant risk to the project.
 
31.5        MINERAL RESOURCES
 
The mineral resources of the Konttijärvi and Ahmavaara deposits were prepared by a very competent geologist with more than 20 years of deposit modeling experience. Many validation steps were undertaken to ensure that mineralized constraints, assay values, grade capping, variography and grade interpolation were within acceptable limits. Numerous checks were incorporated to ensure that total metal content and domain volumes were validated and therefore not generating erroneous results. The risk associated with the mineral resource estimates is considered to be low.
 
31.6        MINING
 
The mining of the Konttijärvi and Ahmavaara deposits is subject to potential scheduling difficulties if production targets cannot be adequately fulfilled. In the event that skilled labour is not available in the required numbers, the production rate would be reduced. Geotechnical difficulties from any localized pit wall failures could seriously jeopardize short-term mining plans and may result in the loss of some mineable material during remediation. The selection of mining equipment is more than adequate for the task and, with proper maintenance, should not be a major concern for this project. The supply of mine consumables should not be a problem due to the Suhanko project proximity to established mine suppliers in the area. The predictability of mining dilution is somewhat uncertain at this time and only theoretical estimates are available. The overall risk associated with mining is considered to be moderate.
 
31.7        METALLURGY
 
The metallurgy of the Konttijärvi and Ahmavaara ores is very complex. Also, the success of the project is very dependent on the revenues from the sale of copper, nickel, platinum, palladium and, to a lesser extent, gold. Consequently, the highest recoveries possible are required for all these metals. This complicates the selection of the optimum process flowsheet. Extensive testwork that has been carried out on the Suhanko Project including bench-scale and pilot plant work and the metallurgical results have given highly variable results. Further work is required on fresh samples of the ore that are as representative as possible of the Konttijärvi and Ahmavaara deposits. Consequently, from an operational point of view, the most significant risk is associated with the metallurgy of the project.
 
183

 
31.8        CAPITAL COST/OPERATING COST
 
The capital cost estimate for the Scoping Study has an accuracy level of ±30% and includes a contingency of 20%. This is considered to be consistent with the level of the design work that has been carried out on the project. The estimate is based on the 2005 Lycopodium Study estimate that had an accuracy of ±15%. The costs were escalated to cover the increases from late 2004 to third quarter 2007. Some changes were made to the process flow-sheet based on the results of the recent metallurgical testwork and the plant capacity was reduced to 7.5 Mt/a. A visit was made to the Suhanko Project to review the infrastructure and some cost reductions were made. The risk associated with the capital cost is considered to be moderate.
 
The operating cost estimate for the Scoping Study has an accuracy of ±30% and is considered to be consistent with the level of design work that has been carried out on the project. Recent labour rates and electrical power cost have been used and the maintenance supplies are based on the latest capital cost estimate. The reagent consumptions are based on the latest metallurgical testwork and operating experience from Lac des Îles. The reagent costs were based on the 2005 Lycopodium Feasibility Study and checked against current pricing. The risk associated with the operating cost is considered to be moderate.
 
31.9        SALEABLE CONCENTRATE
 
The risk associated with producing a saleable concentrate is quite high. Testwork todate has shown that it has been difficult to obtain a concentrate with sufficient grades of the economically valuable metals while also obtaining acceptable recoveries. Gold Fields approached several potential concentrate buyers as part of the 2005 Feasibility Study and was apparently unable to obtain attractive terms. The marketing of the concentrate is an area that needs to be given close attention. An alternative is to consider installing a hydrometallurgical plant to treat the concentrate and produce higher quality products for sale. However, the option introduces some additional risks in that the Platsol process that is the most likely hydrometallurgical process to be chosen has not been commercialized.
 
184

 
32.0        RECOMMENDATIONS FOR FUTURE WORK
 
The Suhanko Project economics at the 18-month trailing average metal prices used in this study are considered to be sufficiently attractive to justify further investigation. Aker Kvaerner recommends that the following work be done to improve the project return.
 
32.1        OREBODY ISSUES
 
(a)                                   An estimated diluted potentially mineable portion of the mineral resources has been developed, which is a significant step forward from the 2005 Lycopodium Feasibility Study. Additional infill drilling is required to convert as much as possible of the Inferred resources to Measured and Indicated resources and this work is currently underway.
 
(b)                                  Additional investigations are required into a better understanding of the mineralogical characteristics of the orebody.
 
(c)                                   The Scoping Study has been based on mining Konttijärvi first followed by Ahmavaara. However, due to the variability of the ore lithologies, a detailed mine plan is required to better understand the metallurgical characteristics of the ore that will be fed to the concentrator for each operating period of the schedule. This work is required before any additional metallurgical testwork is carried out so that the samples used reflect the mining schedule.
 
32.2        MINERALOGY/PLANT DESIGN
 
(a)                                   Extensive testwork that has been carried out by Gold Fields and the recent work by Aker Kvaerner on behalf of GFAP, points out that this appears to be a difficult ore to process. Piloting of the flow sheet is recommended. Particular care needs to be taken in the selection of the samples and close supervision of the sample collection and the operation of the pilot plant as recommended.
 
(b)                                  The metal recoveries used in this report seem low. This may have been due to the use of oxidized ore samples in the testwork. It is recommended that more metallurgical testwork be done on both ore deposits using fresh samples. This work should include the following:
 
•                   bench scale testing on both ores emulating the Aker Kvaerner flow sheet
 
•                   locked cycle work once a productive reagent scheme has been established
 
•                   work to confirm the need for concentrate regrinding on the Konttijärvi ore.
 
(c)                                   The reagent suite needs to be optimized so as to reduce operating costs.
 
(d)                                  Additional mineralogical work should be done on the flotation residue to examine the location of metal values within the tailings. This should lead to improved metal recoveries.
 
(e)                                   It may be possible to improve the recoveries at the expense of concentrate grade. This would depend upon finding a smelter that would accept the lower grade concentrate, or the use of a concentrate leach technology such as Platsol. This avenue should be explored further.
 
(f)                                     The use of a high-pressure rolls crusher in the primary comminution circuit seems to improve the project economics. These numbers need to be
 
185

 
confirmed using actual crushing test data as required by the equipment manufacturer.
 
(g)                                  The effects of increasing the annual throughput rate to 10 Mt/a should be evaluated.
 
32.3        MARKETING
 
(a)                                   Additional work on the marketability and smelting cost of the concentrate should be done. This should include developing a detailed marketing strategy
 
186

 
APPENDICES
 
187

 
APPENDIX 1A
 
KONTTIJÄRVI
2PGE + AU RESOURCE INVENTORY
 
188

 

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	2PGE+AU	 	2PGE+AU	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	4.0 g/t	 	0.968	 	4.259	 	1.199	 	0.308	 	0.095	 	0.248	 	5.766	 	0.179	 	
 	 	3.8 g/t	 	1.201	 	3.994	 	1.127	 	0.285	 	0.096	 	0.243	 	5.406	 	0.209	 	
 	 	3.6 g/t	 	1.427	 	3.792	 	1.069	 	0.275	 	0.096	 	0.241	 	5.136	 	0.236	 	
 	 	3.4 g/t	 	1.792	 	3.548	 	0.997	 	0.257	 	0.096	 	0.235	 	4.802	 	0.277	 	
 	 	3.2 g/t	 	2.184	 	3.349	 	0.942	 	0.242	 	0.095	 	0.227	 	4.533	 	0.318	 	
 	 	3.0 g/t	 	2.598	 	3.180	 	0.894	 	0.230	 	0.094	 	0.223	 	4.305	 	0.360	 	
 	 	2.8 g/t	 	3.046	 	3.028	 	0.853	 	0.218	 	0.092	 	0.215	 	4.099	 	0.401	 	
 	 	2.6 g/t	 	3.621	 	2.862	 	0.807	 	0.207	 	0.092	 	0.207	 	3.876	 	0.451	 	
 	 	2.4 g/t	 	4.386	 	2.685	 	0.757	 	0.194	 	0.089	 	0.200	 	3.636	 	0.513	 	
 	 	2.2 g/t	 	5.254	 	2.520	 	0.713	 	0.181	 	0.088	 	0.192	 	3.414	 	0.577	 	
 	 	2.0 g/t	 	6.313	 	2.355	 	0.667	 	0.170	 	0.085	 	0.185	 	3.193	 	0.648	 	
 	 	1.8 g/t	 	7.518	 	2.201	 	0.625	 	0.160	 	0.083	 	0.177	 	2.986	 	0.722	 	
 	 	1.6 g/t	 	8.827	 	2.058	 	0.586	 	0.149	 	0.081	 	0.169	 	2.794	 	0.793	 	
 	 	1.4 g/t	 	10.527	 	1.903	 	0.543	 	0.139	 	0.079	 	0.160	 	2.585	 	0.875	 	
 	 	1.2 g/t	 	12.268	 	1.768	 	0.505	 	0.129	 	0.076	 	0.151	 	2.403	 	0.948	 	
 	 	1.0 g/t	 	14.094	 	1.644	 	0.471	 	0.120	 	0.073	 	0.143	 	2.234	 	1.012	 	
 	 	0.8 g/t	 	16.293	 	1.510	 	0.433	 	0.111	 	0.070	 	0.133	 	2.054	 	1.076	 	
 	 	0.1 g/t	 	32.388	 	0.873	 	0.252	 	0.068	 	0.049	 	0.083	 	1.192	 	1.241	 	
IND	 	4.0 g/t	 	0.943	 	5.139	 	1.520	 	0.406	 	0.062	 	0.250	 	7.065	 	0.214	 	
 	 	3.8 g/t	 	1.032	 	4.948	 	1.459	 	0.386	 	0.065	 	0.250	 	6.793	 	0.225	 	
 	 	3.6 g/t	 	1.215	 	4.610	 	1.355	 	0.367	 	0.069	 	0.247	 	6.332	 	0.247	 	
 	 	3.4 g/t	 	1.453	 	4.269	 	1.249	 	0.343	 	0.069	 	0.245	 	5.862	 	0.274	 	
 	 	3.2 g/t	 	1.798	 	3.917	 	1.143	 	0.311	 	0.073	 	0.236	 	5.372	 	0.311	 	
 	 	3.0 g/t	 	2.070	 	3.704	 	1.076	 	0.293	 	0.073	 	0.229	 	5.073	 	0.338	 	
 	 	2.8 g/t	 	2.488	 	3.442	 	0.995	 	0.271	 	0.074	 	0.224	 	4.707	 	0.377	 	
 	 	2.6 g/t	 	2.988	 	3.197	 	0.922	 	0.249	 	0.076	 	0.218	 	4.369	 	0.420	 	
 	 	2.4 g/t	 	3.709	 	2.934	 	0.845	 	0.227	 	0.076	 	0.210	 	4.006	 	0.478	 	
 	 	2.2 g/t	 	4.479	 	2.722	 	0.782	 	0.208	 	0.076	 	0.203	 	3.712	 	0.535	 	
 	 	2.0 g/t	 	5.563	 	2.493	 	0.713	 	0.190	 	0.073	 	0.196	 	3.396	 	0.607	 	
 	 	1.8 g/t	 	7.013	 	2.268	 	0.645	 	0.172	 	0.071	 	0.184	 	3.085	 	0.696	 	
 	 	1.6 g/t	 	8.681	 	2.073	 	0.589	 	0.156	 	0.069	 	0.175	 	2.818	 	0.786	 	
 	 	1.4 g/t	 	10.982	 	1.872	 	0.530	 	0.140	 	0.065	 	0.165	 	2.542	 	0.898	 	
 	 	1.2 g/t	 	13.658	 	1.692	 	0.479	 	0.126	 	0.062	 	0.153	 	2.298	 	1.009	 	
 	 	1.0 g/t	 	16.611	 	1.536	 	0.434	 	0.115	 	0.058	 	0.144	 	2.086	 	1.114	 	
 	 	0.8 g/t	 	20.591	 	1.367	 	0.386	 	0.104	 	0.054	 	0.132	 	1.856	 	1.229	 	
 	 	0.1 g/t	 	61.382	 	0.582	 	0.166	 	0.048	 	0.030	 	0.064	 	0.796	 	1.570	 	
INF	 	4.0 g/t	 	0.398	 	6.131	 	1.935	 	0.319	 	0.053	 	0.202	 	8.384	 	0.107	 	
 	 	3.8 g/t	 	0.424	 	5.938	 	1.870	 	0.307	 	0.052	 	0.199	 	8.115	 	0.111	 	
 	 	3.6 g/t	 	0.449	 	5.759	 	1.807	 	0.300	 	0.052	 	0.202	 	7.867	 	0.114	 	
 	 	3.4 g/t	 	0.479	 	5.562	 	1.736	 	0.293	 	0.052	 	0.207	 	7.591	 	0.117	 	
 	 	3.2 g/t	 	0.534	 	5.246	 	1.626	 	0.276	 	0.051	 	0.205	 	7.148	 	0.123	 	
 	 	3.0 g/t	 	0.585	 	4.991	 	1.543	 	0.264	 	0.053	 	0.203	 	6.798	 	0.128	 	
 	 	2.8 g/t	 	0.686	 	4.576	 	1.401	 	0.246	 	0.056	 	0.205	 	6.223	 	0.137	 	
 	 	2.6 g/t	 	0.787	 	4.250	 	1.291	 	0.230	 	0.056	 	0.203	 	5.771	 	0.146	 	
 	 	2.4 g/t	 	0.953	 	3.835	 	1.155	 	0.211	 	0.053	 	0.199	 	5.202	 	0.159	 	
 	 	2.2 g/t	 	1.084	 	3.582	 	1.070	 	0.200	 	0.052	 	0.196	 	4.852	 	0.169	 	
 	 	2.0 g/t	 	1.246	 	3.320	 	0.985	 	0.188	 	0.051	 	0.196	 	4.494	 	0.180	 	
 	 	1.8 g/t	 	1.493	 	3.006	 	0.884	 	0.173	 	0.050	 	0.189	 	4.063	 	0.195	 	
 	 	1.6 g/t	 	1.813	 	2.700	 	0.788	 	0.159	 	0.048	 	0.182	 	3.647	 	0.213	 	
 	 	1.4 g/t	 	2.353	 	2.336	 	0.674	 	0.142	 	0.044	 	0.174	 	3.152	 	0.238	 	
 	 	1.2 g/t	 	2.997	 	2.042	 	0.584	 	0.126	 	0.042	 	0.162	 	2.753	 	0.265	 	
 	 	1.0 g/t	 	3.767	 	1.793	 	0.508	 	0.113	 	0.039	 	0.150	 	2.414	 	0.292	 	
 	 	0.8 g/t	 	5.228	 	1.479	 	0.414	 	0.097	 	0.036	 	0.137	 	1.990	 	0.334	 	
 	 	0.1 g/t	 	28.885	 	0.390	 	0.111	 	0.033	 	0.014	 	0.047	 	0.534	 	0.496	 	
 
189

 
APPENDIX 1B
 
KONTTIJÄRVI
PDEQ RESOURCE INVENTORY
 
190

 
APPENDIX 1B: KONTTIJÄRVI PDEQ RESOURCE INVENTORY

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	PDEQ	 	PDEQ	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	4.0 g/t	 	11.179	 	1.822	 	0.523	 	0.135	 	0.084	 	0.156	 	6.704	 	2.409	 	
 	 	3.8 g/t	 	11.819	 	1.776	 	0.510	 	0.131	 	0.083	 	0.153	 	6.552	 	2.490	 	
 	 	3.6 g/t	 	12.480	 	1.729	 	0.496	 	0.128	 	0.082	 	0.150	 	6.402	 	2.569	 	
 	 	3.4 g/t	 	13.257	 	1.674	 	0.481	 	0.124	 	0.080	 	0.146	 	6.232	 	2.656	 	
 	 	3.2 g/t	 	13.942	 	1.631	 	0.469	 	0.121	 	0.079	 	0.143	 	6.087	 	2.729	 	
 	 	3.0 g/t	 	14.748	 	1.582	 	0.455	 	0.117	 	0.077	 	0.139	 	5.924	 	2.809	 	
 	 	2.8 g/t	 	15.648	 	1.526	 	0.440	 	0.113	 	0.075	 	0.135	 	5.750	 	2.893	 	
 	 	2.6 g/t	 	16.422	 	1.483	 	0.427	 	0.110	 	0.074	 	0.132	 	5.606	 	2.960	 	
 	 	2.4 g/t	 	17.340	 	1.433	 	0.413	 	0.107	 	0.072	 	0.129	 	5.442	 	3.034	 	
 	 	2.2 g/t	 	18.231	 	1.388	 	0.400	 	0.103	 	0.071	 	0.125	 	5.288	 	3.099	 	
 	 	2.0 g/t	 	19.364	 	1.333	 	0.385	 	0.100	 	0.069	 	0.121	 	5.101	 	3.176	 	
 	 	1.8 g/t	 	20.473	 	1.282	 	0.370	 	0.096	 	0.067	 	0.117	 	4.928	 	3.244	 	
 	 	1.6 g/t	 	21.744	 	1.227	 	0.354	 	0.092	 	0.065	 	0.113	 	4.739	 	3.313	 	
 	 	1.4 g/t	 	23.206	 	1.167	 	0.337	 	0.088	 	0.063	 	0.108	 	4.535	 	3.383	 	
 	 	1.2 g/t	 	24.661	 	1.112	 	0.321	 	0.084	 	0.061	 	0.104	 	4.344	 	3.444	 	
 	 	1.0 g/t	 	26.031	 	1.064	 	0.307	 	0.081	 	0.059	 	0.100	 	4.173	 	3.493	 	
 	 	0.8 g/t	 	27.278	 	1.023	 	0.295	 	0.078	 	0.057	 	0.096	 	4.024	 	3.529	 	
 	 	0.1 g/t	 	32.339	 	0.874	 	0.252	 	0.068	 	0.049	 	0.083	 	3.456	 	3.593	 	
IND	 	4.0 g/t	 	11.099	 	1.834	 	0.522	 	0.139	 	0.070	 	0.167	 	6.497	 	2.319	 	
 	 	3.8 g/t	 	12.173	 	1.761	 	0.501	 	0.133	 	0.068	 	0.163	 	6.268	 	2.453	 	
 	 	3.6 g/t	 	13.132	 	1.701	 	0.484	 	0.129	 	0.067	 	0.159	 	6.082	 	2.568	 	
 	 	3.4 g/t	 	14.224	 	1.638	 	0.465	 	0.124	 	0.066	 	0.154	 	5.883	 	2.690	 	
 	 	3.2 g/t	 	15.272	 	1.582	 	0.449	 	0.120	 	0.064	 	0.150	 	5.706	 	2.802	 	
 	 	3.0 g/t	 	16.542	 	1.520	 	0.431	 	0.115	 	0.062	 	0.146	 	5.506	 	2.929	 	
 	 	2.8 g/t	 	18.015	 	1.452	 	0.412	 	0.110	 	0.061	 	0.141	 	5.293	 	3.066	 	
 	 	2.6 g/t	 	19.468	 	1.393	 	0.395	 	0.106	 	0.059	 	0.137	 	5.099	 	3.191	 	
 	 	2.4 g/t	 	20.778	 	1.344	 	0.381	 	0.103	 	0.057	 	0.133	 	4.935	 	3.297	 	
 	 	2.2 g/t	 	22.384	 	1.287	 	0.364	 	0.099	 	0.055	 	0.128	 	4.746	 	3.415	 	
 	 	2.0 g/t	 	24.283	 	1.225	 	0.346	 	0.094	 	0.054	 	0.123	 	4.538	 	3.543	 	
 	 	1.8 g/t	 	26.598	 	1.156	 	0.327	 	0.089	 	0.051	 	0.118	 	4.308	 	3.684	 	
 	 	1.6 g/t	 	28.883	 	1.095	 	0.309	 	0.085	 	0.049	 	0.112	 	4.102	 	3.809	 	
 	 	1.4 g/t	 	31.639	 	1.026	 	0.290	 	0.080	 	0.047	 	0.106	 	3.875	 	3.942	 	
 	 	1.2 g/t	 	34.603	 	0.961	 	0.272	 	0.075	 	0.045	 	0.101	 	3.654	 	4.065	 	
 	 	1.0 g/t	 	37.571	 	0.902	 	0.255	 	0.071	 	0.043	 	0.095	 	3.452	 	4.170	 	
 	 	0.8 g/t	 	40.966	 	0.842	 	0.238	 	0.067	 	0.041	 	0.090	 	3.241	 	4.268	 	
 	 	0.1 g/t	 	61.266	 	0.583	 	0.166	 	0.048	 	0.030	 	0.064	 	2.289	 	4.510	 	
INF	 	4.0 g/t	 	1.974	 	2.537	 	0.739	 	0.153	 	0.051	 	0.191	 	7.770	 	0.493	 	
 	 	3.8 g/t	 	2.281	 	2.345	 	0.680	 	0.144	 	0.049	 	0.184	 	7.249	 	0.532	 	
 	 	3.6 g/t	 	2.513	 	2.227	 	0.641	 	0.137	 	0.048	 	0.180	 	6.922	 	0.559	 	
 	 	3.4 g/t	 	2.760	 	2.106	 	0.604	 	0.131	 	0.047	 	0.177	 	6.616	 	0.587	 	
 	 	3.2 g/t	 	3.062	 	1.986	 	0.567	 	0.125	 	0.045	 	0.172	 	6.289	 	0.619	 	
 	 	3.0 g/t	 	3.511	 	1.838	 	0.522	 	0.118	 	0.043	 	0.166	 	5.881	 	0.664	 	
 	 	2.8 g/t	 	3.877	 	1.737	 	0.492	 	0.113	 	0.042	 	0.161	 	5.599	 	0.698	 	
 	 	2.6 g/t	 	4.356	 	1.622	 	0.459	 	0.107	 	0.040	 	0.155	 	5.279	 	0.739	 	
 	 	2.4 g/t	 	4.946	 	1.509	 	0.425	 	0.100	 	0.038	 	0.147	 	4.946	 	0.787	 	
 	 	2.2 g/t	 	5.606	 	1.401	 	0.393	 	0.094	 	0.037	 	0.140	 	4.634	 	0.835	 	
 	 	2.0 g/t	 	6.231	 	1.314	 	0.368	 	0.089	 	0.036	 	0.133	 	4.381	 	0.878	 	
 	 	1.8 g/t	 	7.028	 	1.221	 	0.341	 	0.084	 	0.034	 	0.127	 	4.100	 	0.927	 	
 	 	1.6 g/t	 	8.025	 	1.122	 	0.314	 	0.078	 	0.032	 	0.119	 	3.803	 	0.981	 	
 	 	1.4 g/t	 	8.995	 	1.041	 	0.290	 	0.073	 	0.030	 	0.112	 	3.554	 	1.028	 	
 	 	1.2 g/t	 	10.089	 	0.963	 	0.268	 	0.068	 	0.029	 	0.105	 	3.309	 	1.073	 	
 	 	1.0 g/t	 	11.638	 	0.867	 	0.241	 	0.062	 	0.027	 	0.097	 	3.014	 	1.128	 	
 	 	0.8 g/t	 	13.353	 	0.783	 	0.217	 	0.057	 	0.025	 	0.090	 	2.743	 	1.178	 	
 	 	0.1 g/t	 	28.885	 	0.390	 	0.111	 	0.033	 	0.014	 	0.047	 	1.444	 	1.341	 	
 
191

 
APPENDIX 1C
 
AHMAVAARA
2PGE + AU RESOURCE INVENTORY
 
192

 
APPENDIX 1C: AHMAVAARA 2PGE + AU RESOURCE INVENTORY
AHMAVAARA 2PGE + AU

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	2PGE+AU	 	2PGE+AU	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	5.0 g/t	 	0.040	 	5.656	 	0.725	 	0.762	 	0.335	 	0.754	 	7.144	 	0.009	 	
 	 	4.0 g/t	 	0.167	 	4.268	 	0.555	 	0.415	 	0.368	 	0.832	 	5.239	 	0.028	 	
 	 	3.0 g/t	 	0.880	 	2.871	 	0.513	 	0.359	 	0.225	 	0.474	 	3.743	 	0.106	 	
 	 	2.8 g/t	 	1.154	 	2.691	 	0.503	 	0.349	 	0.208	 	0.431	 	3.543	 	0.131	 	
 	 	2.6 g/t	 	1.344	 	2.601	 	0.489	 	0.331	 	0.196	 	0.420	 	3.421	 	0.148	 	
 	 	2.4 g/t	 	1.680	 	2.477	 	0.461	 	0.300	 	0.198	 	0.423	 	3.238	 	0.175	 	
 	 	2.2 g/t	 	2.419	 	2.242	 	0.439	 	0.268	 	0.174	 	0.387	 	2.949	 	0.229	 	
 	 	2.0 g/t	 	3.338	 	2.059	 	0.412	 	0.244	 	0.162	 	0.375	 	2.715	 	0.291	 	
 	 	1.8 g/t	 	4.283	 	1.919	 	0.385	 	0.229	 	0.152	 	0.355	 	2.533	 	0.349	 	
 	 	1.6 g/t	 	5.626	 	1.769	 	0.357	 	0.210	 	0.141	 	0.333	 	2.336	 	0.422	 	
 	 	1.4 g/t	 	7.661	 	1.598	 	0.322	 	0.189	 	0.129	 	0.309	 	2.110	 	0.520	 	
 	 	1.2 g/t	 	9.752	 	1.466	 	0.296	 	0.174	 	0.119	 	0.291	 	1.936	 	0.607	 	
 	 	1.0 g/t	 	12.825	 	1.313	 	0.265	 	0.156	 	0.109	 	0.267	 	1.734	 	0.715	 	
 	 	0.8 g/t	 	16.333	 	1.176	 	0.238	 	0.141	 	0.098	 	0.245	 	1.554	 	0.816	 	
 	 	0.1 g/t	 	44.880	 	0.624	 	0.125	 	0.076	 	0.054	 	0.142	 	0.824	 	1.189	 	
IND	 	5.0 g/t	 	0.288	 	4.479	 	0.626	 	0.814	 	0.428	 	0.968	 	5.920	 	0.055	 	
 	 	4.0 g/t	 	0.598	 	3.869	 	0.593	 	0.688	 	0.386	 	0.806	 	5.150	 	0.099	 	
 	 	3.0 g/t	 	2.731	 	2.824	 	0.565	 	0.365	 	0.214	 	0.430	 	3.755	 	0.330	 	
 	 	2.8 g/t	 	3.838	 	2.634	 	0.541	 	0.329	 	0.198	 	0.405	 	3.504	 	0.432	 	
 	 	2.6 g/t	 	5.400	 	2.455	 	0.513	 	0.300	 	0.185	 	0.380	 	3.268	 	0.567	 	
 	 	2.4 g/t	 	7.366	 	2.296	 	0.488	 	0.276	 	0.168	 	0.348	 	3.060	 	0.725	 	
 	 	2.2 g/t	 	10.213	 	2.135	 	0.462	 	0.251	 	0.152	 	0.317	 	2.848	 	0.935	 	
 	 	2.0 g/t	 	14.115	 	1.981	 	0.430	 	0.230	 	0.142	 	0.303	 	2.640	 	1.198	 	
 	 	1.8 g/t	 	19.030	 	1.834	 	0.402	 	0.211	 	0.131	 	0.284	 	2.447	 	1.497	 	
 	 	1.6 g/t	 	25.978	 	1.687	 	0.366	 	0.194	 	0.124	 	0.275	 	2.247	 	1.876	 	
 	 	1.4 g/t	 	34.356	 	1.550	 	0.333	 	0.180	 	0.117	 	0.267	 	2.063	 	2.279	 	
 	 	1.2 g/t	 	43.936	 	1.424	 	0.304	 	0.168	 	0.110	 	0.257	 	1.895	 	2.677	 	
 	 	1.0 g/t	 	54.570	 	1.307	 	0.276	 	0.156	 	0.103	 	0.246	 	1.740	 	3.052	 	
 	 	0.8 g/t	 	67.264	 	1.189	 	0.249	 	0.143	 	0.095	 	0.233	 	1.581	 	3.419	 	
 	 	0.1 g/t	 	123.561	 	0.801	 	0.165	 	0.098	 	0.066	 	0.171	 	1.064	 	4.226	 	
INF	 	5.0 g/t	 	0.041	 	5.295	 	0.949	 	0.423	 	0.672	 	0.831	 	6.667	 	0.009	 	
 	 	4.0 g/t	 	0.147	 	3.931	 	0.885	 	0.278	 	0.285	 	0.373	 	5.094	 	0.024	 	
 	 	3.0 g/t	 	0.610	 	2.874	 	0.677	 	0.210	 	0.170	 	0.264	 	3.761	 	0.074	 	
 	 	2.8 g/t	 	0.757	 	2.729	 	0.662	 	0.203	 	0.158	 	0.250	 	3.594	 	0.088	 	
 	 	2.6 g/t	 	1.073	 	2.516	 	0.618	 	0.190	 	0.142	 	0.232	 	3.325	 	0.115	 	
 	 	2.4 g/t	 	1.451	 	2.343	 	0.579	 	0.187	 	0.130	 	0.223	 	3.109	 	0.145	 	
 	 	2.2 g/t	 	2.267	 	2.109	 	0.516	 	0.192	 	0.126	 	0.224	 	2.817	 	0.205	 	
 	 	2.0 g/t	 	3.270	 	1.941	 	0.472	 	0.181	 	0.117	 	0.219	 	2.594	 	0.273	 	
 	 	1.8 g/t	 	4.590	 	1.790	 	0.433	 	0.170	 	0.112	 	0.215	 	2.393	 	0.353	 	
 	 	1.6 g/t	 	6.780	 	1.621	 	0.384	 	0.162	 	0.110	 	0.218	 	2.167	 	0.472	 	
 	 	1.4 g/t	 	9.558	 	1.471	 	0.344	 	0.157	 	0.102	 	0.215	 	1.973	 	0.606	 	
 	 	1.2 g/t	 	13.691	 	1.321	 	0.300	 	0.149	 	0.097	 	0.212	 	1.770	 	0.779	 	
 	 	1.0 g/t	 	18.244	 	1.198	 	0.268	 	0.138	 	0.091	 	0.205	 	1.604	 	0.941	 	
 	 	0.8 g/t	 	23.096	 	1.087	 	0.239	 	0.128	 	0.085	 	0.197	 	1.455	 	1.080	 	
 	 	0.1 g/t	 	35.448	 	0.853	 	0.184	 	0.102	 	0.069	 	0.167	 	1.139	 	1.298	 	
 
193

 
AHMAVAARA EAST 2PGE + AU

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	2PGE+AU	 	2PGE+AU	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	5.0 g/t	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	
 	 	4.0 g/t	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	
 	 	3.0 g/t	 	0.029	 	2.458	 	0.553	 	0.171	 	0.111	 	0.208	 	3.182	 	0.003	 	
 	 	2.8 g/t	 	0.070	 	2.344	 	0.500	 	0.172	 	0.171	 	0.319	 	3.016	 	0.007	 	
 	 	2.6 g/t	 	0.070	 	2.344	 	0.500	 	0.172	 	0.171	 	0.319	 	3.016	 	0.007	 	
 	 	2.4 g/t	 	0.091	 	2.236	 	0.497	 	0.173	 	0.155	 	0.291	 	2.905	 	0.009	 	
 	 	2.2 g/t	 	0.323	 	1.868	 	0.448	 	0.156	 	0.124	 	0.237	 	2.472	 	0.026	 	
 	 	2.0 g/t	 	0.512	 	1.765	 	0.421	 	0.155	 	0.113	 	0.224	 	2.342	 	0.039	 	
 	 	1.8 g/t	 	0.865	 	1.631	 	0.378	 	0.155	 	0.106	 	0.225	 	2.164	 	0.060	 	
 	 	1.6 g/t	 	1.261	 	1.518	 	0.343	 	0.161	 	0.099	 	0.234	 	2.022	 	0.082	 	
 	 	1.4 g/t	 	1.916	 	1.372	 	0.306	 	0.159	 	0.094	 	0.231	 	1.837	 	0.113	 	
 	 	1.2 g/t	 	2.689	 	1.254	 	0.277	 	0.151	 	0.089	 	0.224	 	1.682	 	0.145	 	
 	 	1.0 g/t	 	3.783	 	1.126	 	0.244	 	0.141	 	0.082	 	0.217	 	1.510	 	0.184	 	
 	 	0.8 g/t	 	5.424	 	0.989	 	0.210	 	0.124	 	0.075	 	0.202	 	1.323	 	0.231	 	
 	 	0.1 g/t	 	20.742	 	0.488	 	0.097	 	0.061	 	0.040	 	0.119	 	0.646	 	0.431	 	
IND	 	5.0 g/t	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	
 	 	4.0 g/t	 	0.021	 	3.159	 	0.775	 	0.206	 	0.144	 	0.235	 	4.140	 	0.003	 	
 	 	3.0 g/t	 	0.458	 	2.481	 	0.625	 	0.191	 	0.107	 	0.211	 	3.298	 	0.049	 	
 	 	2.8 g/t	 	0.633	 	2.383	 	0.605	 	0.192	 	0.114	 	0.224	 	3.180	 	0.065	 	
 	 	2.6 g/t	 	0.851	 	2.285	 	0.585	 	0.189	 	0.115	 	0.222	 	3.058	 	0.084	 	
 	 	2.4 g/t	 	1.610	 	2.073	 	0.528	 	0.189	 	0.106	 	0.204	 	2.790	 	0.144	 	
 	 	2.2 g/t	 	2.986	 	1.897	 	0.483	 	0.184	 	0.103	 	0.196	 	2.563	 	0.246	 	
 	 	2.0 g/t	 	4.717	 	1.773	 	0.446	 	0.172	 	0.100	 	0.199	 	2.391	 	0.363	 	
 	 	1.8 g/t	 	7.063	 	1.651	 	0.409	 	0.166	 	0.097	 	0.196	 	2.225	 	0.505	 	
 	 	1.6 g/t	 	10.496	 	1.526	 	0.368	 	0.156	 	0.095	 	0.203	 	2.051	 	0.692	 	
 	 	1.4 g/t	 	15.503	 	1.396	 	0.324	 	0.151	 	0.094	 	0.215	 	1.871	 	0.933	 	
 	 	1.2 g/t	 	21.318	 	1.280	 	0.289	 	0.144	 	0.091	 	0.218	 	1.713	 	1.174	 	
 	 	1.0 g/t	 	27.870	 	1.173	 	0.258	 	0.137	 	0.087	 	0.217	 	1.568	 	1.405	 	
 	 	0.8 g/t	 	36.677	 	1.055	 	0.225	 	0.126	 	0.082	 	0.212	 	1.406	 	1.658	 	
 	 	0.1 g/t	 	80.174	 	0.668	 	0.137	 	0.081	 	0.055	 	0.153	 	0.886	 	2.283	 	
INF	 	5.0 g/t	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	
 	 	4.0 g/t	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	0.000	 	
 	 	3.0 g/t	 	0.127	 	2.400	 	0.594	 	0.147	 	0.105	 	0.159	 	3.141	 	0.013	 	
 	 	2.8 g/t	 	0.148	 	2.371	 	0.601	 	0.141	 	0.107	 	0.158	 	3.113	 	0.015	 	
 	 	2.6 g/t	 	0.295	 	2.162	 	0.565	 	0.158	 	0.111	 	0.168	 	2.884	 	0.027	 	
 	 	2.4 g/t	 	0.526	 	2.035	 	0.518	 	0.159	 	0.100	 	0.175	 	2.712	 	0.046	 	
 	 	2.2 g/t	 	1.136	 	1.836	 	0.469	 	0.185	 	0.102	 	0.183	 	2.490	 	0.091	 	
 	 	2.0 g/t	 	1.766	 	1.738	 	0.439	 	0.170	 	0.100	 	0.188	 	2.347	 	0.133	 	
 	 	1.8 g/t	 	2.607	 	1.632	 	0.405	 	0.162	 	0.100	 	0.194	 	2.199	 	0.184	 	
 	 	1.6 g/t	 	3.870	 	1.516	 	0.365	 	0.155	 	0.101	 	0.206	 	2.036	 	0.253	 	
 	 	1.4 g/t	 	5.828	 	1.379	 	0.324	 	0.153	 	0.097	 	0.209	 	1.856	 	0.348	 	
 	 	1.2 g/t	 	8.955	 	1.237	 	0.280	 	0.146	 	0.092	 	0.209	 	1.663	 	0.479	 	
 	 	1.0 g/t	 	12.448	 	1.122	 	0.249	 	0.134	 	0.087	 	0.204	 	1.505	 	0.602	 	
 	 	0.8 g/t	 	16.311	 	1.016	 	0.221	 	0.124	 	0.081	 	0.196	 	1.361	 	0.714	 	
 	 	0.1 g/t	 	26.130	 	0.792	 	0.169	 	0.098	 	0.066	 	0.167	 	1.059	 	0.889	 	
 
194

 
AHMAVAARA WEST 2PGE + AU

MEAS	 	5.0 g/t	 	0.040	 	5.656	 	0.725	 	0.762	 	0.335	 	0.754	 	7.144	 	0.009	 	
 	 	4.0 g/t	 	0.167	 	4.268	 	0.555	 	0.415	 	0.368	 	0.832	 	5.239	 	0.028	 	
 	 	3.0 g/t	 	0.851	 	2.885	 	0.512	 	0.365	 	0.229	 	0.483	 	3.762	 	0.103	 	
 	 	2.8 g/t	 	1.083	 	2.713	 	0.503	 	0.361	 	0.210	 	0.438	 	3.577	 	0.125	 	
 	 	2.6 g/t	 	1.274	 	2.616	 	0.489	 	0.339	 	0.198	 	0.426	 	3.444	 	0.141	 	
 	 	2.4 g/t	 	1.588	 	2.491	 	0.459	 	0.308	 	0.201	 	0.430	 	3.257	 	0.166	 	
 	 	2.2 g/t	 	2.096	 	2.299	 	0.438	 	0.285	 	0.182	 	0.410	 	3.022	 	0.204	 	
 	 	2.0 g/t	 	2.826	 	2.112	 	0.410	 	0.261	 	0.171	 	0.403	 	2.782	 	0.253	 	
 	 	1.8 g/t	 	3.418	 	1.991	 	0.386	 	0.248	 	0.164	 	0.388	 	2.626	 	0.289	 	
 	 	1.6 g/t	 	4.364	 	1.842	 	0.361	 	0.224	 	0.153	 	0.361	 	2.426	 	0.340	 	
 	 	1.4 g/t	 	5.745	 	1.674	 	0.328	 	0.199	 	0.141	 	0.336	 	2.201	 	0.407	 	
 	 	1.2 g/t	 	7.063	 	1.546	 	0.304	 	0.182	 	0.131	 	0.316	 	2.032	 	0.462	 	
 	 	1.0 g/t	 	9.042	 	1.391	 	0.274	 	0.163	 	0.120	 	0.289	 	1.828	 	0.531	 	
 	 	0.8 g/t	 	10.909	 	1.268	 	0.252	 	0.149	 	0.109	 	0.266	 	1.669	 	0.585	 	
 	 	0.1 g/t	 	24.138	 	0.740	 	0.149	 	0.088	 	0.065	 	0.162	 	0.976	 	0.758	 	
IND	 	5.0 g/t	 	0.288	 	4.479	 	0.626	 	0.814	 	0.428	 	0.968	 	5.920	 	0.055	 	
 	 	4.0 g/t	 	0.577	 	3.895	 	0.586	 	0.706	 	0.394	 	0.827	 	5.187	 	0.096	 	
 	 	3.0 g/t	 	2.273	 	2.893	 	0.553	 	0.400	 	0.235	 	0.474	 	3.847	 	0.281	 	
 	 	2.8 g/t	 	3.205	 	2.684	 	0.528	 	0.356	 	0.215	 	0.441	 	3.568	 	0.368	 	
 	 	2.6 g/t	 	4.549	 	2.487	 	0.499	 	0.320	 	0.198	 	0.409	 	3.307	 	0.484	 	
 	 	2.4 g/t	 	5.755	 	2.358	 	0.477	 	0.301	 	0.185	 	0.388	 	3.136	 	0.580	 	
 	 	2.2 g/t	 	7.227	 	2.234	 	0.453	 	0.279	 	0.173	 	0.367	 	2.966	 	0.689	 	
 	 	2.0 g/t	 	9.398	 	2.085	 	0.422	 	0.259	 	0.163	 	0.355	 	2.765	 	0.836	 	
 	 	1.8 g/t	 	11.967	 	1.943	 	0.398	 	0.237	 	0.152	 	0.336	 	2.578	 	0.992	 	
 	 	1.6 g/t	 	15.482	 	1.796	 	0.365	 	0.219	 	0.144	 	0.325	 	2.380	 	1.184	 	
 	 	1.4 g/t	 	18.854	 	1.677	 	0.340	 	0.204	 	0.136	 	0.310	 	2.221	 	1.346	 	
 	 	1.2 g/t	 	22.619	 	1.559	 	0.317	 	0.190	 	0.128	 	0.294	 	2.066	 	1.503	 	
 	 	1.0 g/t	 	26.700	 	1.447	 	0.296	 	0.177	 	0.119	 	0.276	 	1.919	 	1.647	 	
 	 	0.8 g/t	 	30.587	 	1.349	 	0.277	 	0.165	 	0.112	 	0.260	 	1.791	 	1.761	 	
 	 	0.1 g/t	 	43.387	 	1.048	 	0.216	 	0.128	 	0.088	 	0.205	 	1.393	 	1.943	 	
INF	 	5.0 g/t	 	0.041	 	5.295	 	0.949	 	0.423	 	0.672	 	0.831	 	6.667	 	0.009	 	
 	 	4.0 g/t	 	0.147	 	3.931	 	0.885	 	0.278	 	0.285	 	0.373	 	5.094	 	0.024	 	
 	 	3.0 g/t	 	0.483	 	2.999	 	0.699	 	0.226	 	0.188	 	0.291	 	3.924	 	0.061	 	
 	 	2.8 g/t	 	0.609	 	2.816	 	0.677	 	0.218	 	0.170	 	0.272	 	3.711	 	0.073	 	
 	 	2.6 g/t	 	0.778	 	2.651	 	0.638	 	0.203	 	0.154	 	0.257	 	3.492	 	0.087	 	
 	 	2.4 g/t	 	0.925	 	2.519	 	0.614	 	0.203	 	0.146	 	0.250	 	3.335	 	0.099	 	
 	 	2.2 g/t	 	1.130	 	2.383	 	0.563	 	0.200	 	0.150	 	0.264	 	3.146	 	0.114	 	
 	 	2.0 g/t	 	1.504	 	2.179	 	0.511	 	0.194	 	0.137	 	0.257	 	2.884	 	0.140	 	
 	 	1.8 g/t	 	1.983	 	1.997	 	0.469	 	0.180	 	0.127	 	0.242	 	2.647	 	0.169	 	
 	 	1.6 g/t	 	2.910	 	1.761	 	0.410	 	0.171	 	0.121	 	0.235	 	2.341	 	0.219	 	
 	 	1.4 g/t	 	3.730	 	1.616	 	0.375	 	0.164	 	0.111	 	0.225	 	2.154	 	0.258	 	
 	 	1.2 g/t	 	4.736	 	1.479	 	0.338	 	0.155	 	0.105	 	0.217	 	1.973	 	0.300	 	
 	 	1.0 g/t	 	5.796	 	1.360	 	0.308	 	0.147	 	0.099	 	0.208	 	1.815	 	0.338	 	
 	 	0.8 g/t	 	6.785	 	1.260	 	0.283	 	0.138	 	0.093	 	0.200	 	1.681	 	0.367	 	
 	 	0.1 g/t	 	9.319	 	1.023	 	0.229	 	0.112	 	0.078	 	0.166	 	1.363	 	0.408	 	
 
195

 
APPENDIX 1D
 
AHMAVAARA
PDEQ RESOURCE INVENTORY
 
196

 
APPENDIX 1D: AHMAVAARA PDEQ RESOURCE INVENTORY

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	PDEQ	 	PDEQ	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	5.0 g/t	 	7.412	 	1.576	 	0.302	 	0.191	 	0.141	 	0.343	 	7.614	 	1.814	 	
 	 	4.0 g/t	 	11.191	 	1.364	 	0.267	 	0.166	 	0.119	 	0.295	 	6.554	 	2.358	 	
 	 	3.0 g/t	 	16.123	 	1.174	 	0.235	 	0.141	 	0.101	 	0.252	 	5.615	 	2.910	 	
 	 	2.8 g/t	 	17.566	 	1.127	 	0.226	 	0.136	 	0.097	 	0.243	 	5.392	 	3.045	 	
 	 	2.6 g/t	 	18.731	 	1.093	 	0.219	 	0.132	 	0.094	 	0.236	 	5.225	 	3.147	 	
 	 	2.4 g/t	 	20.365	 	1.046	 	0.209	 	0.126	 	0.089	 	0.227	 	5.008	 	3.279	 	
 	 	2.2 g/t	 	21.937	 	1.004	 	0.201	 	0.121	 	0.086	 	0.220	 	4.814	 	3.395	 	
 	 	2.0 g/t	 	24.020	 	0.953	 	0.190	 	0.115	 	0.082	 	0.210	 	4.577	 	3.535	 	
 	 	1.8 g/t	 	26.428	 	0.900	 	0.179	 	0.109	 	0.077	 	0.200	 	4.332	 	3.681	 	
 	 	1.6 g/t	 	29.087	 	0.849	 	0.169	 	0.103	 	0.073	 	0.190	 	4.091	 	3.826	 	
 	 	1.4 g/t	 	32.125	 	0.797	 	0.158	 	0.096	 	0.068	 	0.179	 	3.845	 	3.972	 	
 	 	1.2 g/t	 	35.058	 	0.751	 	0.149	 	0.091	 	0.064	 	0.170	 	3.633	 	4.095	 	
 	 	1.0 g/t	 	38.197	 	0.707	 	0.141	 	0.086	 	0.061	 	0.161	 	3.425	 	4.206	 	
 	 	0.8 g/t	 	41.344	 	0.666	 	0.133	 	0.081	 	0.057	 	0.152	 	3.233	 	4.298	 	
 	 	0.1 g/t	 	45.471	 	0.616	 	0.123	 	0.075	 	0.053	 	0.140	 	2.995	 	4.378	 	
IND	 	5.0 g/t	 	32.176	 	1.533	 	0.311	 	0.186	 	0.128	 	0.307	 	7.152	 	7.399	 	
 	 	4.0 g/t	 	49.886	 	1.335	 	0.275	 	0.162	 	0.110	 	0.267	 	6.209	 	9.959	 	
 	 	3.0 g/t	 	68.334	 	1.172	 	0.243	 	0.143	 	0.097	 	0.237	 	5.476	 	12.031	 	
 	 	2.8 g/t	 	71.862	 	1.144	 	0.237	 	0.139	 	0.094	 	0.233	 	5.349	 	12.359	 	
 	 	2.6 g/t	 	75.325	 	1.116	 	0.231	 	0.136	 	0.092	 	0.228	 	5.227	 	12.659	 	
 	 	2.4 g/t	 	79.304	 	1.086	 	0.225	 	0.132	 	0.090	 	0.223	 	5.091	 	12.979	 	
 	 	2.2 g/t	 	82.843	 	1.059	 	0.219	 	0.129	 	0.087	 	0.219	 	4.971	 	13.240	 	
 	 	2.0 g/t	 	87.208	 	1.028	 	0.212	 	0.125	 	0.085	 	0.213	 	4.827	 	13.535	 	
 	 	1.8 g/t	 	92.185	 	0.992	 	0.205	 	0.121	 	0.082	 	0.207	 	4.669	 	13.839	 	
 	 	1.6 g/t	 	98.281	 	0.951	 	0.196	 	0.116	 	0.079	 	0.200	 	4.485	 	14.172	 	
 	 	1.4 g/t	 	104.055	 	0.915	 	0.188	 	0.111	 	0.076	 	0.193	 	4.320	 	14.451	 	
 	 	1.2 g/t	 	109.667	 	0.881	 	0.181	 	0.107	 	0.073	 	0.187	 	4.165	 	14.684	 	
 	 	1.0 g/t	 	114.816	 	0.851	 	0.175	 	0.104	 	0.071	 	0.181	 	4.027	 	14.867	 	
 	 	0.8 g/t	 	119.042	 	0.827	 	0.170	 	0.101	 	0.069	 	0.176	 	3.917	 	14.990	 	
 	 	0.1 g/t	 	124.514	 	0.796	 	0.164	 	0.097	 	0.066	 	0.170	 	3.769	 	15.090	 	
INF	 	5.0 g/t	 	8.581	 	1.467	 	0.320	 	0.161	 	0.118	 	0.260	 	6.615	 	1.825	 	
 	 	4.0 g/t	 	15.276	 	1.244	 	0.270	 	0.146	 	0.101	 	0.231	 	5.689	 	2.794	 	
 	 	3.0 g/t	 	23.587	 	1.067	 	0.231	 	0.127	 	0.087	 	0.204	 	4.922	 	3.732	 	
 	 	2.8 g/t	 	24.973	 	1.042	 	0.226	 	0.124	 	0.085	 	0.200	 	4.809	 	3.862	 	
 	 	2.6 g/t	 	26.003	 	1.023	 	0.222	 	0.122	 	0.083	 	0.197	 	4.726	 	3.951	 	
 	 	2.4 g/t	 	27.448	 	0.997	 	0.216	 	0.119	 	0.081	 	0.193	 	4.609	 	4.068	 	
 	 	2.2 g/t	 	28.988	 	0.970	 	0.210	 	0.115	 	0.079	 	0.188	 	4.487	 	4.182	 	
 	 	2.0 g/t	 	29.832	 	0.955	 	0.207	 	0.114	 	0.077	 	0.186	 	4.420	 	4.239	 	
 	 	1.8 g/t	 	30.996	 	0.935	 	0.202	 	0.111	 	0.076	 	0.182	 	4.325	 	4.310	 	
 	 	1.6 g/t	 	32.203	 	0.914	 	0.198	 	0.109	 	0.074	 	0.178	 	4.227	 	4.376	 	
 	 	1.4 g/t	 	33.302	 	0.894	 	0.193	 	0.106	 	0.072	 	0.175	 	4.137	 	4.429	 	
 	 	1.2 g/t	 	34.166	 	0.878	 	0.190	 	0.104	 	0.071	 	0.172	 	4.065	 	4.465	 	
 	 	1.0 g/t	 	34.757	 	0.867	 	0.187	 	0.103	 	0.070	 	0.170	 	4.015	 	4.486	 	
 	 	0.8 g/t	 	35.198	 	0.858	 	0.186	 	0.102	 	0.070	 	0.168	 	3.976	 	4.499	 	
 	 	0.1 g/t	 	35.469	 	0.852	 	0.184	 	0.101	 	0.069	 	0.167	 	3.950	 	4.505	 	
 
197

 
AHMAVAARA EAST PDEQ

ROCKGROUP	 	GRADEGROUP	 	Tonnage	 	PD	 	PT	 	AU	 	NI	 	CU	 	PDEQ	 	PDEQ	 	
 	 	 	 	M mt	 	g/t	 	g/t	 	g/t	 	%	 	%	 	g/t	 	M oz	 	
MEAS	 	5.0 g/t	 	1.726	 	1.356	 	0.281	 	0.155	 	0.107	 	0.271	 	6.210	 	0.345	 	
 	 	4.0 g/t	 	3.137	 	1.167	 	0.242	 	0.147	 	0.092	 	0.246	 	5.433	 	0.548	 	
 	 	3.0 g/t	 	5.399	 	0.980	 	0.204	 	0.124	 	0.079	 	0.212	 	4.626	 	0.803	 	
 	 	2.8 g/t	 	6.149	 	0.932	 	0.193	 	0.118	 	0.075	 	0.204	 	4.417	 	0.873	 	
 	 	2.6 g/t	 	6.490	 	0.913	 	0.189	 	0.115	 	0.073	 	0.201	 	4.326	 	0.903	 	
 	 	2.4 g/t	 	7.372	 	0.864	 	0.178	 	0.109	 	0.070	 	0.193	 	4.111	 	0.974	 	
 	 	2.2 g/t	 	8.273	 	0.821	 	0.168	 	0.103	 	0.066	 	0.185	 	3.914	 	1.041	 	
 	 	2.0 g/t	 	9.464	 	0.769	 	0.156	 	0.097	 	0.063	 	0.176	 	3.684	 	1.121	 	
 	 	1.8 g/t	 	10.807	 	0.720	 	0.145	 	0.091	 	0.059	 	0.167	 	3.460	 	1.202	 	
 	 	1.6 g/t	 	12.476	 	0.669	 	0.134	 	0.085	 	0.054	 	0.158	 	3.225	 	1.294	 	
 	 	1.4 g/t	 	14.140	 	0.625	 	0.125	 	0.079	 	0.051	 	0.149	 	3.020	 	1.373	 	
 	 	1.2 g/t	 	15.969	 	0.583	 	0.116	 	0.073	 	0.048	 	0.140	 	2.824	 	1.450	 	
 	 	1.0 g/t	 	17.718	 	0.546	 	0.108	 	0.068	 	0.045	 	0.132	 	2.653	 	1.511	 	
 	 	0.8 g/t	 	19.207	 	0.517	 	0.103	 	0.065	 	0.043	 	0.126	 	2.518	 	1.555	 	
 	 	0.1 g/t	 	20.978	 	0.483	 	0.096	 	0.060	 	0.040	 	0.118	 	2.358	 	1.590	 	
IND	 	5.0 g/t	 	14.375	 	1.352	 	0.289	 	0.155	 	0.108	 	0.268	 	6.222	 	2.875	 	
 	 	4.0 g/t	 	25.834	 	1.177	 	0.249	 	0.140	 	0.094	 	0.239	 	5.455	 	4.531	 	
 	 	3.0 g/t	 	38.501	 	1.024	 	0.215	 	0.123	 	0.083	 	0.216	 	4.809	 	5.953	 	
 	 	2.8 g/t	 	40.961	 	0.997	 	0.209	 	0.120	 	0.081	 	0.212	 	4.694	 	6.182	 	
 	 	2.6 g/t	 	43.604	 	0.969	 	0.203	 	0.117	 	0.079	 	0.208	 	4.573	 	6.411	 	
 	 	2.4 g/t	 	46.605	 	0.939	 	0.196	 	0.114	 	0.076	 	0.203	 	4.439	 	6.652	 	
 	 	2.2 g/t	 	49.379	 	0.912	 	0.190	 	0.110	 	0.074	 	0.199	 	4.319	 	6.856	 	
 	 	2.0 g/t	 	52.604	 	0.882	 	0.183	 	0.107	 	0.072	 	0.194	 	4.183	 	7.074	 	
 	 	1.8 g/t	 	56.585	 	0.846	 	0.175	 	0.103	 	0.069	 	0.187	 	4.022	 	7.317	 	
 	 	1.6 g/t	 	61.369	 	0.806	 	0.166	 	0.098	 	0.066	 	0.180	 	3.841	 	7.579	 	
 	 	1.4 g/t	 	65.854	 	0.771	 	0.159	 	0.094	 	0.063	 	0.174	 	3.682	 	7.796	 	
 	 	1.2 g/t	 	70.042	 	0.740	 	0.152	 	0.090	 	0.060	 	0.168	 	3.539	 	7.970	 	
 	 	1.0 g/t	 	73.822	 	0.713	 	0.146	 	0.087	 	0.058	 	0.162	 	3.414	 	8.103	 	
 	 	0.8 g/t	 	76.876	 	0.691	 	0.142	 	0.084	 	0.057	 	0.158	 	3.315	 	8.192	 	
 	 	0.1 g/t	 	80.791	 	0.663	 	0.136	 	0.081	 	0.054	 	0.152	 	3.182	 	8.265	 	
INF	 	5.0 g/t	 	5.489	 	1.339	 	0.288	 	0.154	 	0.113	 	0.264	 	6.290	 	1.110	 	
 	 	4.0 g/t	 	10.418	 	1.150	 	0.244	 	0.141	 	0.097	 	0.234	 	5.449	 	1.825	 	
 	 	3.0 g/t	 	17.005	 	0.987	 	0.210	 	0.122	 	0.083	 	0.204	 	4.693	 	2.566	 	
 	 	2.8 g/t	 	18.183	 	0.962	 	0.205	 	0.118	 	0.081	 	0.199	 	4.577	 	2.676	 	
 	 	2.6 g/t	 	19.081	 	0.943	 	0.201	 	0.117	 	0.079	 	0.196	 	4.489	 	2.754	 	
 	 	2.4 g/t	 	20.192	 	0.920	 	0.196	 	0.114	 	0.077	 	0.192	 	4.380	 	2.843	 	
 	 	2.2 g/t	 	21.381	 	0.896	 	0.191	 	0.110	 	0.075	 	0.188	 	4.265	 	2.932	 	
 	 	2.0 g/t	 	22.036	 	0.882	 	0.188	 	0.109	 	0.074	 	0.185	 	4.201	 	2.976	 	
 	 	1.8 g/t	 	22.902	 	0.864	 	0.184	 	0.106	 	0.072	 	0.182	 	4.114	 	3.029	 	
 	 	1.6 g/t	 	23.780	 	0.846	 	0.180	 	0.104	 	0.071	 	0.178	 	4.025	 	3.077	 	
 	 	1.4 g/t	 	24.590	 	0.829	 	0.176	 	0.102	 	0.069	 	0.174	 	3.942	 	3.117	 	
 	 	1.2 g/t	 	25.180	 	0.815	 	0.173	 	0.101	 	0.068	 	0.172	 	3.880	 	3.141	 	
 	 	1.0 g/t	 	25.581	 	0.806	 	0.172	 	0.099	 	0.067	 	0.170	 	3.837	 	3.155	 	
 	 	0.8 g/t	 	25.960	 	0.797	 	0.170	 	0.098	 	0.067	 	0.168	 	3.794	 	3.167	 	
 	 	0.1 g/t	 	26.130	 	0.792	 	0.169	 	0.098	 	0.066	 	0.167	 	3.773	 	3.170	 	
 
198

 
AHMAVAARA WEST PDEQ

MEAS	 	5.0 g/t	 	5.686	 	1.642	 	0.308	 	0.201	 	0.151	 	0.365	 	8.040	 	1.470	 	
 	 	4.0 g/t	 	8.054	 	1.440	 	0.277	 	0.173	 	0.130	 	0.313	 	6.990	 	1.810	 	
 	 	3.0 g/t	 	10.724	 	1.271	 	0.251	 	0.150	 	0.112	 	0.272	 	6.112	 	2.107	 	
 	 	2.8 g/t	 	11.417	 	1.231	 	0.243	 	0.146	 	0.109	 	0.264	 	5.917	 	2.172	 	
 	 	2.6 g/t	 	12.240	 	1.187	 	0.235	 	0.141	 	0.104	 	0.255	 	5.702	 	2.244	 	
 	 	2.4 g/t	 	12.993	 	1.149	 	0.227	 	0.136	 	0.101	 	0.247	 	5.517	 	2.305	 	
 	 	2.2 g/t	 	13.664	 	1.115	 	0.221	 	0.132	 	0.098	 	0.240	 	5.358	 	2.354	 	
 	 	2.0 g/t	 	14.556	 	1.073	 	0.213	 	0.127	 	0.094	 	0.232	 	5.158	 	2.414	 	
 	 	1.8 g/t	 	15.621	 	1.025	 	0.203	 	0.121	 	0.090	 	0.222	 	4.935	 	2.478	 	
 	 	1.6 g/t	 	16.611	 	0.984	 	0.195	 	0.116	 	0.086	 	0.214	 	4.742	 	2.532	 	
 	 	1.4 g/t	 	17.985	 	0.931	 	0.185	 	0.110	 	0.082	 	0.203	 	4.494	 	2.599	 	
 	 	1.2 g/t	 	19.089	 	0.892	 	0.177	 	0.106	 	0.078	 	0.195	 	4.310	 	2.645	 	
 	 	1.0 g/t	 	20.479	 	0.846	 	0.168	 	0.100	 	0.075	 	0.185	 	4.093	 	2.695	 	
 	 	0.8 g/t	 	22.137	 	0.796	 	0.159	 	0.094	 	0.070	 	0.174	 	3.854	 	2.743	 	
 	 	0.1 g/t	 	24.493	 	0.730	 	0.147	 	0.087	 	0.064	 	0.160	 	3.540	 	2.788	 	
IND	 	5.0 g/t	 	17.802	 	1.679	 	0.329	 	0.210	 	0.145	 	0.338	 	7.904	 	4.524	 	
 	 	4.0 g/t	 	24.052	 	1.504	 	0.302	 	0.186	 	0.127	 	0.297	 	7.019	 	5.428	 	
 	 	3.0 g/t	 	29.832	 	1.363	 	0.279	 	0.167	 	0.114	 	0.265	 	6.337	 	6.078	 	
 	 	2.8 g/t	 	30.902	 	1.338	 	0.274	 	0.164	 	0.112	 	0.260	 	6.218	 	6.177	 	
 	 	2.6 g/t	 	31.721	 	1.318	 	0.270	 	0.161	 	0.111	 	0.256	 	6.127	 	6.248	 	
 	 	2.4 g/t	 	32.698	 	1.295	 	0.266	 	0.159	 	0.109	 	0.252	 	6.019	 	6.327	 	
 	 	2.2 g/t	 	33.464	 	1.277	 	0.262	 	0.156	 	0.107	 	0.248	 	5.934	 	6.384	 	
 	 	2.0 g/t	 	34.604	 	1.250	 	0.257	 	0.153	 	0.105	 	0.243	 	5.807	 	6.461	 	
 	 	1.8 g/t	 	35.600	 	1.225	 	0.252	 	0.150	 	0.103	 	0.238	 	5.698	 	6.522	 	
 	 	1.6 g/t	 	36.912	 	1.193	 	0.246	 	0.146	 	0.100	 	0.233	 	5.556	 	6.593	 	
 	 	1.4 g/t	 	38.201	 	1.164	 	0.240	 	0.142	 	0.098	 	0.227	 	5.419	 	6.655	 	
 	 	1.2 g/t	 	39.626	 	1.131	 	0.233	 	0.138	 	0.095	 	0.221	 	5.271	 	6.715	 	
 	 	1.0 g/t	 	40.995	 	1.101	 	0.227	 	0.134	 	0.093	 	0.215	 	5.131	 	6.763	 	
 	 	0.8 g/t	 	42.166	 	1.075	 	0.222	 	0.131	 	0.090	 	0.210	 	5.014	 	6.797	 	
 	 	0.1 g/t	 	43.723	 	1.041	 	0.215	 	0.127	 	0.088	 	0.204	 	4.855	 	6.825	 	
INF	 	5.0 g/t	 	3.092	 	1.694	 	0.377	 	0.174	 	0.127	 	0.253	 	7.193	 	0.715	 	
 	 	4.0 g/t	 	4.858	 	1.447	 	0.324	 	0.156	 	0.108	 	0.225	 	6.205	 	0.969	 	
 	 	3.0 g/t	 	6.583	 	1.275	 	0.286	 	0.140	 	0.095	 	0.204	 	5.512	 	1.167	 	
 	 	2.8 g/t	 	6.790	 	1.256	 	0.282	 	0.138	 	0.094	 	0.201	 	5.432	 	1.186	 	
 	 	2.6 g/t	 	6.922	 	1.243	 	0.279	 	0.136	 	0.093	 	0.199	 	5.380	 	1.197	 	
 	 	2.4 g/t	 	7.256	 	1.213	 	0.272	 	0.132	 	0.091	 	0.195	 	5.248	 	1.224	 	
 	 	2.2 g/t	 	7.607	 	1.180	 	0.265	 	0.129	 	0.089	 	0.190	 	5.112	 	1.250	 	
 	 	2.0 g/t	 	7.796	 	1.162	 	0.260	 	0.127	 	0.087	 	0.187	 	5.040	 	1.263	 	
 	 	1.8 g/t	 	8.094	 	1.133	 	0.254	 	0.124	 	0.085	 	0.183	 	4.924	 	1.281	 	
 	 	1.6 g/t	 	8.424	 	1.104	 	0.247	 	0.121	 	0.083	 	0.179	 	4.797	 	1.299	 	
 	 	1.4 g/t	 	8.712	 	1.077	 	0.241	 	0.118	 	0.082	 	0.175	 	4.687	 	1.313	 	
 	 	1.2 g/t	 	8.986	 	1.053	 	0.236	 	0.115	 	0.080	 	0.171	 	4.584	 	1.324	 	
 	 	1.0 g/t	 	9.176	 	1.035	 	0.232	 	0.113	 	0.079	 	0.169	 	4.511	 	1.331	 	
 	 	0.8 g/t	 	9.238	 	1.030	 	0.230	 	0.113	 	0.078	 	0.168	 	4.487	 	1.333	 	
 	 	0.1 g/t	 	9.339	 	1.021	 	0.228	 	0.112	 	0.077	 	0.166	 	4.446	 	1.335	 	
 
199

Exhibit 23.1
 
October 25, 2007

To:	Alberta Securities Commission	
 	British Columbia Securities Commission	
 	Manitoba Securities Commission	
 	New Brunswick Securities Commission	
 	Department of Government Services, Newfoundland and Labrador	
 	Nova Scotia Securities Commission	
 	Ontario Securities Commission	
 	Securities Office, Prince Edward Island	
 	Autorité des marchés financiers, Québec	
 	Saskatchewan Financial Services Commission	
 
I, Fred Brown, with a business address of Main St., Lynden, WA, 98264 and with a postal address of PO Box 332, Lynden, WA, 98264, USA, am the qualified person who is responsible for preparing for North American Palladium Ltd. (the “Corporation”) portions of a technical report dated October 29, 2007 and entitled “Technical Report, Mineral Resource Estimate and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” (the “Technical Report”). I hereby:
 
(a)                       consent to the public filing of the Technical Report and to extract from, or a summary of, the Technical Report in the written disclosure of the Corporation, being the press release dated October 31, 2007 issued in connection with the filing of the Technical Report; and
 
(b)                      confirm that I have read the written disclosure being filed and that it fairly and accurately represents the information in the Technical Report that supports the disclosure.
 
Dated this 25 th day of October, 2007.

/s/ Fred H. Brown	 	
 	
Fred H. Brown,	
CPG, Pr. Sci. Nat.	

 
Exhibit 23.2
 
October 29, 2007

To:	Alberta Securities Commission	
 	British Columbia Securities Commission	
 	Manitoba Securities Commission	
 	New Brunswick Securities Commission	
 	Department of Government Services, Newfoundland and Labrador	
 	Nova Scotia Securities Commission	
 	Ontario Securities Commission	
 	Securities Office, Prince Edward Island	
 	Autorité des marchés financiers, Québec	
 	Saskatchewan Financial Services Commission	
 
I, Eugene Puritch, residing at 44 Turtlecreek Blvd,, Brampton, Ontario L6W 3X7, am the qualified person who is responsible for preparing for North American Palladium Ltd. (the “Corporation”) portions of a technical report dated October 29, 2007 and entitled “Technical Report, Mineral Resource Estimate and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” (the “Technical Report”). I hereby:
 
(a)                       consent to the public filing of the Technical Report and to extract from, or a summary of, the Technical Report in the written disclosure of the Corporation, being the press release dated October 31, 2007 issued in connection with the filing of the Technical Report; and
 
(b)                      confirm that I have read the written disclosure being filed and that it fairly and accurately represents the information in the Technical Report that supports the disclosure.
 
Dated this 29 th day of October, 2007.

/s/ Eugene Puritch	 	
 	
Eugene Puritch, P.Eng.	

 
Exhibit 23.3
 
October 29, 2007

To:	Alberta Securities Commission	
 	British Columbia Securities Commission	
 	Manitoba Securities Commission	
 	New Brunswick Securities Commission	
 	Department of Government Services, Newfoundland and Labrador	
 	Nova Scotia Securities Commission	
 	Ontario Securities Commission	
 	Securities Office, Prince Edward Island	
 	Autorité des marchés financiers, Québec	
 	Saskatchewan Financial Services Commission	
 
I, Wayne D. Ewert, residing at 10 Langford Court, Brampton, Ontario L6W 4K4, am the qualified person who is responsible for preparing for North American Palladium Ltd. (the “Corporation”) portions of a technical report dated October 29, 2007 and entitled “Technical Report, Mineral Resource Estimate and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” (the “Technical Report”). I hereby:
 
(a)                       consent to the public filing of the Technical Report and to extract from, or a summary of, the Technical Report in the written disclosure of the Corporation, being the press release dated October 31, 2007 issued in connection with the filing of the Technical Report; and
 
(b)                      confirm that I have read the written disclosure being filed and that it fairly and accurately represents the information in the Technical Report that supports the disclosure.
 
Dated this 29 th day of October, 2007.

/s/ Wayne D. Ewert	 	
 	
Wayne D. Ewert, P.Geo.	

 
Exhibit 23.4
 
October 31, 2007

To:	Alberta Securities Commission	
 	British Columbia Securities Commission	
 	Manitoba Securities Commission	
 	New Brunswick Securities Commission	
 	Department of Government Services, Newfoundland and Labrador	
 	Nova Scotia Securities Commission	
 	Ontario Securities Commission	
 	Securities Office, Prince Edward Island	
 	Autorité des marchés financiers, Québec	
 	Saskatchewan Financial Services Commission	
 
I, Jason Rickard, residing at 417 Strand Ave., Thunder Bay, Ontario P7B 5A8, am the qualified person who is responsible for preparing for North American Palladium Ltd. (the “Corporation”) portions of a technical report dated October 29, 2007 and entitled “Technical Report, Mineral Resource Estimate and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” (the “Technical Report”). I hereby:
 
(a)                       consent to the public filing of the Technical Report and to extract from, or a summary of, the Technical Report in the written disclosure of the Corporation, being the press release dated October 31, 2007 issued in connection with the filing of the Technical Report; and
 
(b)                      confirm that I have read the written disclosure being filed and that it fairly and accurately represents the information in the Technical Report that supports the disclosure.
 
Dated this 31 st day of October, 2007.

/s/ Jason Rickard	 	
 	
Jason Rickard, P.Geo.	

 
Exhibit 23.5
 
October 31, 2007

To:	Alberta Securities Commission	
 	British Columbia Securities Commission	
 	Manitoba Securities Commission	
 	New Brunswick Securities Commission	
 	Department of Government Services, Newfoundland and Labrador	
 	Nova Scotia Securities Commission	
 	Ontario Securities Commission	
 	Securities Office, Prince Edward Island	
 	Autorité des marchés financiers, Québec	
 	Saskatchewan Financial Services Commission	
 
I, David King, residing at Unit 10, 102 Bronte Road, Oakville, Ontario L6L 6J5, am the qualified person who is responsible for preparing for North American Palladium Ltd. (the “Corporation”) portions of a technical report dated October 29, 2007 and entitled “Technical Report, Mineral Resource Estimate and Preliminary Economic Assessment (Scoping Study) of the Suhanko Project, Northern Finland” (the “Technical Report”). I hereby:
 
(a)                       consent to the public filing of the Technical Report and to extract from, or a summary of, the Technical Report in the written disclosure of the Corporation, being the press release dated October 31, 2007 issued in connection with the filing of the Technical Report; and
 
(b)                      confirm that I have read the written disclosure being filed and that it fairly and accurately represents the information in the Technical Report that supports the disclosure.
 
Dated this 31 st day of October, 2007.

/s/ David King	 	
 	
David King, P.Eng.